Critical Mineral Resources of the United States—

Economic and Environmental Geology and 

Prospects for Future Supply

Professional Paper 1802

U.S. Department of the Interior

U.S. Geological Survey
Periodic Table of Elements

1A 8A

2A

3B 4B 5B 6B 7B 8B 11B 12B

3A 4A 5A 6A 7A

element names in blue are liquids at room temperature

element names in red are gases at room temperature

element names in black are solids at room temperature

11

1

3 4

12

19 20 21 22 23 24 25 26 27 28 29 30

37 38 39 40 41 42 43 44 45 46 47 48

55 56

58 59 60

72 73 74 75 76 77 78 79 80

87 88

90

57

89 91 92 93 94 95 96

104 105 106 107 108 109 110 111 112

61 62 63 64 65 66 67

97 98 99

68 69 70 71

100 101 102 103

31

13 14 15 16 17 18

32 33 34 35 36

49 50 51 52 53 54

81 82 83 84 85 86

5 6 7 8 9 10

2

113 114 115 116 117 118

39.10

9.012

24.31

40.08 44.96 47.88 50.94 52.00 54.94 55.85 58.93 58.69 63.55 65.39

10.81

26.98

12.01

28.09

14.01

69.72 72.64

30.97

74.92

16.00 19.00 20.18

4.003

32.06 35.45 39.95

78.96 79.90 83.79

85.47

132.9

87.62

137.3

88.91 91.22

178.5

92.91

180.9

95.96 

183.9

(98)

186.2

101.1

190.2

102.9 106.4

195.1

107.9 112.4 114.8 118.7 121.8

(209) (210) (222)

127.6 126.9 131.3

192.2 197.0 200.5 204.4 207.2 209.0

(223) (226) (265) (268) (271) (270) (277) (276) (281) (280) (285)

140.1 140.9 144.2 (145) 150.4 152.0 157.2 158.9 162.5 164.9 167.3 168.9 173.0 175.0

232

138.9

(227) 231 238 (237) (244) (243) (247) (247) (251) (252) (257) (258) (259) (262)

1.008

6.94

22.99

(284) (289) (288) (293) (294) (294)

hydrogen

barium

francium radium

strontium

sodium

vanadium

lithium beryllium

magnesium

potassium calcium

rubidium

cesium

helium

boron carbon nitrogen oxygen fluorine neon

aluminum silicon phosphorus sulfur chlorine argon

scandium titanium chromium manganese iron cobalt nickel copper zinc gallium germanium arsenic selenium bromine krypton

yttrium zirconium niobium molybdenum technetium ruthenium rhodium palladium silver cadmium indium tin antimony tellurium iodine xenon

hafnium

cerium praseodymium neodymium promethium samarium europium gadolinium terbium dysprosium holmium erbium thulium ytterbium lutetium

tantalum tungsten rhenium osmium iridium platinum gold mercury thallium lead bismuth polonium astatine radon

thorium

lanthanum

actinium protactinium uranium neptunium plutonium americium curium berkelium californium einsteinium fermium mendelevium nobelium lawrencium 

rutherfordium dubnium seaborgium bohrium hassium meitnerium darmstadtium roentgenium copernicium flerovium livermorium

Lanthanide Series*

Actinide Series**

Modified from Los Alamos National Laboratory Chemistry Division; available at http://periodic.lanl.gov/images/periodictable.pdf.

Front cover (clockwise starting at top left).

Dry Lake Wind Power Project. This 63-megawatt-capacity facility in Arizona was the first utility-scale power project. The large wind turbines 

use generators that contain strong permanent magnets composed of neodymium-iron-boron. Photograph courtesy of Iberdrola Renewables, 

Inc., NREL 16701.

U.S. Air Force F-35 Lightning II Joint Strike Fighter. This and many other aerospace vehicles rely on electrical and mechanical components 

made of beryllium alloys. Photograph courtesy of the U.S. Air Force.

Flat-panel display and touchscreen devices. The screens are coated with indium-tin oxide, for which there are few chemical substitutes. 

Photograph by C.N. Mercer, U.S. Geological Survey.

Solar photovoltaic cells. This 2-megawatt ground-mounted array occupies the site of a former landfill at Fort Carson, Colorado. Tellurium is a 

critical component for the development of efficient thin-film photovoltaic cells that are needed for the production of electricity from sunlight. 

Photograph courtesy of the U.S. Department of Energy Western Area Power Administration.

Back cover. A handful of stibnite, which is the primary ore mineral for antimony. Antimony is widely used in plastics, rubbers, paints, 

and textiles, including industrial safety suits and some children’s clothing, to make them resistant to the spread of flames. Photograph by 

Niki Wintzer, U.S. Geological Survey.

Front and back covers background image. Lithium-brine evaporating ponds at Clayton Valley, Nevada. Lithium has many uses, the most 

prominent being in batteries for cell phones, laptop computers, and electric and hybrid vehicles. Photograph by Doc Searls/CC-BY-2.0, 

http://commons.wikimedia.org/wiki/File:Chemetall_Foote_Lithium_Operation.jpg.
Critical Mineral Resources of the 

United States—Economic and 

Environmental Geology and 

Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II,

and Dwight C. Bradley

Professional Paper 1802

U.S. Department of the Interior

U.S. Geological Survey
U.S. Department of the Interior

RYAN K. ZINKE, Secretary

U.S. Geological Survey

William H. Werkheiser, Acting Director

U.S. Geological Survey, Reston, Virginia: 2017

For more information on the USGS—the Federal source for science about the Earth, its natural and living 

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For an overview of USGS information products, including maps, imagery, and publications, 

visit https://store.usgs.gov/.

Any use of trade, firm, or product names is for descriptive purposes only and does not imply endorsement by the 

U.S. Government.

Although this information product, for the most part, is in the public domain, it also may contain copyrighted materials 

as noted in the text. Permission to reproduce copyrighted items must be secured from the copyright owner.

Suggested citation:

Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., 2017, Critical mineral resources of the 

United States—Economic and environmental geology and prospects for future supply: U.S. Geological Survey 

Professional Paper 1802, 797 p., https://doi.org/10.3133/pp1802.

Library of Congress Cataloging-in-Publication Data

Names: Schulz, K. J., editor. | Geological Survey (U.S.), issuing body.

Title: Critical mineral resources of the United States : economic and

 environmental geology and prospects for future supply / edited by Klaus J.

 Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley.

Other titles: Economic and environmental geology and prospects for future

 supply | U.S. Geological Survey professional paper ; 1802.

Description: Reston, Virginia : U.S. Geological Survey, [2017] | Series:

 Professional paper ; 1802 | Includes bibliographical references.

Identifiers: LCCN 2017005489| ISBN 9781411339910 | ISBN 1411339916

Subjects: LCSH: Mines and mineral resources--United States. | Strategic

 materials--United States. | Industrial minerals--United States.

Classification: LCC TN23 .C67 2017 | DDC 333.8/50973--dc23 | SUDOC I

 19.16:1802

LC record available at https://lccn.loc.gov/2017005489

ISSN 1044-9612 (print)

ISSN 2330-7102 (online)
iii

Foreword

From the Stone Age to the present, mineral commodities have been essential ingredients for 

building and advancing civilization. Products built with materials derived from mineral resources 

include homes and office buildings; cars and roads; computers, televisions, and smart phones; 

and jet fighters and other military hardware needed to defend the Nation. In short, minerals are 

essential to advance and protect modern society. 

When the periodic table of elements was first established in the latter half of the 19th century, 

many of the elements were known to exist in nature, but relatively few were being used by 

society. Today, discovery of new uses for an increasing number of elements is enabling rapid 

innovations in technology and materials science. Advances in telecommunications, information 

technology, health care, energy production, and national defense systems have all been possible 

through the use of new mineral materials. 

As the importance and dependence of specific mineral commodities increase, so does 

concern about their supply. The United States is currently 100 percent reliant on foreign 

sources for 20 mineral commodities and imports the majority of its supply of more than 50 

mineral commodities. Mineral commodities that have important uses and face potential 

supply disruption are critical to American economic and national security. However, a mineral 

commodity’s importance and the nature of its supply chain can change with time; a mineral 

commodity that may not have been considered critical 25 years ago may be critical today, and 

one considered critical today may not be so in the future. 

The U.S. Geological Survey has produced this volume to describe a select group of mineral 

commodities currently critical to our economy and security. For each mineral commodity 

covered, the authors provide a comprehensive look at (1) the commodity’s use; (2) the geology 

and global distribution of the mineral deposit types that account for the present and possible 

future supply of the commodity; (3) the current status of production, reserves, and resources in 

the United States and globally; and (4) environmental considerations related to the commodity’s 

production from different types of mineral deposits. The volume describes U.S. critical mineral 

resources in a global context, for no country can be self-sufficient for all its mineral commodity 

needs, and the United States will always rely on global mineral commodity supply chains. This 

volume provides the scientific understanding of critical mineral resources required for informed 

decisionmaking by those responsible for ensuring that the United States has a secure and 

sustainable supply of mineral commodities.

The Nation was largely built on the products produced from its mineral deposits. The future 

will also be built on a foundation of minerals, many of which will continue to be discovered and 

produced from across the country.

Ryan K. Zinke

Secretary of the Interior

v

Preface

Mineral commodities are vital for economic growth, improving the quality of life, providing 

for national defense, and the overall functioning of modern society. Minerals are being used 

in larger quantities than ever before and in an increasingly diverse range of applications— 

from telecommunications (cell phones and computers), to renewable-energy generation (wind 

turbines, solar photovoltaics, and fuel cells), to clean forms of transportation (electric and hybrid 

cars). Until the mid-20th century, only about 15 metallic elements had much practical use. 

Today, nearly all the natural elements in the periodic table of elements have several significant 

uses. For example, the manufacture of a modern computer chip requires more than one-half of 

the elements in the periodic table. Even though many of the elements may be present in only 

small amounts, each is essential to the function and performance of the chip.

With the increasing demand for a considerably more diverse suite of mineral commodities 

has come renewed recognition that competition and conflict over mineral resources can pose 

significant risks to the manufacturing industries that depend on them. In addition, although 

mineral deposits may occur in many places around the world, production of many mineral 

commodities has become concentrated in relatively few countries (for example, tungsten, 

rare-earth elements, and antimony in China; niobium in Brazil; and platinum-group elements 

in South Africa and Russia), thus increasing the risk for supply disruption owing to political, 

social, or other factors. At the same time, an increasing awareness and sensitivity to potential 

environmental and health issues caused by the mining and processing of many mineral 

commodities may place additional restrictions on mineral supplies. These factors have led a 

number of Governments, including the Government of the United States, to attempt to identify 

those mineral commodities that are viewed as most “critical” to the national economy and (or) 

security if supplies should be curtailed. The lists of critical minerals compiled by Governments 

and other organizations vary in the number and individual rankings of mineral commodities 

included on them, but many of the lists include several of the same commodities. Rare-earth 

elements and platinum-group elements particularly are broadly viewed as critical.

This book presents resource and geologic information on the following 23 mineral commodities 

currently among those viewed as important to the national economy and national security of the 

United States: antimony (Sb), barite (barium, Ba), beryllium (Be), cobalt (Co), fluorite or fluorspar 

(fluorine, F), gallium (Ga), germanium (Ge), graphite (carbon, C), hafnium (Hf), indium (In), lithium (Li),

manganese (Mn), niobium (Nb), platinum-group elements (PGE), rare-earth elements (REE), 

rhenium (Re), selenium (Se), tantalum (Ta), tellurium (Te), tin (Sn), titanium (Ti), vanadium (V), and 

zirconium (Zr). For a number of these commodities–for example, graphite, manganese, niobium, 

and tantalum—the United States is currently wholly dependent on imports to meet its needs.

The first two chapters deal with general information pertinent to the study of mineral resources. 

The introductory chapter (A) discusses the purposes of the volume, the distinctions between 

reserves and various categories of resources, and issues related to the classification of mineral 

resource “criticality.” The second chapter (B) provides an overview of some of the environmental 

considerations related to the mining of nonfuel mineral resources, including the modern 

regulatory framework and development of geoenvironmental mineral-deposit models.
vi

Chapters C through V describe individual mineral commodities and include an overview of 

current uses of the commodity, identified resources and their distribution nationally and 

globally, the state of current geologic knowledge, the potential for finding additional deposits 

nationally and globally, and geoenvironmental issues that may be related to the production and 

uses of the commodity. These chapters are updates of the commodity chapters published in 

1973 in U.S. Geological Survey Professional Paper 820, “United States Mineral Resources.” In 

1973, many of these commodities were only of minor importance, and resource and geologic 

information was often very limited and incomplete. In addition, little was generally known 

about geoenvironmental issues related to their production and use.

We would like to thank our colleagues for their contributions to and cooperation in all phases 

of the preparation of this book. The descriptions of geology, the origin of mineral deposits, 

and geoenvironmental chemistry in each chapter necessarily involve some scientific jargon, 

but much of the discussion is cast in less-technical language. Our hope is that the information 

provided will be of use to scientists and nonscientists alike.

Klaus J. Schulz

John H. DeYoung, Jr.

Robert R. Seal II

Dwight C. Bradley
vii

Contents

Foreword........................................................................................................................................................iii

Preface ............................................................................................................................................................v

Chapter A. Critical Mineral Resources of the United States—An Introduction

By Klaus J. Schulz, John H. DeYoung, Jr., Dwight C. Bradley, and Robert R. Seal II................A1

Chapter B. Environmental Considerations Related to Mining of Nonfuel Minerals

By Robert R. Seal II, Nadine M. Piatak, Bryn E. Kimball, and Jane M. Hammarstrom.............B1

Chapter C. Antimony

By Robert R. Seal II, Klaus J. Schulz, and John H. DeYoung, Jr.

With contributions from David M. Sutphin, Lawrence J. Drew, James F. Carlin, Jr.,

and Byron R. Berger............................................................................................................................C1

Chapter D. Barite (Barium)

By Craig A. Johnson, Nadine M. Piatak, and M. Michael Miller .................................................D1

Chapter E. Beryllium

By Nora K. Foley, Brian W. Jaskula, Nadine M. Piatak, and Ruth F. Schulte ............................. E1

Chapter F. Cobalt

By John F. Slack, Bryn E. Kimball, and Kim B. Shedd......................................................................F1

Chapter G. Fluorine

By Timothy S. Hayes, M. Michael Miller, Greta J. Orris, and Nadine M. Piatak........................G1

Chapter H. Gallium

By Nora K. Foley, Brian W. Jaskula, Bryn E. Kimball, and Ruth F. Schulte .................................H1

Chapter I. Germanium and Indium

By W.C. Pat Shanks III, Bryn E. Kimball, Amy C. Tolcin, and David E. Guberman....................... I1

Chapter J. Graphite

By Gilpin R. Robinson, Jr., Jane M. Hammarstrom, and Donald W. Olson ................................. J1

Chapter K. Lithium

By Dwight C. Bradley, Lisa L. Stillings, Brian W. Jaskula, LeeAnn Munk, 

and Andrew D. McCauley...................................................................................................................K1

Chapter L. Manganese

By William F. Cannon, Bryn E. Kimball, and Lisa A. Corathers ......................................................L1

Chapter M. Niobium and Tantalum

By Klaus J. Schulz, Nadine M. Piatak, and John F. Papp .............................................................M1

Chapter N. Platinum-Group Elements

By Michael L. Zientek, Patricia J. Loferski, Heather L. Parks, Ruth F. Schulte, 

and Robert R. Seal II............................................................................................................................N1

Chapter O. Rare-Earth Elements

By Bradley S. Van Gosen, Philip L. Verplanck, Robert R. Seal II, Keith R. Long, 

and Joseph Gambogi...........................................................................................................................O1

Chapter P. Rhenium

By David A. John, Robert R. Seal II, and Désirée E. Polyak..........................................................P1

Chapter Q. Selenium

By Lisa L. Stillings.................................................................................................................................Q1
viii

Chapter R. Tellurium

By Richard J. Goldfarb, Byron R. Berger, Micheal W. George, and Robert R. Seal II............R1

Chapter S. Tin

By Robert J. Kamilli, Bryn E. Kimball, and James F. Carlin, Jr. ...................................................S1

Chapter T. Titanium

By Laurel G. Woodruff, George M. Bedinger, and Nadine M. Piatak........................................ T1

Chapter U. Vanadium

By Karen D. Kelley, Clinton T. Scott, Désirée E. Polyak, and Bryn E. Kimball...........................U1

Chapter V. Zirconium and Hafnium

By James V. Jones III, Nadine M. Piatak, and George M. Bedinger.........................................V1

Figures

A1. Diagram showing increases in the use of elements over two decades of 

computer chip technology development...............................................................................A3

A2. Diagram showing the mineral resource classification system used 

in this volume..............................................................................................................................A5

B1. Graph showing population growth and the change in supporting land area 

from 3500 B.C. to 2100 A.D., with projections to 2050 ..........................................................B3

B2. Graph showing dates associated with all the mine sites on the 

U.S. Environmental Protection Agency’s National Priorities List ....................................B10

C1. World map showing locations of selected antimony deposits, mines, 

and major occurrences ............................................................................................................C2

C2. Pie chart showing major end uses of antimony as a percentage of antimony

consumption in the United States in 2012..............................................................................C3

C3. Graph showing world production, U.S. apparent consumption, and U.S. mine

production of antimony from 1900 to 2012 .............................................................................C4

C4. Pie charts showing percentages of contained antimony in U.S. imports for the

period 2008–11, by source country.........................................................................................C5

D1. Graph showing barite world production, U.S. production, and U.S. consumption

from 1950 to 2011........................................................................................................................D2

D2. World map showing locations of selected barite deposits and districts,

color-coded by deposit type ....................................................................................................D4

D3. Pie chart showing average annual barite production for the period 2007–11, by

country and amount in thousand metric tons .......................................................................D7

E1. World map showing locations of selected deposits of beryllium by the two 

major beryllium-bearing mineral types .................................................................................. E2

E2. Photographs showing the minerals bertrandite, which can contain up to 

42 percent beryllium oxide, and industrial beryl, which can contain up to 

about 5 percent beryllium......................................................................................................... E3

E3. Photographs illustrating some of the many uses of beryllium........................................... E4

E4. Pie charts showing reported end uses of beryllium consumed in the 

United States in 2011 for the two main classes of products—performance 

alloys, and beryllium and composites .................................................................................... E6

E5. A generalized cross section showing the geologic setting of and some 

example deposits for the major types of beryllium resources associated 

with rare-metal magma systems........................................................................................... E12
ix

E6. Location map and simplified geologic map of the Hellroaring Creek prospect, 

which is located west of Kimberley, British Columbia, Canada....................................... E13

E7. Generalized geologic cross section showing the setting of volcanogenic 

beryllium deposits and related deposit types, gelogic map of Spor Mountain 

area in Utah, photograph of beryllium tuff at Spor Mountain, and photograph 

of nodule from the Spor Mountain tuff................................................................................. E15

E8. Graph showing general estimates of grade and tonnage for a variety of 

beryllium deposits, districts, and belts................................................................................. E17

E9. Graph showing estimated global production of beryllium for 2000–11 .......................... E18

F1. Pie chart showing major end uses of cobalt as a percentage of consumption

worldwide in 2011 .......................................................................................................................F2

F2. Graph showing world cobalt mine and refinery production and apparent

consumption from 2007 to 2011 ................................................................................................F2

F3. Graph showing world cobalt mine production from 1950 to 2011 ......................................F3

F4. Pie chart showing percentage of world cobalt mine production in 2011, 

by country .....................................................................................................................................F3

F5. Pie chart showing percentage of world cobalt mine production in 2011, 

by deposit type............................................................................................................................F10

F6. Grade-tonnage plot for 214 cobalt deposits worldwide ....................................................F11

F7. World map showing global distribution of major cobalt-bearing mineral 

deposits and selected smaller deposits that represent minor types...............................F12

F8. Pie charts showing proportions of cobalt contained in mineral deposits 

worldwide, by deposit type .....................................................................................................F13

G1. Photograph of a fluorite specimen from the Number 1 (Minerva) Mine,

Cave-in-Rock subdistrict, Illinois-Kentucky fluorspar district ...........................................G3

G2. Graph showing the solubility of fluorite as a function of temperature for 

complex Na-Ca-Mg-Cl brines from ambient temperatures to 260 °C.............................G10

G3. Chart showing eight minerals or mineral groups from which fluorine has been

produced or may be produced in the future and a preliminary classification of

hydrothermal fluorspar deposits by tectonic and magmatic association......................G11

G4. World map showing locations of selected fluorspar deposits according to 

their tectonic and magmatic class as listed in figure G3B...............................................G14

G5. Map showing the locations of deposits in the Cantabrian salt-related

carbonate-hosted mineral district in north-central Spain................................................G19

G6. Plot of fluorite grade versus tonnage for fluorspar deposits related to 

strongly differentiated granites, for carbonatite-related fluorspar deposits, 

and for veins from all classes of fluorspar deposits..........................................................G24

G7. Graph showing sources of U.S. fluorspar supply from 1970 to 2011...............................G29

H1. Photograph of gallium metal with an inset showing its position on the periodic 

table of elements .......................................................................................................................H1

H2. Photographs showing examples of some current uses for gallium..................................H3

H3. Charts showing U.S. and world gallium production and consumption from 2007 

to 2012..........................................................................................................................................H4

H4. World map showing locations of selected mineral deposits, by type; gallium 

has been produced from these types of deposits................................................................H5

H5. Cross sections showing the general geologic environments for the types of 

mineral deposits with which gallium is most commonly associated and from 

which gallium is typically extracted .......................................................................................H6
x

H6. Plot of the ratio of aluminum to gallium versus gallium content for bauxite 

deposits, volcanic rocks, and hydrothermally altered rocks of the McDermitt 

caldera, Nevada .......................................................................................................................H12

H7. Photographs showing samples of bauxite ore and sphalerite ore, which are the

primary mineralogical sources of gallium...........................................................................H14

I1. Photograph of a concentrator photovoltaic solar power system ...................................... I3

I2. Pie charts showing major end uses of germanium and indium as a percentage 

of world consumption in 2012 ................................................................................................... I3

I3. Photographs showing indium-tin oxide, which is a transparent conducting 

oxide, and examples of flat-panel display screens and touchscreens ............................. I4

I4. Map and schematic cross section showing the geology of the Red Dog mining 

district in Alaska and the stratigraphy of selected deposits in the district ...................... I7

I5. Map showing the locations and geologic settings of selected volcanogenic 

massive sulfide, sedimentary exhalative, Mississippi Valley-type, and coal 

deposits and other types of deposits in southern China...................................................... I9

I6. Map showing the location of Kipushi-type deposits (including the Kabwe 

deposits) and major Neoproterozoic orogenic belts and basins in the 

Precambrian tectonic framework of southern Africa ........................................................ I10

I7. Map showing indium-bearing tin-polymetallic ore deposits in Bolivia........................... I12

I8. Graph showing worldwide production of germanium and indium from 

1995 to 2012................................................................................................................................ I14

J1. Diagram showing the arrangement of carbon atoms in graphite ..................................... J6

J2. World map showing locations of major graphite deposits and districts in the 

world, by commodity type ...................................................................................................... J14

J3. Plot of grade and tonnage for some of the amorphous and crystalline graphite

deposits listed in table J3, by deposit type.......................................................................... J16

J4. Pie chart showing average annual natural graphite production for the period

2006–10, by country or region and amount.......................................................................... J17

K1. Graph showing world lithium production from 1900 to 2007, by deposit type 

and year.......................................................................................................................................K1

K2. Pie chart showing major end uses of lithium as a percentage of world 

consumption in 2013..................................................................................................................K2

K3. Photographs showing some sources and uses of lithium..................................................K3

K4. World map showing locations of selected lithium-cesium-tantalum pegmatites 

and lithium granites...................................................................................................................K6

K5. World map showing locations of selected closed-basin lithium-brine,

lithium-enriched oilfield brine, geothermal brine, lithium-clay, and 

lithium-zeolite deposits.............................................................................................................K7

K6. Histograms showing the broad correspondence in the age distributions of 

lithium-cesium-tantalum pegmatites, detrital zircons from modern river sands, 

and lithium resources in pegmatites......................................................................................K9

K7. Schematic cross section showing the concentric arrangement of

lithium-cesium-tantalum pegmatites around a parental granite pluton.........................K10

K8. Conceptual ore-deposit model for lithium brine.................................................................K11

K9. Plots of lithium grade and tonnage for selected world deposits.....................................K13

L1. Charts showing world production, apparent consumption, and distribution 

of production of manganese ore from 2007 to 2011 ..............................................................L3
xi

L2. Pie charts showing the sources and the annual average amounts of gross weight

of U.S. imports of manganese ore, ferromanganese, and silicomanganese for the 

period 2008–11.............................................................................................................................L3

L3. World map showing the location, relative size, and type of the major terrestrial 

manganese deposits listed in table L2 as well as subeconomic deposits in the 

United States ...............................................................................................................................L7

L4. Photographs showing examples of contrasting types of manganese ore .......................L8

L5. Schematic diagram of the oceanic conditions necessary to form sedimentary

manganese deposits that are not enriched in iron...............................................................L8

L6. Photograph showing a dense carpet of ferromanganese nodules on the 

seabed off Johnston Island within the United States Exclusive Economic Zone 

near Hawaii................................................................................................................................L10

L7. Photograph showing ferromanganese crust on carbonate rock from the Blake

Plateau off the southeastern coast of the United States...................................................L11

L8. Photograph of the Mamatwan open pit mine in South Africa ..........................................L13

L9. Graphs showing the cumulative frequency of tonnages and grades of 

39 marine sedimentary manganese deposits ......................................................................L14

M1. Photograph (central view) of the ATLAS detector in the Large Hadron 

Collider showing its eight superconducting barrel toroid magnets around 

the calorimeter..........................................................................................................................M3

M2. Pie charts showing percentage of reported world consumption of niobium 

and tantalum, by material produced.........................................................................................M3

M3. Criticality matrix for niobium, tantalum, and selected other mineral 

commodities ..............................................................................................................................M4

M4. Photographs showing centimeter-size pyrochlore crystals from Uganda, 

and a tantalite crystal ..............................................................................................................M6

M5. World map showing locations of selected niobium and tantalum mines, 

 deposits, and occurrences, by deposit type .......................................................................M8

M6. Log-log plots of deposit grades and tonnages of tantalum and niobium, by 

deposit type ...............................................................................................................................M9

M7. Diagrams showing the subsurface geology of the Saint-Honoré carbonatite 

complex in southern Quebec, Canada, and a schematic north-south crosssection along line A–A’ in A..................................................................................................M14

M8. Schematic cross-section of the Lovozero alkaline intrusion, Kola Peninsula, 

Russia, showing the relation among the three intrusive phases and the 

niobium mineralization contained in eudialyte and loparite............................................M17

M9. Schematic representation of regional lithium-cesium-tantalum rare-metalbearing pegmatite zoning above a parental granite.........................................................M18

M10. Schematic cross-section of a concentrically zoned lithium-cesium-tantalum 

rare-metal-bearing pegmatite ..............................................................................................M18

M11. Bar chart showing niobium resources and reserves in Brazil, Canada, and 

the United States ....................................................................................................................M20

M12. Bar chart showing global tantalum resources and reserves .........................................M21

M13. Bar charts showing global mine production of niobium and tantalum, from 

2000 to 2011...............................................................................................................................M23

M14. Pie charts showing percentage of average annual world production of niobium 

 and tantalum, for the period 2007–11, by country.................................................................M23
xii

N1. Graphs showing platinum and palladium consumption, by category of use, 

from 2000 to 2012 for the world, North America, and China...............................................N3

N2. Photograph of gold mask with platinum highlights, from the period of 

La Tolita culture, Ecuador.........................................................................................................N4

N3. World map showing locations of igneous intrusions and intrusive complexes 

that contain most of the world’s platinum-group-element deposits, as well as 

the placer deposits that are mentioned in the text..............................................................N7

N4. Schematic block diagram showing changes in the form of igneous intrusions 

with depth and the relative occurrence of conduit-type, contact-type, and 

reef-type magmatic ore deposits............................................................................................N9

N5. Map showing the geology of the Siberian flood basalt province in Russia, 

which is the largest flood basalt province in the world....................................................N10

N6. Map showing the geology of the Noril’sk-Talnakh area and the location of

nickel–copper–platinum-group-element deposits.............................................................N11

N7. Maps showing nickel–copper–platinum-group-element deposits in the 

Talnakh area, Russia ...............................................................................................................N13

N8. Photograph of copper-rich massive sulfide ore exposed in a stope in the 

Oktyabr’sk Mine in the Talnakh area, Russia ......................................................................N14

N9. Map showing the Rustenburg Layered Suite of the Bushveld Complex, 

South Africa, the surface trace of significant orebodies, and cross sections 

through the central area and northeastern limb................................................................N15

N10. Photograph of the UG2 Chromitite at the Karee Mine in the western part of 

the Bushveld Complex, South Africa....................................................................................N16

N11. Photograph of the base of the Merensky cyclic unit, a pegmatoidal pyroxenite, which 

contains the platinum-group-element-rich Merensky Reef.............................................N16

N12. Geologic map and cross sections of the Great Dyke, Zimbabwe ...................................N17

N13. Geologic map and cross section of the Stillwater Complex, Montana ..........................N19

N14. Photograph of the Stillwater Mine in south-central Montana, looking 

southeast...................................................................................................................................N20

N15. Map showing the geology along the western margin of the Duluth Complex,

Minnesota, with the surface projection of nickel–copper–platinum-groupelement deposits and exploration targets...........................................................................N21

N16. Maps illustrating the distribution of platinum deposits in the Ural 

Mountains, Russia ...................................................................................................................N24

N17. Geology and imagery of the Uralian-type Kondyor Massif, which is located in

eastern Siberia, Russia, north of the city of Khabarovsk..................................................N25

N18. Photograph and lithograph showing the morphology of platinum-iron-alloy 

nuggets derived from Uralian-type intrusions ....................................................................N26

N19. Maps illustrating platinum-group-element resources in southeastern Alaska............N27

N20. Graph showing world platinum-group-element production, by country and 

year, from 1960 to 2011............................................................................................................N28

N21. Pie chart showing world platinum-group-element production from 1960 to 2011, 

by country and amount, in metric tons.................................................................................N29

N22. Plot showing the relation between tonnage and grade of remaining resources 

for conduit-type, reef-type, and other types of deposits enriched in platinumgroup elements ........................................................................................................................N30

N23. Graphs showing the percent of contained metal against percent of deposits 

for the world’s platinum-group-element (PGE) and porphyry copper deposits and for 

the top 30 percent of the world’s PGE deposits..................................................................N31
xiii

N24. Graph showing contained platinum-group-element (PGE) and gold metal against 

the ratio of palladium to platinum for the major PGE deposits of the world..................N32

N25. Three-dimensional block diagram showing the Merensky Reef interpolated 

down to 2 kilometers in the southern area of the western limb of the Bushveld 

Complex, South Africa ............................................................................................................N33

N26. Graph illustrating the exposed area and stratigraphic thickness of cumulates 

in more than 200 intrusions from around the world ...........................................................N34

N27. Geologic map of the Amphitheater Mountains and south-central Alaska 

showing the location and names of mafic-ultramafic complexes that are 

part of the Nikolai large igneous province ..........................................................................N36

N28. Graphs showing platinum-group-element prices for platinum, palladium, 

rhodium, iridium, ruthenium, and osmium from 1880 to 2013............................................N44

O1. Graph showing world mine production of rare-earth oxides, by country and 

year, from 1960 to 2012..............................................................................................................O4

O2. Graph showing radii of the trivalent ions of the rare-earth elements as well 

as of cerium in the +4 valence state and europium in the +2 valence state ...................O6

O3. Photograph of the Mountain Pass Mine in San Bernardino County, California,

which was the only active producer of rare-earth elements in the United States 

in 2013 ..........................................................................................................................................O8

O4. World map showing locations of active or recently active rare-earth-element

mines and ongoing advanced exploration projects.............................................................O9

O5. Chondrite-normalized plot showing the rare-earth-element (REE) distribution in six

different types of North American REE deposits................................................................O11

P1. Photographs of rhenium and rhenium compounds .............................................................P2

P2. Pie chart showing major end uses of primary rhenium as a percentage of 

world consumption in 2012.......................................................................................................P3

P3. World map showing locations of major rhenium-bearing deposits listed in 

table P1........................................................................................................................................P5

P4. Plot of rhenium grade versus deposit tonnage for major rhenium-bearing 

deposits in the world, including those shown in figure P3 ...............................................P10

P5. Cross sections illustrating rhenium occurrences in major deposit types from 

which rhenium is recovered or potentially recoverable...................................................P11

P6. Pie chart showing world rhenium mine production in 2012, by country and 

percent of world total..............................................................................................................P18

Q1. Graph showing relative abundance of the chemical elements in Earth’s 

upper crust..................................................................................................................................Q2

Q2. Graph showing end uses for selenium in the United States from 1975 to 2012 ..............Q3

Q3. Graph illustrating a mass balance for selenium in the United States from 

1943 to 2015.................................................................................................................................Q5

Q4. Graph showing average annual prices of commercial-grade selenium from 

1970 to 2010.................................................................................................................................Q6

Q5. Map showing selenium concentrations in coal samples, by region of the 

United States ............................................................................................................................Q15

Q6. Maps showing locations of seleniferous sedimentary outcrops and deposits 

and plant samples with significant selenium content.......................................................Q18

Q7. A predictive map of selenium source rocks associated with organic-rich 

depositional marine basins ....................................................................................................Q22

Q8. Maps showing selenium concentrations in soils of the conterminous 

United States ............................................................................................................................Q24
xiv

R1. World map showing locations of selected tellurium-enriched mineral 

occurrences discussed in the text, by deposit type ............................................................R2

R2. Pie chart showing major end uses of tellurium as a percentage of world

consumption in 2010..................................................................................................................R3

R3. Phase diagram showing speciation calculations for tellurium in a hydrothermal

fluid at 300 degrees Celsius (°C), as a function of pH and oxygen fugacity.....................R4

R4. Pie chart showing percentage of estimated world tellurium refinery production 

in 2013, by country or region ....................................................................................................R13

S1. Pie chart showing major end uses of tin as a percentage of total consumption 

in the United States in 2014......................................................................................................S4

S2. Graph showing the average annual prices of tin metal from 1970 to 2010 ......................S5

S3. Ball-and-stick model of part of the crystal structure of cassiterite..................................S6

S4. Photograph of cassiterite crystals..........................................................................................S6

S5. Photograph of wood tin cassiterite ........................................................................................S6

S6. Schematic vertical section across a typical hydrothermal mineralized 

granite cupola showing salient features of a shallow granite-related 

tin-mineralized system..............................................................................................................S7

S7. World map showing locations of major tin deposits and districts in the world, 

by deposit type .........................................................................................................................S10

S8. Graph showing grades and resource tonnages of major tin deposits in the 

world, by deposit type or source...........................................................................................S11

S9. Pie chart showing percentage of world tin reserves in 2016, by country ............................S12

S10. Pie chart showing estimated world mine production of tin in 2015, by 

country and amount ................................................................................................................S12

S11. Map of Alaska showing the locations of selected tin deposits or occurrences ..........S15

T1. Chart showing common titanium-bearing oxide minerals and common 

titanium-bearing silicate minerals with their approximate titanium content .................. T5

T2. Reflected-light photographs showing lamellae of hematite in ilmenite in 

hemo-ilmenite from the magmatic Lac Tio hemo-ilmenite deposit; and 

titaniferous magnetite and ilmenite from the magmatic La Blache

iron-titanium-oxide deposit...................................................................................................... T5

T3. World map showing global distribution of titanium deposits, by size and type 

of depost...................................................................................................................................... T7

T4. Bar chart comparing approximate titanium dioxide content in titanium

deposits of different deposit types ......................................................................................... T8

T5. Plot of titanium dioxide grade and tonnage for selected igneous, metamorphic, 

and sedimentary deposits in the world, by deposit type................................................... T10

T6. Pie charts showing percentage of estimated 2012 world mine production of 

ilmenite and rutile, and percentage of global reserves of ilmenite and rutile, 

by country .................................................................................................................................. T11

U1. World map showing locations of major vanadium deposits of the world, 

by deposit type ...........................................................................................................................U3

U2. Pie chart showing percentage of world vanadium production in 2012, 

by country ....................................................................................................................................U9

U3. Graph showing major end uses of vanadium in the United States from 

1970 to 2011.................................................................................................................................U9

U4. Photographs showing four examples of vanadium............................................................U12

U5. Plot of grade and tonnage of vanadium deposits for which data 

were available..........................................................................................................................U19
xv

Tables

A1. Crustal abundances of mineral commodities (elements) included in this volume.........A7

B1. Summary of selected Federal laws relevant to mine permitting.......................................B4

B2. Historical summary of mine sites placed on the U.S. Environmental Protection

Agency’s National Priorities List. ............................................................................................B8

B3. Selected consortia devoted to identifying and implementing environmental 

best practices associated with mining ................................................................................B11

C1. Selected antimony minerals ....................................................................................................C7

C2. Estimated world production and reserves of antimony in 2013.........................................C9

C3. Potential additional sources of antimony ore and concentrate, by country .................C10

E1. Selected properties of beryllium, which is a Group 2 alkaline earth metal..................... E5

E2. Selected beryllium minerals .................................................................................................... E8

E3. Global and domestic types of magmatic-related beryllium resources .......................... E10

E4. Location, grade, tonnage, and other data for selected beryllium deposits of 

the world ................................................................................................................................... E22

F1. Location, grade, tonnage, and other data for selected cobalt deposits 

of the world................................................................................................................................F34

F2. Cobalt concentrations in rocks, soils, waters, and air.........................................................F6

G1. Fluorine concentrations in various types of rocks...............................................................G5

G2. Concentrations of fluorine and other elements in waters..................................................G6

G3. Selected examples of the chemistry and mineralogy of mine waste 

generated at fluorspar mines ................................................................................................G34

H1. Selected properties of gallium, a Group 13 post-transition metal.....................................H8

H2. Gallium concentrations in rocks, ore, coal, soils, and natural waters .............................H9

H3. Selected gallium-bearing minerals—Formula, content, and occurrence.....................H13

H4. Significant global and domestic deposit types from which gallium is obtained 

or is potentially extractable ...................................................................................................H16

I1. Classification of deposits that host germanium and indium resources ............................ I6

I2. Average estimated annual refinery production of germanium and indium, 

by area, for 2011 and 2012 ....................................................................................................... I14

I3. Germanium concentrations in rocks, soils, and waters..................................................... I16

I4. Indium concentrations in rocks, soils, waters, and air...................................................... I17

J1. Characteristics of graphite commodities, deposits, and uses, by commodity type....... J2

J2. Selected physical properties of graphite .............................................................................. J3

J3. Location, grade, tonnage and other data for selected graphite deposits and

districts of the world ................................................................................................................. J8

V1. Examples of sources and uses of zirconium and hafnium .................................................V3

V2. Graphs showing global zirconium production and U.S. trade information for 

zirconium ores and concentrates ...........................................................................................V4

V3. World map showing locations of selected zirconium and hafnium deposits 

and regions with modern coastal placer systems ...............................................................V8

V4. Diagrams showing examples of coastal depositional systems with an emphasis 

on barrier island location, morphology, and depositional environments.......................V10

V5. Diagrams showing influences on and locations of placer formation in coastal

environments............................................................................................................................V11
xvi

J4. Estimates of world graphite resources, by country, commodity type, and 

resource category, in thousand metric tons of recoverable graphite............................ J18

K1. Commercially and (or) scientifically important lithium-bearing minerals........................K5

K2. Lithium concentrations in soils developed on various types of bedrock........................K15

L1. Estimated world manganese ore reserves in 2012...............................................................L3

L2. Identified resources estimated for major land-based manganese deposits of 

the world ....................................................................................................................................L13

L3. Grade, tonnage, and quantity of contained manganese for eight manganese

deposits and districts in the United States...........................................................................L16

L4. Background and above-background concentrations of manganese in rocks, 

soils, waters, and air................................................................................................................L19

M1. Selected properties of niobium and tantalum .....................................................................M2

M2. Selected niobium and tantalum oxide minerals and their end-member Nb2O5

and Ta2O5 contents or compositional range .........................................................................M5

M3. Major types of niobium and tantalum deposits, with key characteristics and

examples ....................................................................................................................................M7

M4. Location, grade, tonnage, and other data for selected niobium-tantalum 

deposits of the world..............................................................................................................M10

N1. Chemical formulas for selected platinum-group minerals as well as other 

common rock-forming minerals mentioned in this chapter ...............................................N6

N2. Examples of rocks and ores with anomalous platinum-group-element

concentrations that are not associated with magmatic deposits, by 

deposit type. .............................................................................................................................N72

N3. Areas with significant placer platinum production, and estimates of 

cumulative production and grades.......................................................................................N23

N4. Identified platinum-group-element and gold resources, summarized by 

deposit type and location .......................................................................................................N29

N5. Igneous intrusions and intrusive complexes that contain more than 97 percent 

of the world’s identified platinum-group-element (PGE) and gold resources, 

in order of total contained PGEs ...........................................................................................N76

N6. Areal extent and stratigraphic thickness of layered intrusions with reef-type

platinum-group-element deposits and some examples of large intrusions with 

no known deposits, in order of areal extent........................................................................N77

N7. Platinum-group-element concentrations in samples of upper crust, loess, 

river sediment, and marine pelagic sediment.....................................................................N38

N8. Trace element geochemistry of waters from selected reef-type, contact-type, 

and conduit-type deposits......................................................................................................N79

N9. Acid-base accounting for selected reef-type, contact-type, and conduit-type

deposits .....................................................................................................................................N81

N10. Grade and tonnage of mineralized rock remaining in

platinum-group-element-bearing mineral deposits...........................................................N82

O1. List of the rare-earth elements found in natural deposits—the “lanthanides” 

plus yttrium .................................................................................................................................O2

O2. List of selected rare-earth-element-bearing and yttrium-bearing ore minerals ..........O10

O3. Active rare-earth mines, by deposit type ............................................................................O12

O4. Advanced rare-earth-element (REE) exploration projects and the reported 

estimates of their REE resources, by deposit type. ...........................................................O14
xvii

P1. Summary of rhenium, copper, and molybdenum grades, deposit tonnage, 

and amount of contained rhenium in the rhenium-bearing deposits shown 

in figure P3 ..................................................................................................................................P6

P2. Rhenium data for selected porphyry copper and porphyry molybdenum 

deposits of the world...............................................................................................................P38

P3. Rhenium concentrations in rocks, soils, biota, waters, and sediments.........................P20

Q1. Selenium minerals recognized by the International Mineralogical Association..........Q44

Q2. A summary of selenium concentrations in various selenides and sulfides from 

deposits around the world .....................................................................................................Q49

Q3. Selenium concentrations in sulfide minerals and other phases, in various 

deposit types ............................................................................................................................Q10

Q4. Selenium concentrations in copper-nickel ores in the Sudbury basin, 

Ontario, Canada .......................................................................................................................Q12

Q5. Selenium concentrations in selected Earth and lunar materials ....................................Q16

Q6. Estimated world selenium reserves in 2014, in metric tons..............................................Q21

Q7. National Institutes of Health recommended dietary reference intakes 

(DRIs) for selenium..................................................................................................................Q31

R1. Tellurium-bearing minerals, many of which are mentioned in this chapter....................R5

S1. List of selected tin-bearing minerals......................................................................................S3

S2. Location, grade, tonnage, and other data for selected tin deposits of the 

world with greater than 1,000 metric tons of contained tin..............................................S34

S3. Tin reserves of the world in 2016, in metric tons of contained tin...................................S11

S4. Tin concentrations in rocks, soils, waters, and air.............................................................S17

T1. Classification of selected titanium mineral deposits based on their geologic 

setting and the processes through which they were formed............................................ T3

T2. Titanium resources of the United States ............................................................................. T13

T3. Titanium mineral resources of the world (excluding the United States) for 

ilmenite (including titanomagnetite and leucoxene) and rutile (including 

anatase and brookite). ........................................................................................................... T14

U1. Location, grade, tonnage, and other data for selected vanadium deposits 

of the world.................................................................................................................................U4

U2. Selected vanadium-bearing minerals, by deposit type.....................................................U11

U3. Vanadium concentrations in rocks, soils, waters, and air................................................U23
xviii

Conversion Factors

International System of Units to Inch/Pound

Multiply By To obtain

Length

angstrom (Å) (0.1 nanometer) 0.003937 microinch

angstrom (Å) (0.1 nanometer) 0.000003937 mil

micrometer (μm) [or micron] 0.03937 mil

millimeter (mm) 0.03937 inch (in.)

centimeter (cm) 0.3937 inch (in.)

meter (m) 3.281 foot (ft) 

meter (m) 1.094 yard (yd) 

kilometer (km) 0.6214 mile (mi)

Area

hectare (ha) 2.471 acre

square kilometer (km2

) 247.1 acre

square meter (m2

) 10.76 square foot (ft2) 

square centimeter (cm2

) 0.1550 square inch (ft2) 

square kilometer (km2

) 0.3861 square mile (mi2)

Volume

milliliter (mL) 0.03381 ounce, fluid (fl. oz)

liter (L) 33.81402 ounce, fluid (fl. oz)

liter (L) 1.057 quart (qt)

liter (L) 0.2642 gallon (gal)

cubic meter (m3

) 264.2 gallon (gal) 

cubic centimeter (cm3

) 0.06102 cubic inch (in3) 

cubic meter (m3

) 1.308 cubic yard (yd3) 

cubic kilometer (km3

) 0.2399 cubic mile (mi3) 

Mass

microgram (μg) 0.00000003527 ounce, avoirdupois (oz)

milligram (mg) 0.00003527 ounce, avoirdupois (oz)

gram (g) 0.03527 ounce, avoirdupois (oz)

gram (g) 0.03215075 ounce, troy

kilogram (kg) 32.15075 ounce, troy

kilogram (kg) 2.205 pound avoirdupois (lb)

ton, metric (t) 1.102 ton, short [2,000 lb]

ton, metric (t) 0.9842 ton, long [2,240 lb]

Deposit grade

gram per metric ton (g/t) 0.0291667 ounce per short ton (2,000 lb) (oz/T)

Pressure

megapascal (MPa) 10 bar

gigapascal (GPa) 10,000 bar

Density

gram per cubic centimeter (g/cm3

) 62.4220 pound per cubic foot (lb/ft3) 

milligram per cubic meter (mg/m3

) 0.00000006243 pound per cubic foot (lb/ft3)

Energy

joule (J) 0.0000002 kilowatthour (kWh)

joule (J) 6.241×1018 electronvolt (eV)

joule (J) 0.2388 calorie (cal)

kilojoule (kJ) 0.0002388 kilocalorie (kcal)
xix

International System of Units to Inch/Pound—Continued

Multiply By To obtain

Radioactivity

becquerel (Bq) 0.00002703 microcurie (μCi)

kilobecquerel (kBq) 0.02703 microcurie (μCi)

Electrical resistivity

ohm meter (Ω-m) 39.37 ohm inch (Ω-in.)

ohm-centimeter (Ω-cm) 0.3937 ohm inch (Ω-in.)

Thermal conductivity

watt per centimeter per degree 

Celsius (watt/cm °C)

693.1798 International British thermal unit 

inch per hour per square foot per 

degree Fahrenheit (Btu in/h ft2

 °F)

watt per meter kelvin (W/m-K) 6.9318 International British thermal unit 

inch per hour per square foot per 

degree Fahrenheit (Btu in/h ft2

 °F)

Inch/Pound to International System of Units

Length

mil 25.4 micrometer (μm) [or micron]

inch (in.) 2.54 centimeter (cm)

inch (in.) 25.4 millimeter (mm)

foot (ft) 0.3048 meter (m)

mile (mi) 1.609 kilometer (km)

Volume

ounce, fluid (fl. oz) 29.57 milliliter (mL)

ounce, fluid (fl. oz) 0.02957 liter (L) 

Mass

ounce, avoirdupois (oz) 28,350,000 microgram

ounce, avoirdupois (oz) 28,350 milligram

ounce, avoirdupois (oz) 28.35 gram (g) 

ounce, troy 31.10 348 gram (g)

ounce, troy 0.03110348 kilogram (kg)

pound, avoirdupois (lb) 0.4536 kilogram (kg) 

ton, short (2,000 lb) 0.9072 ton, metric (t) 

ton, long (2,240 lb) 1.016 ton, metric (t) 

Deposit grade

ounce per short ton (2,000 lb) (oz/T) 34.285714 gram per metric ton (g/t)

Energy

kilowatthour (kWh) 3,600,000 joule (J)

electronvolt (eV) 1.602×10–19 joule (J)

Radioactivity

microcurie (μCi) 37,000 becquerel (Bq)

microcurie (μCi) 37 kilobecquerel (kBq)

Temperature in degrees Celsius (°C) may be converted to degrees Fahrenheit (°F) as follows:

°F =(1.8×°C) +32

Temperature in degrees Celsius (°C) may be converted to kelvin (K) as follows:

K=°C+273.15

Temperature in degrees Fahrenheit (°F) may be converted to degrees Celsius (°C) as follows:

°C= (°F–32)/ 1.8
xx

Datum

Unless otherwise stated, vertical and horizontal coordinate information is referenced to the 

World Geodetic System of 1984 (WGS 84). Altitude, as used in this report, refers to distance 

above the vertical datum.

Supplemental Information

Specific conductance is given in microsiemens per centimeter at 25 degrees Celsius (μS/cm 

at 25 °C).

Concentrations of chemical constituents in soils and (or) sediment are given in milligrams per 

kilogram (mg/kg), parts per million (ppm), or parts per billion (ppb).

Concentrations of chemical constituents in water are given in milligrams per liter (mg/L), 

micrograms per liter (μg/L), nanogams per liter (ng/L), nanomoles per kilogram (nmol/kg), 

parts per million (ppm), parts per billion (ppb), or parts per trillion (ppt).

Concentrations of suspended particulates in water are given in micrograms per gram (μg/g), 

milligrams per kilogram (mg/kg), or femtograms per gram (fg/g).

Concentrations of chemicals in air are given in units of the mass of the chemical (milligrams, 

micrograms, nanograms, or picograms) per volume of air (cubic meter).

Activities for radioactive constituents in air are given in microcuries per milliliter (μCi/mL).

Deposit grades are commonly given in percent, grams per metric ton (g/t)—which is equivalent 

to parts per million (ppm)—or troy ounces per short ton (oz/T).

Geologic ages are expressed in mega-annum (Ma, million years before present, or 106 years ago)

or giga-annum (Ga, billion years before present, or 109 years ago).

For ranges of years, “to” and (or) the en dash (“–”) mean “up to and including.”

Concentration unit Equals

milligram per kilogram (mg/kg) part per million

microgram per gram (μg/g) part per million

microgram per kilogram (μg/kg) part per billion (109

)

Equivalencies

part per million (ppm): 1 ppm=1,000 ppb=1,000,000 ppt=0.0001 percent

part per billion (ppb): 0.001 ppm=1 ppb=1,000 ppt=0.0000001 percent

part per trillion (ppt): 0.000001 ppm=0.001 ppb=1 ppt=0.0000000001 percent

Metric system prefixes

tera- (T-) 1012 1 trillion

giga- (G-) 109 1 billion

mega- (M-) 106 1 million

kilo- (k-) 103 1 thousand

hecto- (h-) 102 1 hundred

deka- (da-) 10 1 ten

deci- (d-) 10–1 1 tenth

centi- (c-) 10–2 1 hundredth

milli- (m-) 10–3 1 thousandth

micro- (μ-) 10–6 1 millionth

nano- (n-) 10–9 1 billionth

pico- (p-) 10–12 1 trillionth

femto- (f-) 10–15 1 quadrillionth

atto- (a-) 10–18 1 quintillionth
xxi

Abbreviations and Symbols

[This list is intended to help the reader identify the meanings of various abbreviations and symbols used in the text. To 

avoid unnecessary duplication, not all abbreviations and symbols are included. For example, symbols that are used in 

only a single table and that are identified in that table’s headnote are not included here. Likewise, mineral names and 

their chemical formulas are not listed here]

α-tin alpha-tin (also known as gray tin)

ß-tin beta-tin (also known as white tin)

°C degree Celsius

δ-MnO2 delta-manganese dioxide

μCi/mL microcurie per milliliter

μg Al/L microgram of aluminum per liter

μg/d microgram per day

μg/g microgram per gram

μg/kg microgram per kilogram

μg/L microgram per liter

μg/m3 microgram per cubic meter

μg Mn/L microgram of manganese per liter

μg Mn/m3 microgram of manganese per cubic meter

μg V/L microgram of vanadium per liter

μm micrometer

3TG tantalum, tin, tungsten, and (or) gold

Å angstrom

AC alternating current

ADTI Acid Drainage Technology Initiative

AFRG alkali-feldspar rhyolite-granite

AIDS acquired immunodeficiency syndrome

AMD acid mine drainage

AMT audio magnetotelluric

API American Petroleum Institute

APS American Physical Society

ATPC Association of Tin Producing Countries

ATSDR Agency for Toxic Substances and Disease Registry

BCMA beryllium-copper master alloy

Bq/kg becquerel per kilogram

ca. circa (about)

CBMM Companhia Brasileira de Metalurgia e Mineração

CCD carbonate compensation depth

CCZ Clarion-Clipperton zone

CERCLA Comprehensive Environmental Response, Compensation, and 

Liability Act
xxii

CIGS copper-indium-gallium-selenide

cm centimeter

cm3 cubic centimeter

CMI Critical Minerals Index

Congo (Kinshasa) Democratic Republic of the Congo

CPV concentrator photovoltaic

CRD carbonate replacement deposit

DC direct current

DOD U.S. Department of Defense

DOE U.S. Department of Energy

EA environmental assessment

EC50 effective concentration 50 (concentration that results in 

50 percent exhibiting decreased functionality)

EEZ Exclusive Economic Zone

Eh oxidation potential

EIS environmental impact statement

EMPA electron microprobe analysis

EPA U.S. Environmental Protection Agency

FCC fluid catalytic cracking

FONSI finding of no significant impact

ft foot

g gram

g CO2-equiv/kWh gram of CO2 equivalent per kilowatthour

g/cm3 gram per cubic centimeter

g/d gram per day

g/L gram per liter

g/t gram per metric ton

Ga giga-annum

GARD Global Acid Rock Drainage

GWh gigawatt-hour

HDC Hicks Dome Corp.

HFSE high-field-strength element

HIV human immunodeficiency virus

HMIS Hazardous Materials Identification System

HREE heavy rare-earth element

HSLA high-strength, low-alloy

IARC International Agency for Research on Cancer

ICP–MS inductively coupled plasma-mass spectrometry

IED improvised explosive device
xxiii

IMA

INAP

IOCG

IREL

ISA

ISMI

ITO

JORC

kBq/kg

kg

kg CaCO3  / t

kg/cm2

kg/m3

kg/t

km

km2

km3

KMML

lb

LC50

LCT

LED

LIP

LOEL

log f O2

LREE

m

M

m.y.

m3

Ma

MC–ICP–MS

MDEQ

MEND

mg

mg Al/kg

mg/d

mg/kg

International Mineralogical Association

International Network for Acid Prevention

iron oxide-copper-gold

Indian Rare Earths Ltd.

International Seabed Authority

International Strategic Minerals Inventory

indium-tin oxide

Joint Ore Reserves Committee (Australia)

kilobecquerel per kilogram

kilogram

kilogram of calcium carbonate per metric ton

kilogram per square centimeter

kilogram per cubic meter

kilogram per metric ton

kilometer

square kilometer

cubic kilometer

Kerala Minerals and Metals Ltd.

pound

lethal concentration 50 (concentration that kills 50 percent 

of test population within a given timeframe)

lithium-cesium-tantalum

light-emitting diode

large igneous province

lowest observable effect limit

log of oxygen fugacity

light rare-earth element

meter

molarity

million years

cubic meter

mega-annum

multiple collector-inductively coupled plasma-mass spectrometry

Michigan Department of Environmental Quality

Mine Environment Neutral Drainage

milligram

milligram of aluminum per kilogram

milligram per day

milligram per kilogram
xxiv

mg/L milligram per liter

mg/m3 milligram per cubic meter

Mkg/yr million kilograms per year

mm millimeter

MOFCOM Ministry of Commerce (China)

MRDS Mineral Resources Data System (USGS)

MRI magnetic resonance imaging

MRS Materials Research Society

MSS monosulfide solution

MTR Maderia-Tore Rise

MVT Mississippi Valley-type

NEPA National Environmental Policy Act

NFPA National Fire Protection Association

ng/L nanogram per liter

ng/m3 nanogram per cubic meter

NI National Instrument

NiMH nickel-metal-hydride

NIOSH National Institute for Occupational Safety and Health

nm nanometer

NMRI nuclear magnetic resonance instruments

NPL National Priorities List (Superfund)

NREL National Renewable Energy Laboratory

NSTC National Science and Technology Council

NTP National Toxicology Program

NYF niobium-yttrium-fluorine

OPEC Organization of the Petroleum Exporting Countries

org-Se(-II) organo-selenide

OSHA Occupational Safety and Health Administration

PEM proton exchange membrane

PET polyethylene terephthalate

pg/L picogram per liter

pg/m3 picogram per cubic meter

PGE platinum-group element

PGM platinum-group metal

pm picometer

ppb part per billion

ppm part per million

ppt part per trillion
xxv

QSP quartz-sericite-pyrite

RDA recommended daily allowance

REE rare-earth element

REO rare-earth oxide

RF radio-frequency

SAF submerged-arc furnace

SAMREC South African Code for Reporting Exploration Results

Se(II) selenide

Se(IV) selenite

Se(VI) selenate

SEC U.S. Securities and Exchange Commission

SEDEX sedimentary exhalative

SG specific gravity

SHRIMP sensitive high-resolution ion microprobe

SSV sandstone-hosted vanadium

SX–EW solvent extraction-electrowinning

TBT tributyltin

TBTCl tributyltin chloride

TBTO bis(tri-n-butyltin) oxide

Te(II) telluride

Te(IV) tellurite

Te(VI) tellurate

TPT triphenyltin

TRE Tantalus Rare Earths

TREO total rare-earth oxide

TWA time-weighted average

U.S.C. U.S. Code

USGS U.S. Geological Survey

UST unidirectional solidification texture

VLF very low frequency

VMS volcanogenic massive sulfide

VRB vanadium redox-flow battery

VTM vanadiferous titanomagnetite

WGS 84 World Geodetic System of 1984

WHO World Health Organization

wt. % weight percent

Critical Mineral Resources of the

United States—An Introduction

By Klaus J. Schulz, John H. DeYoung, Jr., Dwight C. Bradley, and Robert R. Seal II

Chapter A of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–A

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter A

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................A1

Background..................................................................................................................................................A1

U.S. Mineral Supply Situation ...................................................................................................................A2

Mineral Resource Classifications, Terms, and Definitions Used in This Volume .............................A4

Joint Products, Byproducts, and Coproducts ...............................................................................A5

Scarcity and Rarity ............................................................................................................................A6

Mineral Commodities Selected for Inclusion in This Volume ..............................................................A6

Changes Since the Mid-1970s ..................................................................................................................A7

Back to the Future.......................................................................................................................................A8

References Cited.........................................................................................................................................A9

Appendix A1. What is Meant by “Critical” Minerals (or Materials)? .............................................A11

Suggested citation:

Schulz, K.J., DeYoung, J.H., Jr., Bradley, D.C., and Seal, R.R., II, 2017, Critical mineral resources of the United 

States—An introduction, chap. A of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical 

mineral resources of the United States—Economic and environmental geology and prospects for future supply: 

U.S. Geological Survey Professional Paper 1802, p. A1– A14, http://dx.doi.org/10.3133/pp1802A.
Critical Mineral Resources of the United States—

An Introduction

By Klaus J. Schulz, John H. DeYoung, Jr., Dwight C. Bradley, and Robert R. Seal II

Abstract

Many changes have taken place in the mineral resource 

sector since the publication by the U.S. Geological Survey of 

Professional Paper 820, “United States Mineral Resources” 

(Brobst and Pratt, 1973), which is a review of the long-term 

United States resource position for 65 mineral commodities or 

commodity groups. For example, since 1973, the United States 

has continued to become increasingly dependent on imports 

to meet its demands for an increasing number of mineral 

commodities. The global demand for mineral commodities is at 

an alltime high and is expected to continue to increase, and the 

development of new technologies and products has led to the 

use of a greater number of mineral commodities in increasing 

quantities to the point that, today, essentially all naturally 

occurring elements have several significant industrial uses. 

Although most mineral commodities are present in sufficient 

amounts in the earth to provide adequate supplies for many 

years to come, their availability can be affected by such factors 

as social constraints, politics, laws, environmental regulations, 

land-use restrictions, economics, and infrastructure.

This volume presents updated reviews of 23 mineral 

commodities and commodity groups viewed as critical to a 

broad range of existing and emerging technologies, renewable 

energy, and national security. The commodities or commodity 

groups included are antimony, barite, beryllium, cobalt, fluorine, 

gallium, germanium, graphite, hafnium, indium, lithium, 

manganese, niobium, platinum-group elements, rare-earth 

elements, rhenium, selenium, tantalum, tellurium, tin, titanium, 

vanadium, and zirconium. All these commodities have been 

listed as critical and (or) strategic in one or more of the recent 

studies based on assessed likelihood of supply interruption and 

the possible cost of such a disruption to the assessor. For some 

of the minerals, current production is limited to only one or 

a few countries. For many, the United States currently has no 

mine production or any significant identified resources and is 

largely dependent on imports to meet its needs. As a result, the 

emphasis in this volume is on the global distribution and availability of each mineral commodity. The environmental issues 

related to production of each mineral commodity, including 

current mitigation and remediation approaches to deal with 

these challenges, are also addressed.

This introductory chapter provides an overview of the 

mineral resource classifications, terms, and definitions used in 

this volume. A review of the history of the use and meaning 

of the term “critical” minerals (or materials) is included as an 

appendix to the chapter.

Background

In 1973, the U.S. Geological Survey (USGS) produced 

Professional Paper 820, “United States Mineral Resources” 

(Brobst and Pratt, 1973), a review of the long-term 

United States resource position for 65 mineral commodities 

or commodity groups. A main purpose of that volume was 

to provide nonspecialists with easily understandable factual 

information on the resources of the many mineral commodities 

that are important to the national economy, national security, 

and the everyday lives of U.S. citizens. More specifically, 

that volume addressed three basic questions about each 

mineral commodity:

“(1) How important is it to our present industrial 

civilization and standard of living? (2) how much of it 

do we have and to what extent is it economically and 

technologically available? and (3) how and where can 

we find more?” (Brobst and Pratt, 1973, p. 1).

The emphasis was on domestic mineral resources, but many

chapters also include information on resources in other countries.

The economy and the national security of the 

United States are based directly or indirectly on minerals 

and, in the early 1970s, there was increasing recognition of 

and concern about the fact that the United States did not have 

adequate domestic supplies of many of the nonfuel minerals 

needed to sustain its economy (U.S. Geological Survey, 1975). 

In an analysis of the principal developments in U.S. mineral 

history and world mineral history since 1939, Cameron (1973) 

concluded that:

“(1) United States mineral production has greatly 

increased but has not kept pace with consumption. 

Self-sufficiency in minerals has declined, both 

overall and in numbers of minerals involved.
A2 Critical Mineral Resources of the United States—An Introduction

(2) Since 1945, world mineral production has

increased far more rapidly than United States

production. The relative importance of the

United States as a supplier of world minerals,

raw and manufactured, has rapidly declined.

(3) World mineral consumption has increased far

more rapidly than United States consumption. The

United States is no longer the world’s principal

market for mineral raw materials, and it faces

increasing competition for the world’s mineral

supplies.” (Cameron, 1973, p. 25).

The volume “United States Mineral Resources” was produced

to provide factual information on the Nation’s mineral 

resources at the time and to inform policymakers and the 

public in efforts to frame national mineral policy.

U.S. Mineral Supply Situation

The conclusions reached by Cameron (1973) about the 

United States mineral position at that time are even more pertinent now, more than 40 years later. In 1980, the United States 

was more than 50 percent dependent on imports to meet its 

annual requirements for more than 20 major nonfuel mineral 

commodities; in 2014, that number had increased to more than 

40 commodities (Fortier and others, 2015; U.S. Geological 

Survey, 2015b, p. 6). The United States was the world’s leading 

producer of copper for much of the 20th century, but Chile 

became the leading copper-producing country in 1982 and 

has remained so. In 2014, China accounted for 20 percent or 

more of the world’s mine production of more than 40 mineral 

commodities; these included the rare-earth elements, of which 

China accounted for 85 percent of world production; tungsten, 

82 percent; antimony, 76 percent; germanium, 73 percent; 

mercury, 68 percent; graphite, 66 percent; fluorspar, 59 percent; 

and bismuth, 56 percent (U.S. Geological Survey, 2016).

As world population increases and the average standard 

of living improves, the global demand for mineral commodities is at an all-time high and is expected to continue to 

increase. Much of this demand is the result of industrialization 

in large developing countries, such as Brazil, China, and India.

In addition, the development of new technologies and 

products has led to the use of a greater number of mineral 

commodities in increasing quantities. In 1932, uranium and 

rare-earth elements had only minor uses, and the production 

or uses of such elements as gallium, germanium, rhenium, and 

several others were not even tracked. As of the first decades 

of the 21st century, essentially all the naturally occurring 

elements have several significant industrial uses (Price, 2013). 

By way of example, in the 1980s, 12 elements were used in the 

manufacture of computer chips. A decade later, 16 elements 

were employed, and by 2006, as many as 60 elements were 

used in the manufacture of high-speed, high-capacity integrated 

circuits (fig. A1; National Research Council, 2008, p. 56–58).

In the 1970s, concerns regarding the adequacy of future 

supplies of minerals were raised owing to a perceived lack 

of world resources. Based on their finite occurrence, “limits 

to supply” of nonrenewable resources, such as minerals, was 

most famously predicted in a report by the Club of Rome 

(Meadows and others, 1972). The concerns raised in that 

report were based partly on results of previous studies. For 

example, the time-series analysis of metal production in 

mining districts of Europe by Hewett (1929) demonstrated 

cyclical patterns that were subsequently applied to the 

production histories of other nonfuel minerals and energy 

minerals (Lasky, 1951, 1955; Hubbert, 1956; Pazik, 1976). 

This type of analysis formed the basis of numerous studies 

that predicted that the highest level of production of some 

mineral commodities may soon be reached and that production 

is likely to decline from then on. This approach to studying 

future resource availability is often termed “peak minerals.” 

Many of these studies, however, used estimates of mineral 

reserves (the economically extractable portion of resources at 

the time of determination) as a proxy for the Earth’s resources, 

including those that have been discovered (identified 

resources), which, in turn, include those that are economic to 

produce (reserves) (Meadows and others, 1972; Cohen, 2007). 

This approach is flawed, because estimates of mineral reserves 

are a function of economic factors, such as metal prices and 

costs of recovery, which can vary considerably over time. 

New geologic knowledge can result in changes to estimates of 

identified resources, including those identified resources that 

are economic (reserves) changing as well (Rustad, 2012; Gold, 

2014; Meinert and others, 2016). Likewise, both the development of new extractive technologies and increased metal 

prices have resulted in mineral deposits previously considered 

subeconomic or marginally economic becoming viable sources 

for mineral production. In addition, improved methodologies 

for estimating undiscovered mineral resources, the discovery 

of new mineral deposit types, improved recycling technologies, higher processing efficiency, and longer product life 

have all helped allay fears that we are near peak supply of 

most mineral commodities. Today, therefore, it is generally 

recognized that, although mineral commodities are mostly 

nonrenewable on human time scales and are inherently finite, 

fears of resource depletion for most mineral commodities 

anytime soon are unwarranted.

Nonetheless, although most mineral commodities 

may be present in sufficient amounts in the earth to provide 

adequate supplies for some years to come, they are generally 

found concentrated only in small volumes in the crust and 

are not distributed evenly across the planet. As a result, no 

country today can be fully self-sufficient in meeting all its 

mineral resource needs. In addition, the availability of mineral 

commodities is not just a function of geologic accessibility but 

of such factors as social constraints, politics, laws, environmental regulations, land restrictions, economics, and infrastructure. Large deposits of a particular mineral commodity 

tend to account for the bulk of production, but these deposits 
U.S. Mineral Supply Situation A3

1A

2A

3B 4B 5B 6B 7B 8B 11B 12B

3A 4A 5A 6A 7A

1A

2A

3B 4B 5B 6B 7B 8B 11B 12B

3A 4A 5A 6A 7A

1A

2A

3B 4B 5B 6B 7B 8B 11B 12B

3A 4A 5A 6A 7A

11

1

3 4

12

19 20 21 22 23 24 25 26 27 28 29 30

37 38 39 40 41 42 43 44 45 46 47 48

55 56

58 59 60

72 73 74 75 76 77 78 79 80

87 88

90

57

89 91 92 93 94 95 96

104 105 106 107 108 109 110 111 112

61 62 63 64 65 66 67

97 98 99

68 69 70 71

100 101 102 103

31

13 14 15 16 17 18

32 33 34 35 36

49 50 51 52 53 54

81 82 83 84 85 86

5 6 7 8 9 10

2

113 114 115 116 117 118

11

1

3 4

12

19 20 21 22 23 24 25 26 27 28 29 30

37 38 39 40 41 42 43 44 45 46 47 48

55 56

58 59 60

72 73 74 75 76 77 78 79 80

87 88

90

57

89 91 92 93 94 95 96

104 105 106 107 108 109 110 111 112

61 62 63 64 65 66 67

97 98 99

68 69 70 71

100 101 102 103

31

13 14 15 16 17 18

32 33 34 35 36

49 50 51 52 53 54

81 82 83 84 85 86

5 6 7 8 9 10

2

113 114 115 116 117 118

11

1

3 4

12

19 20 21 22 23 24 25 26 27 28 29 30

37 38 39 40 41 42 43 44 45 46 47 48

55 56

58 59 60

72 73 74 75 76 77 78 79 80

87 88

90

57

89 91 92 93 94 95 96

104 105 106 107 108 109 110 111 112

61 62 63 64 65 66 67

97 98 99

68 69 70 71

100 101 102 103

31

13 14 15 16 17 18

32 33 34 35 36

49 50 51 52 53 54

81 82 83 84 85 86

5 6 7 8 9 10

2

113 114 115 116 117 118

39.10

9.012

24.31

40.08 44.96 47.88 50.94 52.00 54.94 55.85 58.93 58.69 63.55 65.39

10.81

26.98

12.01

28.09

14.01

69.72 72.64

30.97

74.92

16.00 19.00 20.18

4.003

32.06 35.45 39.95

78.96 79.90 83.79

85.47

132.9

87.62

137.3

88.91 91.22

178.5

92.91

180.9

95.96 

183.9

(98)

186.2

101.1

190.2

102.9 106.4

195.1

107.9 112.4 114.8 118.7 121.8

(209) (210) (222)

127.6 126.9 131.3

192.2 197.0 200.5 204.4 207.2 209.0

(223) (226) (265) (268) (271) (270) (277) (276) (281) (280) (285)

140.1 140.9 144.2 (145) 150.4 152.0 157.2 158.9 162.5 164.9 167.3 168.9 173.0 175.0

232

138.9

(227) 231 238 (237) (244) (243) (247) (247) (251) (252) (257) (258) (259) (262)

1.008

6.94

22.99

(284) (289) (288) (293) (294) (294)

39.10

9.012

24.31

40.08 44.96 47.88 50.94 52.00 54.94 55.85 58.93 58.69 63.55 65.39

10.81

26.98

12.01

28.09

14.01

69.72 72.64

30.97

74.92

16.00 19.00 20.18

4.003

32.06 35.45 39.95

78.96 79.90 83.79

85.47

132.9

87.62

137.3

88.91 91.22

178.5

92.91

180.9

95.96 

183.9

(98)

186.2

101.1

190.2

102.9 106.4

195.1

107.9 112.4 114.8 118.7 121.8

(209) (210) (222)

127.6 126.9 131.3

192.2 197.0 200.5 204.4 207.2 209.0

(223) (226) (265) (268) (271) (270) (277) (276) (281) (280) (285)

140.1 140.9 144.2 (145) 150.4 152.0 157.2 158.9 162.5 164.9 167.3 168.9 173.0 175.0

232

138.9

(227) 231 238 (237) (244) (243) (247) (247) (251) (252) (257) (258) (259) (262)

1.008

6.94

22.99

(284) (289) (288) (293) (294) (294)

39.10

9.012

24.31

40.08 44.96 47.88 50.94 52.00 54.94 55.85 58.93 58.69 63.55 65.39

10.81

26.98

12.01

28.09

14.01

69.72 72.64

30.97

74.92

16.00 19.00 20.18

4.003

32.06 35.45 39.95

78.96 79.90 83.79

85.47

132.9

87.62

137.3

88.91 91.22

178.5

92.91

180.9

95.96 

183.9

(98)

186.2

101.1

190.2

102.9 106.4

195.1

107.9 112.4 114.8 118.7 121.8

(209) (210) (222)

127.6 126.9 131.3

192.2 197.0 200.5 204.4 207.2 209.0

(223) (226) (265) (268) (271) (270) (277) (276) (281) (280) (285)

140.1 140.9 144.2 (145) 150.4 152.0 157.2 158.9 162.5 164.9 167.3 168.9 173.0 175.0

232

138.9

(227) 231 238 (237) (244) (243) (247) (247) (251) (252) (257) (258) (259) (262)

1.008

6.94

22.99

(284) (289) (288) (293) (294) (294)

hydrogen

barium

francium radium

strontium

sodium

vanadium

lithium beryllium

magnesium

potassium calcium

rubidium

cesium

helium

boron carbon nitrogen oxygen fluorine neon

aluminum silicon phosphorus sulfur chlorine argon

scandium titanium chromium manganese iron cobalt nickel copper zinc gallium germanium arsenic selenium bromine krypton

yttrium zirconium niobium molybdenum technetium ruthenium rhodium palladium silver cadmium indium tin antimony tellurium iodine xenon

hafnium

cerium praseodymium neodymium promethium samarium europium gadolinium terbium dysprosium holmium erbium thulium ytterbium lutetium

tantalum tungsten rhenium osmium iridium platinum gold mercury thallium lead bismuth polonium astatine radon

thorium

lanthanum

actinium protactinium uranium neptunium plutonium americium curium berkelium californium einsteinium fermium mendelevium nobelium lawrencium 

rutherfordium dubnium seaborgium bohrium hassium meitnerium darmstadtium roentgenium copernicium flerovium livermorium

hydrogen

barium

francium radium

strontium

sodium

vanadium

lithium beryllium

magnesium

potassium calcium

rubidium

cesium

helium

boron carbon nitrogen oxygen fluorine neon

aluminum silicon phosphorus sulfur chlorine argon

scandium titanium chromium manganese iron cobalt nickel copper zinc gallium germanium arsenic selenium bromine krypton

yttrium zirconium niobium molybdenum technetium ruthenium rhodium palladium silver cadmium indium tin antimony tellurium iodine xenon

hafnium

cerium praseodymium neodymium promethium samarium europium gadolinium terbium dysprosium holmium erbium thulium ytterbium lutetium

tantalum tungsten rhenium osmium iridium platinum gold mercury thallium lead bismuth polonium astatine radon

thorium

lanthanum

actinium protactinium uranium neptunium plutonium americium curium berkelium californium einsteinium fermium mendelevium nobelium lawrencium 

rutherfordium dubnium seaborgium bohrium hassium meitnerium darmstadtium roentgenium copernicium flerovium livermorium

hydrogen

barium

francium radium

strontium

sodium

vanadium

lithium beryllium

magnesium

potassium calcium

rubidium

cesium

helium

boron carbon nitrogen oxygen fluorine neon

aluminum silicon phosphorus sulfur chlorine argon

scandium titanium chromium manganese iron cobalt nickel copper zinc gallium germanium arsenic selenium bromine krypton

yttrium zirconium niobium molybdenum technetium ruthenium rhodium palladium silver cadmium indium tin antimony tellurium iodine xenon

hafnium

cerium praseodymium neodymium promethium samarium europium gadolinium terbium dysprosium holmium erbium thulium ytterbium lutetium

tantalum tungsten rhenium osmium iridium platinum gold mercury thallium lead bismuth polonium astatine radon

thorium

lanthanum

actinium protactinium uranium neptunium plutonium americium curium berkelium californium einsteinium fermium mendelevium nobelium lawrencium 

rutherfordium dubnium seaborgium bohrium hassium meitnerium darmstadtium roentgenium copernicium flerovium livermorium

Actinide Series**

Lanthanide Series*

Actinide Series**

Lanthanide Series*

Actinide Series**

Lanthanide Series*

1980s

(12 elements)

1990s

(+4 elements)

2000s

(+45 elements)

8A

8A

8A

Figure A1. Diagram showing 

increases in the use of elements 

over two decades of computer chip 

technology development. Highspeed, high-capacity integrated 

circuits have gone from being 

made with A, 12 minerals or their 

elemental components in the 1980s 

to B, 16 in the 1990s to C, more 

than 60 by the 2000s. Modified from 

National Research Council (2008, 

fig. 2.2).
A4 Critical Mineral Resources of the United States—An Introduction

are few in number and geographically restricted. As a result, 

there is potential for local social and environmental issues or 

political manipulation to restrict or deny access to supply.

The sometimes-tenuous nature of the mineral supply 

chain received world attention in 2010 when China suddenly 

drastically cut its export quota for the rare-earth elements. The 

move highlighted the fact that China had a virtual monopoly 

on the short-term supply of rare-earth elements—elements that 

are essential to the renewable energy sector and many other 

high-tech applications globally. The rest of the world was left 

scrambling to find alternative and secure supplies. China is 

also the world’s major producer of a number of other mineral 

commodities that are essential in high-tech applications, 

renewable energy, and national security, including antimony, 

bismuth, fluorspar, germanium, graphite, and indium (Price, 

2013). These and other mineral commodities that are largely 

controlled by one country, such as cobalt (Democratic 

Republic of the Congo), niobium (Brazil), and platinum 

(South Africa), are also considered to be at high risk of supply 

disruption and would have high impact if supply restrictions 

should take place (McGroarty and Wirtz, 2012).

Mineral Resource Classifications, 

Terms, and Definitions Used in

This Volume

Through the years, a variety of terms have been used to 

describe and classify mineral resources. Although some terms 

have gained wide use and acceptance, they are not always 

used with precisely the same meaning. Some basic terms and 

definitions are listed below. No attempt has been made to 

include a detailed listing of all mineral-resource-related terms 

and definitions; rather, emphasis is on those basic terms that 

are essential to proper understanding of the chapters in this 

volume. This terminology is intended to represent standard 

definitions and usage by the mineral industry and resource 

assessment community today. A more comprehensive discussion of mineral resource terms and definitions is presented in 

U.S. Bureau of Mines and U.S. Geological Survey (1980); 

Cox and others (1986); and U.S. Geological Survey National 

Mineral Resource Assessment Team (2000).

Mineral resources are classified based on a number of 

factors, such as their geologic and physical and (or) chemical 

characteristics (such as grade, mineralogy, tonnage, thickness, 

and depth) of the material in place; the economic viability 

of production based on costs of extracting, processing, 

and marketing the material at a given time; the level of 

certainty of estimates of physical and economic factors; and 

environmental and legal constraints on resource development. The mineral resource classification system used in this 

volume is shown in figure A2 (U.S. Bureau of Mines and 

U.S. Geological Survey, 1980).

When used in this volume, the terms listed below have 

the following meanings:

mineral occurrence. A mineral concentration that is considered valuable by someone somewhere, or that is of scientific 

or technical interest.

mineral deposit. A mineral concentration of sufficient size and

grade that it might, under the most favorable of circumstances, 

be considered to have potential for economic development.

undiscovered mineral deposit. A mineral deposit

believed to exist 1 kilometer or less below the surface

of the ground, or an incompletely explored mineral

occurrence or prospect that could have sufficient size

and grade to be classified as a deposit.

ore deposit. A mineral deposit that has been tested

and is known to be of sufficient size, grade, and

accessibility to be producible to yield a profit.

resource. A mineral concentration of sufficient size and 

grade and in such a form and amount that economic extraction 

of a commodity from the concentration is currently or potentially feasible.

	 identified resources. Resources whose location, grade, 

quality, and quantity are known or estimatable from 

specific geologic evidence.

 reserves. Identified resources that meet

specified minimum physical and chemical

 criteria related to current mining and production

 practices and that can be economically extracted

 or produced at the time of determination.

undiscovered resources. Resources in undiscovered 

mineral deposits whose existence is postulated on the 

basis of indirect geologic evidence.

economic. Profitable extraction or production under defined 

investment assumptions has been established, analytically 

demonstrated, or assumed with reasonable certainty.

A mineral occurrence becomes a mineral resource only if 

there is a demand to use it—that is, inherent in the definition 

of a mineral resource is the necessity for there to be present 

or expected future demand for a mineral commodity in order 

for an occurrence of that mineral to be considered a resource. 

Note also that the quantity of the resource within each 

category is a stock variable (a measurement at a specific point 

in time), as contrasted with a flow variable, such as production 

(a measurement that covers a period of time). The quantity of 

resources (including reserves) is not static, and the assignment 

of the resource to a given category often changes over time.

The development of mineral resources into the category 

of reserves requires the creation of facilities to extract the ore, 

process it into a mineral product, and transport that product 

to the user. Although the quantity of production is a flow 

measure (amount per unit time—for example, tons per month 

or tons per year), the estimates of the amount available in each 

resource category (a stock variable) also changes over time as 
Mineral Resource Classifications, Terms, and Definitions Used inThis Volume A5

Identified resources Undiscovered resources

Cumulative

production Demonstrated Probability range

Measured Indicated Hypothetical Speculative Inferred

Economic

Marginally

economic

Subeconomic

Reserves

Marginal reserves

Demonstrated

subeconomic resources

Inferred

reserves

Inferred

marginal 

reserves

Inferred

subeconomic

resources

Increasing geologic certainty

Increasing economic feasibility

Figure A2. Diagram showing the mineral resource classification system used in this volume. Modified from 

U.S. Bureau of Mines and U.S. Geological Survey (1980).

production removes part of the resource; additional exploration results in the discovery of new deposits and provides new 

information about previously discovered deposits; advances in 

technology affect the costs and methods of mining, processing, 

and use; and changes in market conditions affect the economic 

viability of production. For example, in 1973, world reserves 

of cobalt were estimated to be 2.5 million metric tons of 

contained cobalt (U.S. Bureau of Mines, 1973, p. 39). Since 

then, more than 2.0 million metric tons of cobalt has been 

mined worldwide, yet world reserves of cobalt in 2015 were 

estimated to be 7.2 million metric tons of cobalt, which 

is almost triple the amount estimated in 1973, despite the 

depletion by mining of an amount equivalent to 80 percent of 

the 1973 reserve estimate (Shedd, 2015, 2016; U.S. Geological 

Survey, 2015a).

E.W. Zimmermann saw resources as a function of human 

wants and abilities. In the revised edition of his 1933 book, 

“World Resources and Industries—A Functional Appraisal of 

the Availability of Agricultural and Industrial Resources,” he 

discussed the debate between “the static school who insists 

that ‘resources are’,” and “the dynamic, functional, operational 

school who insists that ‘resources become’” (Zimmermann, 

1951, p. 11). He succinctly stated the relation between demand 

and resources as “Resources are not, they become.”

Joint Products, Byproducts, and Coproducts

Many of the mineral commodities included in this volume 

are not produced as independent (also called individual, primary, or 

main) products; rather, they are produced (mined and [or] processed) 

along with other minerals. The terms used to describe the mineral 

commodities produced at a particular site can give an indication 

of their relative economic importance at that site. The terms, 

however, are not always used consistently across the industry.

The term “byproduct,” for example, has been used to 

characterize any mineral commodity produced along with 

another as a “secondary product” (Kesler, 1994, p. 371) and 

(or) as a “secondary or additional product” (Thrush and others, 

1968, p. 157). Landsberg and others (1963, p. 479–483) used 

the terms “byproduct” and “coproduct” interchangeably in a 

detailed study of the future availability of natural resources 

to meet projected material use. Brooks (1965, p. 25–31), who 

was a senior contributor to that study, later described a “joint 

product” relationship when he developed his classification 

system for minor metals.

Brooks defined an individual product as one that 

is produced alone or with other minerals that are “of 

comparatively insignificant value.” Joint products are called 

“byproducts” or “coproducts” depending upon their effect 

on the economic viability of the production process. If the 

joint product is needed to make the operation viable, it is 

a coproduct; if not, it is a byproduct. Changes in market 

conditions (prices and costs of production processes) and in 

applicable mining and processing technologies can change 

the status of joint products. With this caveat, the uses of the 

terms joint product, byproduct, and coproduct in this volume 

follow the definitions established by Brooks (1965). The 

revised version of the 1968 U.S. Bureau of Mines dictionary 

of mining and mineral terms published by the American 

Geological Institute in 1997 added a definition of “coproduct” 

that is cross-referenced to the definition of “byproduct” that 

is generally consistent with the classification developed by 

Brooks (American Geological Institute, 1997, p. 76, 124).
A6 Critical Mineral Resources of the United States—An Introduction

Scarcity and Rarity

Analysis of mineral resources is linked to concerns 

about scarcity because resources are defined as physical 

quantities that can be extracted at a cost that is commensurate 

with current or potential profitable marketability. Brooks 

(1965, p. 22) explained the important distinction between 

physical rarity and economic scarcity; that is, rarity is 

determined by “...an element’s relative physical abundance 

in some specified portion of the earth” and scarcity is determined by the “...cost of acquisition under given conditions 

of time and place.” The concepts of rarity and scarcity derive 

from different origins—rarity from the differentiation of the 

universe, over which humans have no control, and scarcity 

from many causes, of which human activity is often the sole 

determinant, such as monopolistic or oligopolistic market 

forces (Brooks, 1965, p. 22).

Three reasons why an element or mineral commodity 

may be scarce are rooted in geochemistry. First, rarity may 

result in scarcity; however, elements that have similar rarity 

as measured by average crustal abundance, such as cobalt 

and scandium, may have quite different scarcity as measured 

by market price. For example, in 2012, the price of cobalt 

was about 2.9 cents per gram ($13 per pound), whereas 

the price of scandium was $169 per gram for metal ingots 

(Gambogi, 2014; Shedd, 2014). Because of market prices, 

cobalt was rare but not scarce, whereas scandium was both 

rare and scarce.

A second factor affecting the scarcity of an element is 

the concentration of the element above its average crustal 

abundance that is required to create an ore deposit (where 

the mineral can be extracted at a profit). The concentration 

above average crustal abundance needed to enable profitable 

mining is also dependent upon the byproducts and coproducts 

present in the deposit, ore mineralogy, grain size, consolidation of the material to be mined, and deposit depth and 

location. These characteristics are not accounted for in a ratio 

of the grade of an element in a deposit to its average crustal 

abundance. The necessary concentration with respect to 

average crustal abundance to make mining profitable varies 

considerably. Brooks (1976, p. 149) credited former USGS 

Director V.E. McKelvey with pointing out that titanium has 

been mined from deposits having grades below its average 

crustal abundance, whereas, for antimony, the concentration 

has needed to be as high as more than 300,000 times its 

crustal abundance for mining to be economic.

Finally, a mineral commodity may be scarce because 

of the lack of technology needed to effectively and 

economically mine and process ore material into a usable 

(and marketable) product. The need for such technology is 

dictated by the intended use. If a mineral commodity can 

be used effectively while containing some impurities, then 

the absence of processing technology to separate impurities 

is not a factor that contributes to scarcity. If removal of 

impurities is necessary, then the absence of such technology 

increases scarcity (Brooks, 1965, p. 23–24).

In “United States Mineral Resources,” Erickson (1973) 

estimated the resource base—that is, the total amount of 

a mineral or metal in the earth (Lusty and Gunn, 2015, 

p. 266)—by using estimates of elemental concentrations in

Earth’s crust. In the present volume, the mineral commodities

included have crustal abundances that range from 5,650 parts

per million (ppm) (titanium) to 0.0007 ppm (rhenium). The

crustal abundance of the mineral commodities (elements)

discussed in this volume are listed in table A1.

Some of the chapters in this volume give different 

elemental abundances than those presented in table A1. This 

reflects that different sources were used for the information 

and, in some cases, abundances were for the upper continental crust versus the total crust. Rudnick and Gao (2003) 

reviewed current estimates of the composition of continental 

crust and discussed the methods employed to derive 

the estimates.

Mineral Commodities Selected for 

Inclusion in This Volume

The use of terms such as “critical,” deficient,” “essential,” and “strategic” to identify minerals or lists of minerals 

that are considered important by some observer(s) has a 

long and complicated history (appendix A1). In addition to 

implying a general level of importance, the word “critical” 

has been used to describe minerals or materials whose 

“criticality,” as quantified by various analytical approaches, 

has been determined to exceed some specified limit. Some of 

the numerous studies that have been conducted to quantify 

or rank the criticality of selected mineral commodities have 

focused on the broad resource needs of current and emerging 

technologies (National Research Council, 2008; European 

Commission, 2010), whereas others have focused on certain 

sectors only, such as the energy sector (Massachusetts 

Institute of Technology, 2010; U.S. Department of Energy, 

2011) or the national security sector (McGroarty and Wirtz, 

2012). As a result, the mineral commodities classified as 

critical by each of these studies differ, although many of 

the findings include some of the same mineral commodities, such as some or all of the rare-earth elements and 

platinum-group elements.

The 23 mineral commodities or commodity groups 

selected for coverage in this volume (antimony, barite, beryllium, cobalt, fluorite, gallium, germanium, graphite, hafnium, 

indium, lithium, manganese, niobium, platinum-group 

elements, rare-earth elements, rhenium, selenium, tantalum, 

tellurium, tin, titanium, vanadium, and zirconium) have 

been listed as critical and (or) strategic in one or more of the 

recent studies based on assessed risks to their supply and (or) 
Changes Since the Mid-1970s A7

Table A1. Crustal abundances of mineral commodities 

(elements) included in this volume.

[Source: Jefferson Science Associates, LLC (2014). ppm, part per million]

Element Symbol Crustal abundance 

(ppm)

Titanium Ti 5,650

Manganese Mn 950

Fluorine F 585

Barium Ba 425

Carbon C 200

Zirconium Zr 165

Vanadium V 120

Cerium Ce 66.5

Neodymium Nd 41.5

Lanthanum La 39

Yttrium Y 33

Cobalt Co 25

Scandium Sc 22

Lithium Li 20

Niobium Nb 20

Gallium Ga 19

Praseodymium Pr 9.2

Samarium Sm 7.05

Gadolinium Gd 6.2

Dysprosium Dy 5.2

Erbium Er 3.5

Ytterbium Yb 3.2

Hafnium Hf 3

Beryllium Be 2.8

Tin Sn 2.3

Europium Eu 2

Tantalum Ta 2

Germanium Ge 1.5

Holmium Ho 1.3

Terbium Tb 1.2

Lutetium Lu 0.8

Thulium Tm 0.52

Indium In 0.25

Antimony Sb 0.2

Selenium Se 0.05

Palladium Pd 0.015

Platinum Pt 0.005

Osmium Os 0.0015

Iridium Ir 0.001

Rhodium Rh 0.001

Ruthenium Ru 0.001

Tellurium Te 0.001

Rhenium Re 0.0007

impact of potential supply restrictions. They are viewed as 

critical to a broad range of existing and emerging technologies, 

renewable energy, and national security (National Research 

Council, 2008; Massachusetts Institute of Technology, 2010; 

U.S. Department of Energy, 2011). They all have specialized 

and important applications in high-tech industrial, energy, 

defense, and (or) medical sectors, often with no effective 

substitutes. For some, current production is largely limited 

to only one or a few countries (for example, cobalt, niobium, 

rare-earth elements, and platinum-group elements), making 

their supply vulnerable to potential influence and disruption by 

such factors as civil unrest, political changes, natural disasters, 

environmental issues, and market manipulation. For many, 

the United States currently has neither mine production nor 

significant identified deposits and is largely dependent on 

imports to meet its resource needs.

This volume does not cover all 65 mineral commodities 

and commodity groups presented in “United States Mineral 

Resources” because the global resource picture for many 

of the major base, ferrous, and precious metals is already 

relatively well understood and documented, and the information on future sources of supply has not changed significantly 

in the past 40 years. Many of the commodities or commodity 

groups covered in the present volume, on the other hand, 

were considered less significant in the early 1970s because 

of their limited uses, and the documentation on them was 

sparse. In this volume, greater emphasis is placed on the 

global distribution of resources of these minerals than was 

the case in “United States Mineral Resources.” In addition, 

for each mineral commodity, the text includes a discussion of 

the environmental issues related to its production, including 

current mitigation and remediation approaches to deal with 

these challenges.

Work on this volume began in 2013, and the writing and 

technical reviews of the chapters were completed between 

2014 and 2016. As a result, there are some differences 

between chapters, particularly with respect to the most 

recent information regarding identified resources, reserves, 

and production.

Changes Since the Mid-1970s

The status of global mineral resources and mining has 

changed significantly since the publication of “United States 

Mineral Resources.” The annual world production of most 

mineral commodities has increased markedly—in the case 

of the minerals covered in the chapters of this volume, the 

changes from 1973 to 2015 range from a modest 21 percent 

increase in annual tin production to a 2,800-percent increase 

(more than 29 times as much production!) for gallium. In light 

of this increased demand, concerns about possible weak links 

in the supply chains for mineral commodities have highlighted 

the need for reliable, regularly updated minerals information 
A8 Critical Mineral Resources of the United States—An Introduction

that is both broad in the scope of the mineral commodities 

covered and expansive in geographic reach. In addition, there 

is need for reliable assessments of the global distribution of 

mineral resources, their potential for supply disruption, and 

the environmental consequences of their production and use 

(Herrington, 2013).

Geologic knowledge and research technologies have also 

changed significantly since the publication of “United States 

Mineral Resources.” During the 1960s, the theory of plate 

tectonics had been formulated and most geologists had 

embraced it, but as of the early 1970s, the implications for 

ore-deposit geology were still only dimly perceived. Today, 

several plate-tectonic processes are seen to control the genesis 

of most mineral-deposit types, and such other factors as 

paleoclimate and time in Earth history are also recognized 

as important.

Geologic maps remain the main prerequisite in the 

exploration for ore deposits, and the past four decades have 

seen huge strides in the amount and quality of geologic 

map coverage of the world, especially in the less developed 

countries. New software developments have made it possible 

for all geologists to use the once-specialized toolkit of 

satellite remote sensing. Methods of isotopic analysis have 

vastly improved, and it is now possible to date many ore 

deposits accurately and precisely, so that their origins can 

be understood in terms of the regional geologic and tectonic 

history. It is also possible to analyze a tiny sample of rock or 

ore chemically for virtually any element in the periodic table, 

including those elements that are the focus of this volume. 

All these scientific advances have helped to bring down 

the cost of mineral exploration and raise the probability of 

success. Despite these advances, however, continued research 

is essential for future geologists to find, and engineers to 

develop, mine, and process, increasing amounts of mineral 

resources to meet future needs.

Furthermore, as discussed by Seal and others (this 

volume, chap. B), people, governments, and corporations are 

much more aware than before of the environmental impacts of 

mineral resource extraction and use. Environmental challenges 

for the future remain, such as increased mine-waste and water 

management issues, including those caused by the effects 

of climate change, and the aspect of more energy-intensive 

mining operations.

Back to the Future

Despite the many changes in the status of global mineral 

resources and mining and the advances in geologic knowledge 

during the past 40 years, much of the previous analysis—and 

cautions—about mineral supplies and knowledge of them 

remains relevant. Cameron’s (1973) statements about the 

U.S. mineral position, based on his analysis of U.S. mineral 

history and world mineral history since 1939, are perhaps 

even more pertinent today. McDivitt and Manners (1974, 

p. 10–11) wrote that in spite of the inescapable facts “that the 

total amount of each and every mineral is fixed and that use 

diminishes this given stock, ...the availability of minerals is 

expanded steadily by a growing knowledge of the world’s 

geology, by the falling real costs of transport that allow 

minerals to be moved over ever-increasing distances, and 

by the development of techniques that permit the processing 

of different types of ore, often without more costly effort.” 

They tempered expectations that these factors would keep 

major problems of inadequate mineral supplies at bay with 

cautions about how far into the future these expectations could 

be extended. At some point, increases in real costs for some 

minerals might result in changes in consumption, including 

the use of substitute materials.

Tilton (2003, p. 101–102) concluded that “depletion 

raises the specter of a world where resources are too costly 

to use rather than a world with no resources,” but that “the 

past is not necessarily a good guide to the future,” especially 

to the more distant future. He added that “More geologic 

information on the incidence and nature of mineral deposits, 

particularly subeconomic mineral deposits, could go a long 

way toward resolving this critical issue by providing useful 

insights on the nature and shape of cumulative supply curves” 

but that this information is not likely to become available soon 

because of the lack of economic incentives to gather information on deposits that are not currently profitable to develop 

(Tilton, 2003, p. 103).

Former mining industry executive Simon Strauss drew 

on his 1986 book, “Trouble in the Third Kingdom,” when he 

delivered a presentation titled “The Appetite for Minerals” 

at a seminar at the USGS National Center on April 18, 1989. 

Mr. Strauss concluded that continued success in addressing 

mineral supply problems could be expected by stimulation 

of domestic production; increases in recycling, conservation, 

and substitution; and the use of stockpiles. When asked if that 

view might indicate a lack of need for further research on 

the geology of mineral deposits and on mining and mineral 

processing technology, such as that done by the USGS and the 

U.S. Bureau of Mines, Mr. Strauss replied that the contributions of those institutions were an important reason for his 

optimistic outlook, but that the role of those institutions in 

helping ensure adequate past and future mineral supplies was 

widely appreciated and that they would continue to perform 

those roles. The closure of the U.S. Bureau of Mines in 

1996 indicates that the appreciation of the value of scientific 

research that provides the foundation for solving future 

resource problems should not be taken for granted.
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at https://pubs.er.usgs.gov/publication/cir831.]

U.S. Department of Energy, 2011, Critical materials strategy:

Washington, D.C., U.S. Department of Energy, 

DOE/PI-0009, 191 p. [Also available at http://energy.gov/

sites/prod/files/DOE_CMS2011_FINAL_Full.pdf.]

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1975: U.S. Geological Survey Professional Paper 940, 24 p.

U.S. Geological Survey, 2015a, Cobalt statistics [through 

2013; last modified January 30, 2015], in Kelly, T.D., and 

Matos, G.R., comps., Historical statistics for mineral and 

material commodities in the United States (2015 version): 

U.S. Geological Survey Data Series 140, 4 p., accessed 

March 13, 2016, at http://minerals.usgs.gov/minerals/pubs/

historical-statistics/. [Supersedes Open-File Report 01– 006.]

U.S. Geological Survey, 2015b, Mineral commodity summaries

2015: U.S. Geological Survey, 196 p. [Also available 

at http://minerals.usgs.gov/minerals/pubs/mcs/2015/

mcs2015.pdf.]

U.S. Geological Survey, 2016, Mineral commodity summaries 

2016: U.S. Geological Survey, 202 p., accessed March 7, 2016,

at http://minerals.usgs.gov/minerals/pubs/mcs/2016/

mcs2016.pdf.

U.S. Geological Survey National Mineral Resource Assessment

Team, 2000, 1998 assessment of undiscovered deposits 

of gold, silver, copper, lead, and zinc in the United States: 

U.S. Geological Survey Circular 1178, 21 p. [A Portable 

Document (PDF) recompilation of USGS Open-File Report 

96–96 and Circular 1178 was prepared by Paul Schruben 

and released as U.S. Geological Survey Open-File Report 

02–198 in 2002.] [Also available at http://pubs.usgs.gov/

circ/c1178/c1178.pdf.]

Zimmermann, E.W., 1951, World resources and industries—

A functional appraisal of the availability of agricultural and 

industrial materials (2d ed.): New York, N.Y., Harper & 

Brothers, 832 p.
Appendix A1 A11

Appendix A1. What is Meant by “Critical” Minerals (or Materials)?

Problems resulting from the lack of domestic production 

of certain mineral commodities became apparent in World War I 

when “Minerals that were essential to the war effort ... could 

not be produced in adequate amounts from domestic 

sources...” (Cameron, 1986, p. 254). These became known as 

“strategic minerals.” Charles Kenneth Leith has been credited 

with compiling the first of many lists of “strategic and critical 

minerals” when he was serving as a mineral adviser to the 

War Industries Board in 1917—this “unofficial” list was 

industrial in character, rather than defense oriented (Roush, 

1939, p. 10). The first “official” list was prepared under 

the auspices of the Supply Division of the General Staff on 

January 20, 1921 (Pehrson, 1944, p. 339). The plethora of lists 

that followed seems to bear out an observation of a university 

administrator that ranking (and listing) things is “a part of the 

DNA of America” (Pope, 2013, p. 31). Roush (1939, p. 2–10) 

documented 15 different lists of strategic and critical minerals 

and materials created from 1917 through 1939; there were 

both official lists and unofficial lists that categorized minerals 

as strategic (from the standpoints of industrial needs, defense 

needs, or both), critical (“those minerals which might be 

expected to develop a shortage with increased demand, but 

which, with careful control of consumption and proper stimulation of production, might be maintained on a self-supporting 

basis”), essential (“essential materials neither strategic nor 

critical”), and deficient (“in a major degree” and “to a lesser 

degree”). In addition, there were several changes in the stated 

meaning of the terms and differences in their application over 

time, depending upon the perspective of the person or groups 

of persons compiling the lists (Roush, 1939, p. 3).

Pehrson (1944, p. 339–340) attempted to set the record 

straight in this regard by defining strategic materials as those 

that were essential for defense and for which the United States 

relied on foreign supply sources, and critical materials as those 

that were essential to defense but less difficult to obtain because 

they were less essential or had domestic supply sources. 

In 1944, the Army and Navy Munitions Board approved a 

new definition of “strategic and critical materials” (with no 

distinction between the two adjectives) as “...those materials 

required for essential uses in a war emergency, the procurement 

of which in adequate quantities, quality, and time is sufficiently 

uncertain for any reason to require prior provision for the supply 

thereof.” In the Strategic and Critical Materials Stock Piling 

Act (50 U.S.C. §98), “The term ‘strategic and critical materials’ 

means materials that (A) would be needed to supply the military, industrial, and essential civilian needs of the United States 

during a national emergency, and (B) are not found or produced 

in the United States in sufficient quantities to meet such need.” 

Pehrson (1944, p. 341) documented the changes as to which 

minerals were classified as strategic, critical, essential, or 

strategic and critical in several lists since the first official list 

was compiled in January 1921. McKinstry (1948, p. xvii–xix) 

attributed the increased attention concerning strategic minerals 

in the 1940s to “geological economists” and observed that, in 

time of war, “almost all useful minerals become strategic” and 

information about where minerals occur and how soon and for 

how long they can be produced is “critical.”

The American Geological Institute “Dictionary of 

Geological Terms” (1962, p. 113, 475) defined critical 

[materials] and strategic [materials] separately. The Commerce 

Department representative on the staff of the 1978–79 

Presidential Review of Nonfuel Minerals Policy succinctly 

summarized the two-definition concept by saying, “Critical 

means you need it; strategic means you don’t have it!” 

(William J. Kaestner, U.S. Department of Commerce, oral 

commun., 1978). In contrast, Evans (1993, p. 10) stated, 

“A material needed for military purposes is considered strategic and a material is termed critical if future events involving 

its supply from abroad threaten to inflict serious harm on a 

nation’s economy” (DeYoung and others, 2006, p. 486).

The number of minerals classified as strategic, critical, 

vital, deficient, or strategic and critical (with frequent changes 

in definitions of the terms and as viewed from the perspectives 

of various stakeholders) increased from 4 in World War I, to 

9 in 1939, to 52 later in World War II (Hewett, 1959, p. 191). 

In 1974, the President’s Council on International Economic 

Policy reported on a list of 19 critical materials, but did not 

provide a definition of critical (Council on International 

Economic Policy, Executive Office of the President, 1974). 

A 1982 National Indicators System report on the domestic 

supply of critical minerals explained that classification of 

minerals as strategic or as critical changes over time across 

countries, industries, and users and that the determination of 

which minerals were “strategic and critical” was made by the 

executive branch of the Government. For the purposes of that 

report, the 15 “critical” minerals chosen were based on one or 

more of several considerations, including the large amounts 

used and resulting importance to the economy, strategic 

importance to national defense, special properties not readily 

found in other materials, reliance on imports for domestic 

consumption, and alloying properties (U.S. Bureau of Mines, 

1982, p. 2–3). In advocating a national geochemical census 

and increased support for basic research, Yoder (1982, p. 229) 

stated that all the minerals for which the United States was 

more than 50 percent import reliant in 1979 were critical to 

the economy and that they could be classified as “strategic” if 
A12 Critical Mineral Resources of the United States—An Introduction

they met one or more of four criteria describing their geologic 

availability, technology necessary to bring them to market in a 

timely fashion, the “possibility that unforeseen essential uses 

may develop,” and the reliability of foreign sources. When the 

National Critical Materials Council was established in 1984, 

the enabling legislation (30 U.S.C. §§1801–1811) specified a 

definition for “materials,” but not for “critical.” Some lists that 

use “critical” as a modifier explicitly restrict the definition of 

critical to that applied to a particular end use, such as “energycritical” minerals (American Physical Society and the Materials 

Research Society, 2011; U.S. Department of Energy, 2011).

Conflicting uses of terms to describe critical and (or) 

strategic minerals led Archer (1980, p. 2) to question whether 

the cobalt in the turbine blades of a jet engine was more strategic or critical than the petroleum-based fuel or the materials 

used in constructing the runway. He concluded that the term 

“strategic” was used by some as an impressive synonym for 

“important” and that “criticality” as measured by the part that 

a mineral plays in everyday life, industry, and defense, is not 

related to its source of supply.

When they addressed their question, “How critical are 

critical materials?”, Clark and Field (1985, p. 38–40) did 

not present a way to measure criticality but concluded, after 

examining statistics on net import reliance, that mechanisms 

to address short-term and long-term problems that might 

result from disruptions in mineral supplies were already in 

place (stockpiles, conservation, recycling, and development 

of alternative materials). Scientists in the Department of the 

Interior’s Office of Minerals Policy and Research Analysis had 

already begun to develop a new measure, the Critical Minerals 

Index (CMI). They recognized that the “imports market share” 

measure ignored the two primary determinants of mineral 

criticality—supply risk and economic importance—and was 

thus of little use to policymakers in focusing attention on 

supply problems for specific minerals (Adams and others, 

1979, p. 1–2). Two CMI components were identified for 

each mineral commodity—a Mineral Disruption Index (the 

likelihood of disruption of U.S. imports during the time period 

of consideration) and a Mineral Cost Index (the annual cost 

imposed by the import disruption; Adams and others, 1979, 

p. 4–5).

In 2008, these two components were used by the National

Research Council’s Committee on Critical Mineral Impacts 

on the U.S. Economy to develop a two-dimensional graphical 

approach to define and measure criticality, the “criticality 

matrix,” which was applied to several nonfuel mineral 

commodities (National Research Council, 2008). Subsequent 

studies have added a third dimension (environmental implications) to the matrix, considered the implications of specifying 

a time dimension, added net present value calculations to 

the economic value (cost) dimension, and incorporated 

the determinants of risk into the vulnerability dimension 

(Erdmann and Graedel, 2011; Graedel and others, 2012; 

Glöser and others, 2015; Graedel and Reck, 2015; Mayer and 

Gleich, 2015). Studies that have applied these approaches to 

produce quantitative measures of criticality have incorporated 

existing data series as surrogates for the components; 

for example, Silberglitt and others (2013, p. 4) used the 

Herfindahl-Hirschman Index (a measure of the concentration 

of market power and international trade) as a measure of 

the geographic concentration of mineral production and the 

World Bank’s Worldwide Governance Indicators (a measure 

of the quality of governance in six dimensions) as a measure 

of supply risk where reliability of supply might be affected by 

political instability, government control of mineral production, 

or restrictive trade policies.

In 2016, the National Science and Technology Council 

(NSTC) issued a report that provided and applied a systematic 

methodology to screen for potentially critical minerals. 

The report summarized an interagency effort to develop 

a screening process to assess “potential criticality” based 

on three fundamental indicators: supply risk, production 

growth, and market dynamics. The screening was applied 

to 78 minerals or mineral commodities for each year from 

1996 through 2013. This study defined critical minerals to be 

“those that have a supply chain that is vulnerable to disruption, and that serve an essential function in the manufacture 

of a product, the absence of which would cause significant 

economic or security consequence,” and strategic minerals as 

“a subset of critical minerals and are those that are essential 

for national security applications” (National Science and 

Technology Council, 2016, p. ix).

In wrestling over the question of “Which materials are 

‘critical’ and which are ‘strategic’?”, Simandl and others 

(2015, p. 59) concluded that “misunderstanding, miscommunications and potentially misrepresentations” can result 

from the lack of consistency in use of the terms “critical” and 

“strategic” and that “Which materials are considered critical 

depends to a large extent on the priorities and objectives of 

the organization or country that commissions the study.” 

This observation about the necessarily subjective nature of 

determining criticality—depending upon who commissions a 

study about critical minerals, makes a list of critical minerals, 

or devises a process to define and measure criticality of 

minerals—has been stated explicitly in many studies and 

is implicit in all the rest. The economic concept of utility, 

how useful or satisfying a good, service, or action is to an 

individual, underlies the formation of priorities and objectives. 

In fact, as noted by Sears (2011), “it is impossible to compare 

utility levels of different people” and thus “modern utility 

theory does not allow the economist to combine individual 

utilities into one number for all society.” Recognition of 

the role of perspective in determining what is “critical” or 

important and the attendant limitations of such determinations 

is a critical element in understanding what is meant by 

“critical” minerals.
Appendix A1 A13

Appendix A1 References Cited

Note: All Web links listed were active as of the access date but may no 

longer be available.

Adams, R.L., White, B.A., and Grichar, J.S., 1979, Developing a critical minerals index—A pilot study: U.S. Department of the Interior, Office of Minerals Policy and Research 

Analysis, July, 52 p.

American Geological Institute, 1962, Dictionary of geological 

terms (3d ed.): Garden City, N.Y., Dolphin Books, 545 p.

American Physical Society and the Materials Research 

Society, 2011, Energy critical elements—Securing 

materials for emerging technologies: Washington, D.C., 

American Physical Society, 24 p., accessed March 20, 2016,

at https://www.aps.org/policy/reports/popa-reports/upload/

elementsreport.pdf.

Archer, A.A., 1980, Definition of strategic materials and 

the means currently available to the United Kingdom to 

identify strategic sources and requirements, in Jones, M.J., 

ed., Availability of strategic minerals—Proceedings of the 

National Symposium on the Availability of Strategic Minerals, London, United Kingdom, November 20–21, 1979: 

London, United Kingdom, The Institution of Mining and 

Metallurgy, p. 1–3.

Cameron, E.N., 1986, At the crossroads—The mineral problems of the United States: New York, N.Y., John Wiley & 

Sons, 320 p.

Clark, J.P., and Field, F.R., III, 1985, How critical are critical 

minerals?: Technology Review (Massachusetts Institute of 

Technology), v. 88, no. 6, August/September, p. 38–46.

Council on International Economic Policy, Executive Office 

of the President, 1974, Special report—Critical imported 

materials: Washington, D.C., U.S. Government Printing

Office, 49 p. plus 1 appendix. [Also available at 

https://catalog.hathitrust.org/Record/000035158.]

DeYoung, J.H., Jr., McCartan, Lucy, and Gambogi, Joseph, 

2006, What’s been (and what will be) strategic—My metal 

or your paint?, in Reid, J.C., ed., Proceedings of the 42d 

Forum on the Geology of Industrial Minerals (Asheville, 

N.C., May 7–13, 2006): Raleigh, N.C., North Carolina

Geological Survey Information Circular 34, p. 482–497,

CD–ROM.

Erdmann, Lorenz, and Graedel, T.E., 2011, Criticality of 

non-fuel minerals—A review of major approaches and 

analyses: Environmental Science and Technology, v. 45, 

no. 18, p. 7620–7630, accessed March 27, 2016, at 

http://dx.doi.org/10.1021/es200563g.

Evans, A.M., 1993, Ore geology and industrial minerals—

An introduction (3d ed.): Oxford, United Kingdom, Blackwell Science, 389 p.

Glöser, Simon, Espinoza, L.T., Gandenberger, Carsten, and 

Faulstich, Martin, 2015, Raw material criticality in the 

context of classical risk assessment: Resources Policy, 

v. 44, p. 35–46. [Also available at http://dx.doi.org/10.1016/

j.resourpol.2014.12.003.]

Graedel, T.E., Barr, Rachel, Chandler, Chelsea, Chase, Thomas,

Choi, Joanne, Christoffersen, Lee, Friedlander, Elizabeth, 

Henly, Claire, Jun, Christine, Nassar, N.T., Schechner, Daniel, 

Warren, Simon, Yang, Man-yu, and Zhu, Charles, 2012, 

Methodology of metal criticality determination: Environmental Science and Technology, v. 46, no. 2, p. 1063–1070, 

accessed March 27, 2016, at http://dx.doi.org/10.1021/

es203534z.

Graedel, T.E., and Reck, B.K., 2015, Six years of criticality

assessments—What have we learned so far?: Journal

of Industrial Ecology, accessed March 27, 2016, at 

http://dx.doi.org/10.1111/jiec.12305.

Hewett, D.F., 1959, Charles Kenneth Leith, January 20, 

1875–September 13, 1956—A biographical memoir: 

Washington, D.C., National Academy of Sciences, p. 180–204, 

accessed March 16, 2016, at http://www.nasonline.org/

member-directory/deceased-members/20001553.html.

Mayer, Herbert, and Gleich, Benedikt, 2015, Measuring criticality of raw materials—An empirical approach assessing 

the supply risk dimension of commodity criticality: Natural 

Resources, v. 6, p. 56–78, accessed March 27, 2016, at 

http://dx.doi.org/10.4236/nr.2015.61007.

McKinstry, H.E., 1948, Mining geology: Englewood Cliffs, 

N.J., Prentice-Hall, Inc., 680 p.

National Research Council, 2008, Minerals, critical minerals, 

and the U.S. economy: Washington, D.C., The National 

Academies Press, 245 p. [Also available at http://www.nap.edu/

catalog/12034/minerals-critical-minerals-and-the-us-economy.]

National Science and Technology Council, 2016, Assessment 

of critical minerals—Screening methodology and initial 

application: Washington, D.C., Subcommittee on Critical 

and Strategic Mineral Supply Chains of the Committee on 

Environment, Natural Resources, and Sustainability of the 

National Science and Technology Council, March, 47 p.

Pehrson, E.W., 1944, What are strategic and critical materials?:

Mining and Metallurgy, v. 25, no. 451, July, p. 339–341.

Pope, Justin, 2013, A look behind the rankings: Michigan 

Alumnus, Early Fall, p. 30–35.
A14 Critical Mineral Resources of the United States—An Introduction

Roush, G.A., 1939, Strategic mineral supplies: New York, 

N.Y., McGraw-Hill Book Company, Inc., 485 p.

Sears, Louis, 2011, Jargon alert—Utility: Region Focus 

(Federal Reserve Bank of Richmond), v. 15. no. 3, 

third quarter, p. 10., accessed March 7, 2016, via 

https://www.richmondfed.org/publications/.) [Also available 

at https://www.richmondfed.org/publications/research/

region_focus/2011/q3/q3.]

Silberglitt, Richard, Bartis, J.T., Chow, B.G., An, D.L., and 

Brady, Kyle, 2013, Critical materials—Present danger 

to U.S. manufacturing (Research report prepared for the 

National Intelligence Council): Santa Monica, Calif., 

RAND Corp., document RR–133–MC, 46 p., accessed 

March 9, 2016, at http://www.rand.org/pubs/research_

reports/RR133.html.

Simandl, G.J., Akam, Carlee, and Paradis, Suzanne, 2015, 

Which materials are “critical” and which are “strategic”?: 

Industrial Minerals, no. 578, December, p. 57–59.

U.S. Bureau of Mines, 1982, The domestic supply of critical 

minerals: U.S. Bureau of Mines, National Indicators System 

Report Number 9, May, 98 p.

U.S. Department of Energy, 2011, Critical materials strategy:

Washington, D.C., U.S. Department of Energy, 

DOE/PI-0009, 191 p. [Also available at http://energy.gov/

sites/prod/files/DOE_CMS2011_FINAL_Full.pdf.]

Yoder, H.S., Jr., 1982, Strategic minerals—A critical research 

need and opportunity: Proceedings of the American Philosophical Society, v. 126, no. 3, p. 229–241.
Environmental Considerations Related to 

Mining of Nonfuel Minerals

By Robert R. Seal II, Nadine M. Piatak, Bryn E. Kimball, and Jane M. Hammarstrom

Chapter B of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–B

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter B

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................B1

Introduction..................................................................................................................................................B1

Modern Regulatory Framework................................................................................................................B3

Investigations Supporting Permit Applications .....................................................................................B6

The Quest for Best Practices ....................................................................................................................B7

Mine Permitting and Planning Process Overview ..............................................................................B11

Geoenvironmental Models ......................................................................................................................B12

Future Trends.............................................................................................................................................B13

Summary.....................................................................................................................................................B13

References Cited.......................................................................................................................................B14

Suggested citation:

Seal, R.R., II, Piatak, N.M., Kimball, B.E., and Hammarstrom, J.M., 2017, Environmental considerations related to 

mining of nonfuel minerals, chap. B of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical 

mineral resources of the United States—Economic and environmental geology and prospects for future supply: 

U.S. Geological Survey Professional Paper 1802, p. B1– B16, http://dx.doi.org/10.3133/pp1802B.
Environmental Considerations Related to Mining of 

Nonfuel Minerals

By Robert R. Seal II, Nadine M. Piatak, Bryn E. Kimball, and Jane M. Hammarstrom

Abstract

Throughout most of human history, environmental 

stewardship during mining has not been a priority partly 

because of the lack of applicable laws and regulations and 

partly because of ignorance about the effects that mining can 

have on the environment. In the United States, the National 

Environmental Policy Act of 1969, in conjunction with related 

laws, codified a more modern approach to mining, including 

the responsibility for environmental stewardship, and provided 

a framework for incorporating environmental protection into 

mine planning. Today, similar frameworks are in place in 

the other developed countries of the world, and international 

mining companies generally follow similar procedures 

wherever they work in the world. The regulatory guidance 

has fostered an international effort among all stakeholders to 

identify best practices for environmental stewardship.

The modern approach to mining using best practices 

involves the following: (a) establishment of a pre-mining 

baseline from which to monitor environmental effects during 

mining and help establish geologically reasonable closure 

goals; (b) identification of environmental risks related to 

mining through standardized approaches; and (c) formulation 

of an environmental closure plan before the start of mining. 

A key aspect of identifying the environmental risks and 

mitigating those risks is understanding how the risks vary 

from one deposit type to another—a concept that forms the 

basis for geoenvironmental mineral-deposit models.

Accompanying the quest for best practices is the goal 

of making mining sustainable into the future. Sustainable 

mine development is generally considered to be development that meets the needs of the present generation without 

compromising the ability of future generations to meet their 

own needs. The concept extends beyond the availability of 

nonrenewable mineral commodities and includes the environmental and social effects of mine development.

Global population growth, meanwhile, has decreased the 

percentage of inhabitable land available to support society’s 

material needs. Presently, the land area available to supply the 

mineral resources, energy resources, water, food, shelter, and 

waste disposal needs of all Earth’s inhabitants is estimated to 

be 135 square meters per person. Continued global population 

growth will only increase the challenges of sustainable mining.

Current trends in mining are also expected to lead to 

new environmental challenges in the future, among which 

are mine-waste management issues related to mining larger 

deposits for lower ore grade; water-management issues related 

to both the mining of larger deposits and the changes in 

precipitation brought about by climate change; and greenhouse 

gas issues related to reducing the carbon footprint of larger, 

more energy-intensive mining operations.

Introduction

The development of mining was coincident with the 

beginning of the Stone Age when, more than 2.5 million 

years ago, ancestors of the human race first began fashioning 

tools out of stone. At the time, survival was likely the highest 

priority, and the health of the environment and the effects 

that mining had on it were likely of little, if any, concern. The 

world was probably considered vast and able to tolerate the 

seemingly minor effects of mining. This attitude has persisted 

for most of human history.

Early cultural development was punctuated by advances 

in mining and metallurgy (the science of extracting metals from 

ores). The Stone Age gave way to the Bronze Age roughly 

5,300 years ago when use of weapons and implements made 

of bronze—an alloy of copper and tin—became widespread. 

The manufacture of bronze required not only the mining of 

copper and tin, but the smelting of it; that is, the extraction 

of the metal from ore minerals using heat. The Bronze Age 

was followed by the Iron Age roughly 3,200 years ago. The 

Iron Age is defined by the widespread use of weapons and 

implements made of iron. The smelting of iron ores requires 

much higher temperatures than does the smelting of copper or 

tin. Cultural milestones in human history have been accompanied by advances in mining and metallurgical methods and 

technologies, even down to the present day, and the need for 

mining and metallurgy is likely to endure far into the future.
B2 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

Evidence of the historical or prehistoric influence of 

mining on the environment is limited, and indications of 

significant societal concern about the environmental impacts 

of mining are even more scarce. Sediment cores from peat 

bogs, estuaries, lakes, and ice cores indicate that the dispersion of metals from mining and early smelting date back 

at least 5,000 years (Shotyk and others, 1996; Rosman and 

others, 1997; Leblanc and others, 2000; Lee and others, 2008; 

Lottermoser, 2010). In Georgius Agricola’s landmark treatise 

on mining and metallurgy titled “De Re Metallica” [On the 

Nature of Metals (Minerals)], which was published in 1556, 

a year after his death, the author notes that “...when the ores 

are washed, the water which has been used poisons the brooks 

and streams, and either destroys the fish or drives them away” 

(Agricola, 1950/1556).

In the United States, the earliest significant court case 

involving a dispute about the environmental impacts of mining 

concerned hydraulic mining of gravels in the Mother Lode 

gold belt of the Sierra Nevada Mountains. Hydraulic mining 

uses high-pressure “cannons” to direct water to erode ancient 

stream gravels and move the eroded sediment into channels 

where gold can be efficiently separated. Farmers in the Sacramento Valley of California filed lawsuits against hydraulic 

mining companies because of the effects that the sediment 

from the hydraulic gold mining was having on farmland 

downstream from the mines. This legal battle culminated 

with the Sawyer Decision in 1884, which effectively banned 

hydraulic mining (Kelley, 1956).

In the Eastern United States, significant legal action 

surrounded ore processing practices in the Copper Basin (or 

Ducktown Basin) mining district in Tennessee (Quinn, 1993). 

Numerous lawsuits filed between 1890 and 1930 focused on 

air pollution initially related to the open roasting of sulfide 

ores as preparation for copper smelting and, later, to the 

smelting itself. The released sulfur dioxide resulted in acid 

rain throughout the surrounding 250-square-kilometer (km2

)

area, which destroyed much of the vegetation and led to 

extensive soil erosion. The long series of lawsuits ultimately 

reached the U.S. Supreme Court, which ruled against the 

mining companies. The Ducktown case was especially 

noteworthy in that it was the first air pollution case to be 

heard by the U.S. Supreme Court. The numerous lawsuits 

had the positive effect of encouraging the development of 

acid converters, which capture sulfur dioxide emissions from 

roasting and smelting and convert them to sulfuric acid, which 

is a valuable byproduct.

Today, the National Environmental Policy Act of 1969 

(NEPA) (42 U.S.C. §4321 et seq.) serves as the basic national 

charter for environmental protection. Its passage represents 

a milestone in the history of environmental protection with 

respect to mining practices, and its effect has been significant 

for society, the mining industry, and the environment.

The realization that the environmental effects of mining 

done without regard for the environment cannot be absorbed 

by nature has been building for at least the past several 

centuries. The problem can be conceptualized using the idea 

of a “support square,” as described by Skinner (1989), and 

amplified by Rutland (1997). A support square is equal to the 

hypothetically inhabitable area of the world divided by the 

population of the world; hence, the size of a support square is 

determined by the global population density. A support square 

represents the inhabitable land area needed to support all the 

resource needs of one human being. These needs include 

mineral resources, food, water, shelter, recreation, and waste 

disposal. If areas permanently covered by ice and areas above 

5,000 meters (m) are excluded, the inhabitable area of the 

world is roughly 133,000,000 km2

.

Figure B1 illustrates how the size of the support square 

has changed over time as the population density of the planet 

has increased. Beginning with the Bronze Age, roughly 

5,300 years ago (3300 B.C.), when the population of the 

world is estimated to have been 14 million people, each 

support square would have covered an area of 9.5 km2, or be 

approximately 3,080 m on a side. By the time Christopher 

Columbus discovered America (1492), the side of the support 

square was reduced to 560 m. When the Declaration of 

Independence was signed in 1776, the square was down to 

400 m on a side. Today, the support square is only 135 m on 

a side and, by 2050, more than 15 m is projected to be lost, 

leaving a square of only 119 m on a side to support an individual. In other words, society currently has only 11 percent 

of the per capita land area that it had in 1776, 6 percent of the 

per capita land area that it had when Columbus discovered 

America, and 0.2 percent of the per capita land area that 

it had at the beginning of the Bronze Age. This trend will 

continue with the currently projected population growth. 

Improved environmental stewardship during mining and 

all other extractive activities is essential to maintaining the 

quality of the land area available for use.

Responsible environmental stewardship related to 

mining is embodied in the concept of “sustainable mining.” 

The term sustainable mining strikes many as an oxymoron, 

because it is commonly used in reference to the mining of 

nonrenewable mineral or energy resources. Sustainable 

development, or sustainability, however, is generally defined 

as the ability to meet current needs without hindering future 

generations from meeting their needs (Brundtland, 1987). 

Sustainable practices for mining have come to focus on the 

environmental, economic, and social well-being in areas 

where mining is proposed or is taking place. The environmental part typically includes aspects of water management, 

solid-waste management, reagent use and recycling, and 

greenhouse gas emissions (Mudd, 2008, 2010). This chapter 

presents an overview of environmental practices related to 

the permitting and development of active and new mines, 

and provides a context for the environmental considerations 

sections of the commodity-focused chapters in this book.
Modern Regulatory Framework B3

0

2

4

6

8

10

1

3

5

7

9

11

30003500 2500 2000

B.C. A.D.

1500 1000 500 0 500 1000 1500 2000 2500

Population, in billions

Year

0

500

1,000

1,500

2,000

2,500

3,000

3,500

4,000

Support square edge

Population

Beginning of Bronze AgeColumbus discovers AmericaDeclaration of IndependencePresent day

Support square edge, in meters

Environmental Overview figure 1

Figure B1. Graph showing population 

growth and the change in supporting 

land area from 3500 B.C. to 2100 A.D., 

with projections to 2050. The support 

square edge (in meters [m]) represents 

the amount of inhabitable land available 

to support each person alive on Earth. In 

2016, the hypothetical support square was 

135 m on a side; by 2050, the projected 

support square will have decreased to 

119 m on a side. Based on projections by 

Skinner (1989) and Rutland (1997).

Modern Regulatory Framework

The regulatory framework for mine permitting in the 

United States is similar to that of other developed nations, but 

mining regulations in developing nations have tended to be 

absent or relatively weak. With globalization, however, less 

developed countries, such as Chile, where mining is a significant component of the gross domestic product, are reviewing 

and extending existing regulations to become more aligned 

with those of global partners, such as the United States, 

Canada, and the European Union (Espejo and others, 2012). 

Furthermore, major international mining companies generally 

follow similar approaches wherever they work, regardless of 

the presence or absence of local regulations. The following 

describes the development of the modern regulatory framework in the United States.

The National Environmental Policy Act of 1969 (NEPA), 

which was signed into law on January 1, 1970, serves as a 

basic national charter for environmental protection. It led 

to the establishment of the U.S. Environmental Protection 

Agency (EPA) in December 1970. The law provides the 

framework that Federal Government agencies use to evaluate 

the environmental effects of proposed Federal actions, such 

as the issuance of permits for mining or other industrial 

activities. An example of a specific action related to mining 

includes the issuing of permits for such mining-related 

activities as water withdrawal, exploratory drilling, wetland 

and (or) stream modifications, the transport of chemicals, 

and mine construction, among others. The framework put 

in place with the passage of the NEPA is known as the 

“NEPA process”—a process that is used to evaluate the 

environmental effects of Federal projects or actions, including 

alternative actions. The NEPA process is a staged approach in 

which, in the absence of a categorical exclusion, an environmental assessment (EA) is performed and results in either a 

“finding of no significant impact” (FONSI) or a determination 

that an environmental impact statement (EIS) is required. 

A categorical exclusion classifies the action as not having 

a significant effect on environmental quality. An EA is a 

preliminary evaluation of the potential magnitude of environmental effects and alternatives. The identification in the EA of 

a potential for significant environmental impact leads to the 

requirement for an EIS. An EIS is a detailed assessment of 

potential environmental impacts and alternatives. A Federal 

land-management or regulatory agency may skip the initial 

steps of the NEPA process if it feels that a significant potential 

for adverse environmental impact exists and may proceed 

directly to requiring the preparation of an EIS. Proposed 

mining projects that require the action of a Federal agency 

typically result in the requirement for an EIS.

The passage of the NEPA led to the passage of additional 

legislation that provides the modern policy for environmental 

protection associated with mining and other commercial and 

industrial activities. In one example, a previous law—the 

General Mining Law of 1872, which had been enacted to 

promote development and settlement of publicly owned lands 

in the Western United States—had provisions for environmental 

protection; in follow-on legislation, the Federal Land Policy and 

Management Act of 1976 (43 U.S.C. §1701 et seq.) was passed 

to address such issues as the requirement for mine reclamation 
B4 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

and its financial assurance. The NEPA process relies on these 

other laws and regulations to ensure that the potential for 

significant environmental impacts is addressed. The major 

laws relevant to proposed mining projects commonly include 

the Clean Water Act (33 U.S.C. §1251 et seq.), the Safe 

Drinking Water Act (42 U.S.C. §300f et seq.), the Clean Air 

Act (42 U.S.C. §7401 et seq.), and the Endangered Species 

Act (16 U.S.C. §1531 et seq.), although many others may 

be relevant. Salient features of these acts are summarized 

in table B1. The Resource Conservation and Recovery Act 

(1976) (42 U.S.C. §6901 et seq.), which governs the disposal 

of solid and hazardous wastes, specifically exempts waste 

generated during the extraction, beneficiation, and processing 

of minerals from regulation as hazardous waste.

Table B1. Summary of selected Federal laws relevant to mine permitting.

[EPA, U.S. Environmental Protection Agency]

Act (abbreviation) Year 

signed Purpose Implementation

National Environmental Policy Act of 1969 (NEPA)

(42 U.S.C. §4321 et seq.), and amendments

https://www.gpo.gov/fdsys/pkg/USCODE-2014-title42/

pdf/USCODE-2014-title42-chap55.pdf

See also: http://www2.epa.gov/laws-regulations/

summary-national-environmental-policy-act

and

http://www.epa.gov/compliance/nepa/

1970 Promotes efforts that will prevent 

or eliminate damage to the 

environment and biosphere 

and stimulate the health and 

welfare of humankind by 

providing information to

the public.

EPA reviews documents to 

determine if the environmental 

alternative chosen will affect 

the environment and to 

determine if the agency that 

wrote the documents provided 

this information to the public.

Clean Water Act (CWA) (33 U.S.C. §1251 et seq.), 

and amendments

https://www.gpo.gov/fdsys/pkg/USCODE-2014-title33/

pdf/USCODE-2014-title33-chap26.pdf

See also: http://www2.epa.gov/laws-regulations/

summary-clean-water-act

1972 Restores and maintains the 

chemical, physical, and

biological integrity of the

Nation’s waters; also

protects wetlands.

EPA provides guidance to States 

and tribes on the concentrations of metals that will not 

harm aquatic life or human 

health so State and tribal 

governments can set standards.

Safe Drinking Water Act of 1974 (SDWA) 

(42 U.S.C. §300f et seq.), and amendments

https://www.gpo.gov/fdsys/pkg/USCODE-2014-title42/

pdf/USCODE-2014-title42-chap6A.pdf

See also: http://www2.epa.gov/laws-regulations/

summary-safe-drinking-water-act

and

http://water.epa.gov/lawsregs/rulesregs/sdwa/

index.cfm

1974 Ensures safe drinking water

for the public.

EPA sets legal limits on the 

levels of certain contaminants 

in treated drinking water and 

provides oversight for States 

and tribes for implementing 

these standards.

Clean Air Act (CAA) (42 U.S.C. §7401 et seq.), 

and amendments

https://www.gpo.gov/fdsys/pkg/USCODE-2014-title42/

pdf/USCODE-2014-title42-chap85.pdf

See also: http://www2.epa.gov/laws-regulations/

summary-clean-air-act

and

http://www.epa.gov/air/caa/

1963 Ensures clean air at a national 

level by controlling air 

pollution.

EPA develops and enforces 

regulations to protect 

the public from airborne 

contaminants.

Endangered Species Act (16 U.S.C.§1531 et seq.), 

and amendments

https://www.gpo.gov/fdsys/pkg/USCODE-2014-title16/

pdf/USCODE-2014-title16-chap35.pdf

See also: http://www2.epa.gov/laws-regulations/

summary-endangered-species-act

and

http://www.fws.gov/endangered/laws-policies/

1973 Ensures that any action authorized, funded, or carried out by 

a Federal agency is not likely 

to jeopardize the continued 

existence of any listed species 

or result in the destruction or 

adverse modification of critical 

habitat of such species.

The U.S. Fish and Wildlife 

Service (FWS) or the National 

Marine Fisheries Service 

(NMFS), depending upon 

location of the site, evaluates 

potential risks to threatened 

and endangered species 

through consultation with 

other Federal agencies.
Modern Regulatory Framework B5

The Comprehensive Environmental Response, Compensation, and Liability Act (CERCLA) (42 U.S.C. §9601 et seq.), 

commonly known as “Superfund,” can apply to mines during 

production or after closure. The EPA conducts and supervises 

investigations and cleanup actions at mine sites where 

hazardous substances have been released into the environment 

or when there is a threat of release. The CERCLA authorizes 

the following two types of response actions: long-term (remedial) cleanup and short-term (removal) cleanup. Cleanups 

may be done by the EPA; by other Federal agencies, such as 

the Bureau of Land Management or the U.S. Forest Service; 

States or municipalities; or the mining company or parties 

responsible for the contamination.

A key aspect of the NEPA process for mine permitting 

is the prediction of potential environmental impacts related 

to a proposed mining project as well as predictions regarding 

environmental conditions after mine closure. Information 

about the environmental characteristics of mineral deposits, 

mines, and resource extraction is available from a variety of 

sources; however, the types and sources of this information 

should be considered in an appropriate context. All information has value, but its usefulness has limitations. The types of 

information from various sources include studies of historical 

abandoned mines, studies in support of mine permit applications designed to “predict” mine waste behavior, studies of 

modern mines permitted under the NEPA or similar environmental regulations, studies of mines that began operation prior 

to NEPA, and studies of mines in developing countries.

Information from historical abandoned mines is useful for 

several reasons. In many cases, they represent “natural experiments” that have been underway for decades, centuries, or even 

millennia—time scales that are impossible to reproduce in the 

laboratory. Many of these historical mines were developed with 

no consideration for environmental protection because that was 

not a requirement for mine permitting, and because the full 

environmental ramifications of mining activities probably were 

not fully appreciated. As such, these sites may reflect “worst 

case” scenarios. Information from these sites has limitations, 

however, because certain deposit types may no longer be mined 

for specific mineral commodities, or mining and ore-processing 

methods may have changed. Thus, the nature of modern mine 

waste can be quite different from that at historical mines.

For example, the abandoned Elizabeth copper mine in 

Vermont was initially operated in the early to middle 1800s 

as a source of copperas (ferrous sulfate heptahydrate—

FeSO4•7H2O), which served a variety of uses—as a disinfectant, dye, as dye mordant, and as a wood preservative. 

Copperas production ceased at the Elizabeth Mine in the 

1880s as less expensive sources became available from the 

steel industry (Kierstead, 2001). The copperas was produced 

by controlled roasting of pyrrhotite (Fe1–xS) accompanied by 

leaching of the resulting ferrous sulfate, which was then reconstituted as solid crystals in evaporation vats. When built in the 

1800s, these copperas roast beds were designed to encourage 

the oxidation of sulfide minerals and to drain leachate. These 

features contributed to the environmental legacy of the 

abandoned Elizabeth copper mine because acid mine drainage 

is the result of the oxidation of sulfide minerals, principally 

pyrite (FeS2) and pyrrhotite, in the presence of water. Surface 

water and groundwater from the historical copperas roast 

beds had the worst water quality of all waters at the site in 

terms of low pH and high concentrations of potentially toxic 

dissolved metals (Seal and others, 2001). In contrast, seepage 

at the base of a mill tailings pile that was produced using 

more modern techniques from 1942 to 1958 (the pile contains 

fine-grained reject material from which the ore minerals have 

been removed) and deposited lower in the stream valley, had 

near neutral pH, and the water-quality issues were restricted to 

high concentrations of dissolved iron and sulfate. Therefore, 

for the same ore, historical processing techniques and modern 

processing techniques may result in significantly different 

environmental impacts.

Other sources of useful information about the environmental characteristics of ore deposits and mines are the 

environmental impact statements and related documents 

associated with mine permit applications. These documents 

include geochemical and other data about pre-mining conditions, which are crucial for environmental monitoring during 

and after mining and for setting environmental closure goals. 

They also contain information about the environmental 

characteristics of expected mine waste, the acid-generating 

potential of various types of solid mine wastes, and the leachability of trace elements from these wastes. These documents 

make predictions about future environmental conditions to aid 

in the permit decisionmaking process, both for scenarios in 

which mitigation measurements are successfully implemented 

and those in which they are not implemented.

Case studies of modern mines in the United States 

permitted under the NEPA process are scarce in the scientific 

literature. Instead, some data may be found in publicly 

available EISs associated with mine expansion projects; 

for example, the Aqqaluk project at the Red Dog zinc mine 

in Alaska (Tetra Tech, Inc., 2009). These sources generally 

constitute representative information about the environmental 

characteristics of active mines. Mining operations in many 

developing countries do not adhere to the same standards 

followed in the United States and by major international 

mining companies worldwide, however. Therefore, environmental data from developing countries may not be an accurate 

reflection of international consensus on “best practices” for 

environmental protection.

Even more challenging is establishing a proper context 

for the environmental information available for historical 

mines that started mining before NEPA but that have 

continued to produce into more recent times. It is often difficult to distinguish historical legacy issues from effects related 

to modern mining activity, especially if data about specific 

sampling sites are not reported in sufficient detail. The issue 

of distinguishing legacy environmental signatures is especially 

problematic when trying to establish closure goals for mine 

sites that began operations before the era of more-extensive 

regulation (Alpers and Nordstrom, 2000; Nordstrom, 2008).
B6 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

Investigations Supporting Permit 

Applications

Mine permitting is a prescribed process under the NEPA. 

The NEPA process as applied to mining places an emphasis on 

protecting water resources through the Clean Water Act. The 

EPA guidance is summarized in “EPA and Hardrock Mining—

A Source Book for Industry in the Northwest and Alaska” 

(U.S. Environmental Protection Agency, 2003). The source 

book outlines an approach to evaluating the potential risks 

to water quality posed by mining. A key distinction between 

mining and other types of industrial, commercial, or municipal 

development is that a mine needs to be adapted to the site 

conditions where the ore deposit occurs. In contrast, factories, 

shopping malls, airports, and schools tend to be built on tracts 

of land that meet a variety of specific criteria to ensure safe 

development and protection of the environment and the public.

The EPA source book has largely been adopted by the 

international mining community to serve as an outline for premining environmental studies. It also forms the foundation of 

the following discussion. The Source Book describes a variety 

of data collection, testing, mine design, and data analysis 

methods necessary to accomplish a rigorous assessment. The 

approach requires that, as part of the permitting process, premining baseline conditions must be documented, ecosystem 

and human health risks must be assessed, and predictions must 

be made about post-closure environmental conditions through 

a mining impact assessment. Specific topics include impacts 

to surface water and groundwater hydrology, impacts to water 

quality, impacts to aquatic resources, and impacts to wetlands. 

Ecological and human health risks associated with nonaqueous 

pathways, such as windblown dusts, are also considered.

Mining operations have the potential to affect both 

surface-water and groundwater hydrology at the mine site 

and within the broader watershed. Therefore, it is essential 

to evaluate whether or not impacts will be expected and 

what their nature and extent will be. Necessary input for this 

assessment includes meteorological and hydrologic data and 

water-management plans for operational and post-closure 

phases, with the goal of determining a detailed water balance 

for the project. Meteorological and hydrologic data are best 

collected on a regular basis to capture seasonal and year-toyear variations and to characterize the duration and intensity of 

storm events. These pre-mining data provide a baseline from 

which to monitor changes brought about by mining and other 

effects, such as climate change. Surface-water and groundwater resources are required to be thoroughly characterized, 

including the establishment of surface-water monitoring sites 

and the installation of monitoring wells. Water-management 

plans describe how process water, mine drainage, and storm 

water are handled during mining and after closure.

Mining also has the potential to affect surface-water and 

groundwater quality during mining and after closure. A key 

aspect of evaluating potential impacts to water quality and 

other ecosystem services and human health is the detailed 

geochemical characterization of pre-mining baseline conditions in surface water, groundwater, sediments, and soils. Ore 

deposits represent anomalous concentrations of a variety of 

elements that typically go beyond the mineral commodities 

that are the focus of mining. These chemical anomalies can be 

manifested in all media (rock, water, soil, sediment, and biota) 

in the vicinity of the mine site and represent a natural part of 

these ecosystems. Because elevated concentrations are part 

of the natural ecosystem, these concentrations represent more 

realistic closure goals than generic regulatory guidelines.

Mine-waste characterization and designing a mine-waste 

management plan are important aspects of assessing the potential for water-quality impacts. There are several types of mine 

waste, and mine-waste characterization is commonly done in 

a staged approach. Materials that are characterized include 

waste rock (material removed to access the ore), spent ore, 

and the rock that forms the walls of the pit, if open pit mining 

is used to produce the ore. Spent ore can be either tailings or 

leach-pad waste (residue from moderately crushed rock that 

has had metals leached from it by reactive solutions). Tailings 

generally result from the concentration of ores from basemetal (such as copper, lead, zinc, and nickel) and other metal 

(such as iron, and some uranium, gold, and silver) mines. 

Ore is crushed to sand or silt size and transferred to vats that 

contain surfactant solutions. The vats are aerated, which 

produces bubbles in the solution. Specific minerals adhere 

to the bubbles, which rise to the top and are skimmed off. 

Depending on the specific ore and the chemical additives used, 

either ore minerals or gangue (waste) minerals will rise to the 

top. The waste, or tailings, from this process is disposed of in a 

variety of ways. Tailings can be pumped as a slurry to tailings 

ponds, dried and stacked in piles, or mixed with a binder and 

disposed of as a paste, which then solidifies. Leach-pad wastes 

are commonly associated with many copper, gold, silver, and 

uranium deposits. Ore is crushed and placed on impermeable 

liners, where it is leached with solutions designed to extract 

the mineral commodities of interest. Copper is commonly 

leached with a sulfuric acid solution, whereas gold and silver 

are leached with a cyanide solution. The leach pad residues are 

usually left in place for disposal, and a new pad is constructed 

for new ore. Closure of spent piles may involve regrading the 

pile and (or) capping it.

The potential mine waste is characterized using several 

types of tests. It is common for tests to be conducted on cores 

from exploratory drilling and on reject material from preliminary metallurgical testing. Acid-generating potential is a major 

concern for some types of ores and associated wastes because 

of their specific mineralogical and geochemical characteristics. 

One type of test is known as “acid-base accounting” and is 

used to estimate the ability of the material either to generate 

acid or to neutralize it. The mineral pyrite—fool’s gold—is 

the main source of acid mine drainage, so the test estimates its 

abundance on the basis of the amount of sulfur in a sample. 

These tests estimate only the maximum acid-generating 

potential of a sample because a number of factors can limit 

the rate of acid-mine-drainage production.
The Quest for Best Practices B7

Acid-base accounting is categorized as a “static” test, 

because it is based on a single set of tests. In contrast, leaching 

tests of potential mine-waste material that span multiple weeks 

or years are known as “kinetic” tests. Kinetic tests not only 

provide information about the acid-generating potential of 

mine waste but also provide insights about the leachability 

of metals and other trace elements from the mine waste, as 

well as a timeframe for the reactions to occur. Humidity-cell 

tests are an example of a kinetic test. Humidity-cell tests, in 

which a sample is leached successively on a weekly basis 

in laboratory experiments that span weeks or years, provide 

better information about the ability of waste materials to 

produce acid and leach environmentally important trace 

elements. Laboratory kinetic tests can be improved upon by 

constructing similar experiments at a larger scale in the field. 

Here, the waste material is exposed to the natural variations 

of temperature and precipitation that the actual mine waste 

will be subject to rather than the constant, somewhat artificial 

conditions imposed by the bench-top experimental protocols. 

In any case, insights gained from pre-mining mine-waste 

characterization are typically refined as experience is gained 

during mine operation.

A mining impact assessment also requires that a number 

of design features be considered when evaluating potential 

impacts to water quality. Hydrologic characterization of 

mines is essential for understanding the risks associated 

with pit lakes or underground mine waters, and hydrologic 

characterization of the waste facilities provides insights about 

the exposure of surface water and groundwater to solid waste. 

Facilities have to be designed not only to accommodate typical 

seasonal variations in the amount of precipitation, but they 

also must be able to handle less common 500-year floods. The 

water-management, solid-waste-management, and wastewatermanagement plans are also critical. Permit applications are 

required to describe closure and reclamation plans, how these 

plans affect water quantity and quality, and what the long-term 

needs for active or passive treatment of the water will be. 

These details include the final state of waste-rock piles and 

tailing storage facilities, such as the nature of post-closure 

covers, and plans to manage drainage flows off or out of waste 

piles. Likewise, predictions must be made regarding long-term 

water quality in planned pit lakes.

Mining also has the potential to affect aquatic ecosystem 

resources adversely in a variety of ways that must be evaluated and described during the permitting process. Baseline 

characterization is needed to describe pre-mining fish and 

aquatic macro-invertebrate populations. Mine construction can 

affect riparian areas and influence infiltration. Water needed 

for ore processing and beneficiation can influence stream flow. 

Water-management structures, such as ditches and trenches, 

can channel the flow and influence the natural balance of 

erosion and sedimentation in a watershed, which can impair 

aquatic habitat. Dams and other barriers can impede the fish 

movement. Accidental releases of process water or mine 

drainage can degrade quality and affect aquatic biota. Plans 

for water treatment during operation and after closure must 

be described. All these factors are considered as part of the 

permitting process.

Wetlands represent a special case for potential impacts 

associated with mining because they are explicitly regulated 

under the Clean Water Act. They perform a variety of 

important environmental management and filtration functions, 

including groundwater recharge and discharge, flood storage 

and moderation, sediment trapping, and nutrient trapping 

and removal, and they provide habitat. The mine permitting 

process requires that wetland resources be fully documented.

The Quest for Best Practices

As previously noted, 1970 marks both the founding of 

the EPA and the start of the NEPA process that provided an 

environmental planning framework for modern mining sites. 

Passage of this law sparked a review of and an evolution in 

mining practices meant to protect the environment, and that 

review continues today. One measure of the success of the 

law and how it has improved environmental practices associated with metal mining in the United States is the decreased 

number of mining sites being designated as Superfund sites.

The EPA uses the Hazard Ranking System of CERCLA 

to evaluate the level of contamination at a site (U.S. Environmental Protection Agency, 2016). The Hazard Ranking 

System is a numerical screening system that assesses potential 

impacts to both human and environmental (ecological) 

health. It considers groundwater, surface water, soil, and air. 

The assessments are made to determine whether the level of 

contamination at a site is severe enough to warrant inclusion 

on the EPA’s National Priorities List (NPL) (U.S. Environmental Protection Agency, 2015). Placement on the NPL 

designates a site as eligible for cleanup under the Federal 

Superfund program.

The NPL includes both mine sites and related facilities, 

such as smelters and mineral storage facilities. In the 

following discussion, only mine sites are addressed, however, 

because smelters and storage facilities lack the intimate link to 

local geologic and environmental conditions that mines do and 

can be located at some distance from a mine site. Furthermore, 

understanding the environmental impacts at modern mines 

that have also had historical production can be complicated by 

legacy issues that are unrelated to modern mining practices. 

Nonetheless, the history of site placement on the NPL can be 

considered an indication of the progress that has been made in 

environmental protection associated with mine development 

since the passage of the NEPA.

The first listing of mine sites on the NPL (that is, as 

Superfund sites) was published in 1983; all the mines listed 

in that year were old mines that had begun operating in 1900 

or earlier, and some began operating as early as 1800—more 

than 170 years before any sort of guidance on environmental 

protection was available. As of 2016, the NPL includes 

approximately 60 mine sites with diverse mining histories. 

Table B2 is a list of these mine sites, including the dates for 
B8 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

Table B2. Historical summary of mine sites placed on the U.S. Environmental Protection Agency’s National Priorities List.—Continued

[EPA, U.S. Environmental Protection Agency; ID, identification; NPL, National Priorities List; n.a., not applicable]

Site name EPA 

region State EPA site ID number Year mining started

Year modern 

mining started

Year mining 

ceased

Date listed 

on NPL

Callahan Mining Corp. 1 Maine MED980524128 1880 1968 1972 9/5/2002

Elizabeth Mine 1 Vermont VTD988366621 1809 1942 1958 6/14/2001

Ely Copper Mine 1 Vermont VTD988366571 1821 n.a. 1920 9/13/2001

Pike Hill Copper Mine 1 Vermont VTD988366720 1847 n.a. 1918 7/21/2004

Ore Knob Mine 4 North Carolina NCN000409895 1850 n.a. 1962 9/23/2009

Barite Hill/Nevada 

Goldfields

4 South Carolina SCN000407714 1989 1989 1994 4/9/2009

Brewer Gold Mine 4 South Carolina SCD987577913 1828 1987 1995 4/27/2005

Bautsch-Gray Mine 5 Illinois ILN000510407 1850 n.a. 1970 9/18/2012

Torch Lake 5 Michigan MID980901946 1890 n.a. 1969 6/10/1986

Chevron Questa Mine 

(Molycorp)

6 New Mexico NMD002899094 1920 1965 2013 9/16/2011

Jackpile-Paguate 

Uranium Mine

6 New Mexico NMN000607033 1953 1953 1982 3/15/2012

Tar Creek (Ottawa 

County)

6 Oklahoma OKD980629844 1900 n.a. 1967 9/8/1983

Cherokee County 7 Kansas KSD980741862 1850 n.a. 1967 9/8/1983

Annapolis Lead Mine 7 Missouri MO0000958611 1919 n.a. 1940 7/22/2004

Big River Mine Tailings/

St. Joe Minerals Corp.

7 Missouri MOD981126899 1929 n.a. Historical 10/14/1992

Madison County Mines 7 Missouri MOD098633415 1840 n.a. Historical 9/29/2003

Newton County Mine 

Tailings

7 Missouri MOD981507585 1850 n.a. Historical 9/29/2003

Oronogo-Duenweg 

Mining Belt

7 Missouri MOD980686281 1850 n.a. 1967 8/30/1990

Southwest Jefferson 

County Mining

7 Missouri MON000705443 1800 n.a. Historical 9/23/2009

Washington County Lead 

District–Furnace Creek

7 Missouri MON000705842 1720 n.a. 1980 3/10/2011

Washington County Lead 

District–Old Mines

7 Missouri MON000705027 1720 n.a. 1980 3/19/2008

Washington County Lead 

District–Potosi

7 Missouri MON000705023 1720 n.a. 1980 3/19/2008

Washington County Lead 

District–Richwoods

7 Missouri MON000705032 1720 n.a. 1980 3/19/2008

California Gulch 8 Colorado COD980717938 1859 n.a. Historical 9/8/1983

Captain Jack Mill 8 Colorado COD981551427 1860 n.a. 1992 9/29/2003

Central City, Clear Creek 8 Colorado COD980717557 1859 n.a. 1950 9/8/1983

Eagle Mine 8 Colorado COD081961518 1880 n.a. 1984 9/17/2001

Nelson Tunnel/Commodore Waste Rock8 Colorado CON000802630 1890 n.a. 1985 9/3/2008

Smuggler Mountain 8 Colorado COD980806277 1879 n.a. 1920 6/10/1986
The Quest for Best Practices B9

Table B2. Historical summary of mine sites placed on the U.S. Environmental Protection Agency’s National Priorities List.—Continued

[EPA, U.S. Environmental Protection Agency; ID, identification; NPL, National Priorities List; n.a., not applicable]

Site name EPA 

region State EPA site ID number Year mining started

Year modern 

mining started

Year mining 

ceased

Date listed 

on NPL

Standard Mine

Summitville Mine

Barker Hughesville 

Mining District

Basin Mining Area

Carpenter Snow Creek 

Mining District

Flat Creek IMM

Libby Asbestos Site

Silver Bow Creek/ 

Butte Area

Upper Tenmile Creek 

Mining Area

Gilt Edge Mine

Richardson Flat Tailings

Iron King Mine–

Humboldt Smelter

Atlas Asbestos Mine

Blue Ledge Mine

Coalinga Asbestos Mine

Iron Mountain Mine

Klau/Buena Vista Mine

Lava Cap Mine

Leviathan Mine

New Idria Mercury Mine

Sulphur Bank 

Mercury Mine

Salt Chuck Mine

Blackbird Mine

Bunker Hill Mining and 

Metallurgical Complex

Stibnite/Yellow Pine 

Mining Area

Black Butte Mine

Formosa Mine

Fremont National Forest/

White King and Lucky 

Lass Uranium Mines

Midnite Mine

Silver Mountain Mine

8

8

8

8

8

8

8

8

8

8

8

9

9

9

9

9

9

9

9

9

9

10

10

10

10

10

10

10

10

10

Colorado

Colorado

Montana

Montana

Montana

Montana

Montana

Montana

Montana

South Dakota

Utah

Arizona

California

California

California

California

California

California

California

California

California

Alaska

Idaho

Idaho

Idaho

Oregon

Oregon

Oregon

Washington

Washington

CO0002378230

COD983778432

MT6122307485

MTD982572562

MT0001096353

MT0012694970

MT0009083840

MTD980502777

MTSFN7578012

SDD987673985

UTD980952840

AZ0000309013

CAD980496863

CAN000906063

CAD980817217

CAD980498612

CA1141190578

CAD983618893

CAD980673685

CA0001900463

CAD980893275

AK0001897602

IDD980725832

IDD048340921

IDD980665459

OR0000515759 

ORN001002616

OR7122307658

WAD980978753

WAD980722789

1874

1870

1879

1890

1880

1909

1920

1882

1870

1876

1870

1904

1963

1904

1962

1860

1868

1861

1863

1854

1865

1905

1902

1883

1930

1890

1910

1958

1955

1928

n.a.

1984

n.a.

n.a.

n.a.

n.a.

1963

n.a.

n.a.

1986

1953

n.a.

1963

n.a.

1962

n.a.

n.a.

n.a.

1951

n.a.

n.a.

n.a.

n.a.

n.a.

1979

n.a.

1990

1958

1955

n.a.

1974

1992

1975

1960

1940

1953

1990

Active

1930

1992

1981

1969

1979

1930

1974

1963

1970

1943

1962

1975

1957

Historical

1982

Active

1991

Historical

1993

1965

1981

1965

9/14/2005

5/31/1994

9/13/2001

10/22/1999

9/13/2001

9/23/2009

10/1/2002

9/8/1983

10/22/1999

12/1/2000

2/7/1992

9/3/2008

9/21/1984

9/16/2011

9/21/1984

9/8/1983

4/19/2006

1/9/1999

5/11/2000

9/16/2011

8/30/1990

3/4/2010

5/10/1993

9/8/1983

9/13/2011

3/4/2010

9/19/2007

4/25/1995

5/11/2000

6/10/1986
B10 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

the historical start of mining, the start of modern-style mining, 

the end of mining (that is, when mining stopped), and the mine 

site’s year of placement on the NPL. Figure B2 shows this 

information in graphical form.

A few of the sites, such as the Coeur d’Alene mining 

district in Idaho (Bunker Hill) and the Butte mining district in 

Montana, have been mined more or less continuously since the 

late 1800s. Other historical mines saw early production and 

then became inactive because of ore depletion or significant 

decreases in metal prices, only to restart in modern times 

once metal prices increased or technological advances made 

production economically viable once again. Other sites began 

modern-style mining but before environmental protection 

requirements were in place (that is, before 1970). The term 

“modern-style mining” generally refers to mines that began 

operation after 1940, and is meant to include such practices as 

froth flotation for separation of ore concentrates as opposed 

to the hand-sorting of ores employed historically, and cyanide 

leaching of gold as opposed to mercury amalgamation. A final 

group of mines on the NPL is the set that went through the 

NEPA process for permitting.

The first mine that was permitted through the NEPA 

process (Summitville, Colorado) began operation under a 

new permit in 1984 and was designated as a Superfund site 

(placed on the NPL) in 1994 (table B2; fig. B2). Another mine 

permitted in 1979 (Stibnite/Yellow Pine, Idaho) was not listed 

on the NPL until 2011. For the mines listed on the NPL, the 

shortest time between the start of production and listing on 

the NPL is 3 years (Formosa, Oregon), and the longest time is 

32 years (Stibnite/Yellow Pine, Idaho). The most recent mine 

to be permitted through the NEPA process and placed on the 

NPL began operation in 1990. No new mine permitted in the 

United States since 1990 has had significant enough environmental issues to warrant placement on the NPL, even though 

numerous mines have gone into production since 1990 (Long 

and others, 2010, p. 22).

For more than two decades, the quest for best environmental practices associated with mining has seen an increased 

level of coordination and cooperation among public and 

private stakeholders. It has motivated mining companies, 

government regulatory agencies, government land management agencies, government science agencies, and academia 

to work together to identify, investigate, and promote environmental best practices in mining. The level of cooperation 

among mining companies in this effort is especially notable 

because the companies seek competitive advantages over one 

another in most other aspects of their businesses. Collaborations on environmental best practices are generally made 

through consortia that involve various groups of stakeholders, 

professional conferences that facilitate the exchange of 

1800

1820

1840

1860

1880

1900

1920

1940

1960

1980

2000

2020

Start and end year

Year listed in NPL

Historical start

Modern start

End of mining

1982

1985

1990

1995

2000

2005

2010

2015

NEPA

Environmental Overview figure 2

EXPLANATION

Figure B2. Graph showing dates associated with all the mine sites on the U.S. Environmental 

Protection Agency’s National Priorities List (NPL). Dates, which correspond to those listed in table B2, 

include the historical start of mining, the start of modern-style mining, the end of mining, and the date 

of placement on the NPL. Vertical lines connect the dates for individual sites. The figure includes sites 

that started operation before the enactment of the National Environmental Policy Act (NEPA) of 1969 

(42 U.S.C. §4321 et seq.) as well as the one that started operating after its enactment.
Mine Permitting and Planning Process Overview B11

information, technology transfer efforts through workshops 

and other training courses, knowledge-gap-driven research, 

and the publication of print and Web-based resources.

Membership in these consortia spans the range of from 

highly selective membership to open membership (table B3). 

At the highly selective end are such groups as the International Network for Acid Prevention (INAP), which is 

made up of representatives from major mining companies. 

More inclusive organizations, such as the Canadian Mine 

Environment Neutral Drainage (MEND) program or the 

American Acid Drainage Technology Initiative (ADTI), 

include the mining industry, their consultants, government 

agencies at all levels, nongovernment organizations, and 

academia. Because of their overlapping goals, these consortia 

interact extensively, mostly through jointly sponsored 

professional conferences.

Numerous “best practices” publications and case studies 

have resulted from the efforts of these consortia. For example, 

the INAP produced its online Global Acid Rock Drainage 

Guide (GARD Guide), which describes best practices for the 

prediction, prevention, and management of drainage produced 

from sulfide mineral oxidation, including metal leaching and 

acid generation (International Network for Acid Prevention, 

2014). Canada’s MEND program releases best practices and 

case study reports on its Web site (http://mend-nedem.org). 

The American ADTI publishes a series of workbooks on 

management technologies for waters affected by metal mining. 

To date, topics include the basics of metal-mining-influenced 

water (McLemore, 2008); the mitigation of metal-mininginfluenced water (Gusek and Figueroa, 2009); and the 

characteristics, predictive modeling, and sustainability of mine 

pit lakes (Castendyk and Eary, 2009). Other workbooks are 

in preparation.

Mine Permitting and Planning 

Process Overview

Mine permitting is a laborious process that must consider 

not only planning for mine development, but also planning 

for environmental protection during mining and after mine 

closure. The permitting process varies, depending on whether 

the mine will be located on private land, State land, or Federal 

land, although many aspects of the permitting process are 

similar among the location types.

In the United States, numerous permits are required 

to identify, develop, operate, and close a mine (Arizona 

Department of Mines and Mineral Resources, 2015). For 

State and Federal lands, permits may also be needed for 

Table B3. Selected consortia devoted to identifying and implementing environmental best practices associated with mining.

Consortium Host country Founding

year Membership Focus

Acid Drainage Technology 

Initiative (ADTI)

http://www.osmre.gov/

programs/tdt/adti.shtm

United States 1995 (coal);

1998 (metals)

Mining industry, government, 

academia, nongovernmental 

organizations

Promote advances in acidic drainage 

prediction, prevention, control, 

sampling and monitoring, and 

treatment. Includes both coal and 

metal mining sectors.

International Network for 

Acid Prevention (INAP)

http://inap.com.au/

International 1998 Mining companies Reduce liability associated with 

acid mine drainage related to 

sulfide mine materials.

International Council on Mining

and Metals (ICMM)

http://www.icmm.com/

International 2001 Mining companies, industry 

associations

Sustainable development performance in the mining and 

metals industries.

Mine Environment Neutral 

Drainage (MEND)

http://mend-nedem.org

Canada 1989 Mining industry, Federal and 

Provincial governments, 

academia, nongovernment 

organizations

Develop technologies to prevent 

and control acidic drainage.

Partnership for Acid Drainage

Remediation in Europe

(PADRE)

http://www.padre.imwa.info

Europe 2003 Mining industry, government, 

academia, nongovernmental 

organizations

Promote dissemination of knowledge 

of current best practices and 

innovations relating to acidic 

drainage prediction, prevention, 

and remediation with a 

European emphasis.
B12 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

mineral exploration. Permits for groundwater use, surfacewater use, and wastewater treatment and discharge are also 

required. These permits are designed to protect water quality 

and quantity in groundwater aquifers, streams, lakes, and 

wetlands. They cover all aspects of water management at a 

mine site. The storage and use of chemicals, explosives, fuel, 

and oil at mine sites require a number of permits. Lease agreements must be established that include a geologic evaluation; 

an economic feasibility study; an environmental assessment; 

and mine operation, reclamation, and closure plans. The 

agreements must also include archeological and biological 

surveys. A number of permits govern water and solid-waste 

management at mine sites. A variety of air-quality permits may 

be required to cover emissions and dust. Permits also cover 

flood control, construction, and worker safety, among other 

aspects of mine operation.

Bonding, also known as financial assurance, is an 

important part of the mine permitting process. Companies 

are required to furnish a bond to cover the estimated costs of 

mine reclamation and closure. Financial assurance is a topic 

of significant concern during mine permitting because of the 

difficulty in making accurate estimates of reclamation and 

closure costs. Mining companies prefer to have bonds set as 

low as possible, whereas regulators seek to ensure that bonds 

are adequate for successful reclamation. A notable example 

of an inadequate bond is the Summitville Mine, which is 

a heap-leach gold-mining operation in Colorado. Galactic 

Resources Ltd. declared bankruptcy in 1992 and abandoned 

the site. The bond was $4.7 million, and although the EPA and 

the State of Colorado obtained an additional $28 million from 

the bankruptcy proceedings, the total cost of the cleanup has 

exceeded $150 million, which has been funded mostly through 

Federal and State taxes (Warhurst and Mitchell, 2000).

The time required to permit, construct, and commission a 

mine in the United States can be significant. Long and others 

(2010, p. 22) summarized the permitting and production history 

for recent mines (including uranium) in the United States. The 

time from the start of permitting to initial production ranged 

from 2 to 22 years and averaged approximately 10 years. The 

comprehensive regulatory environment in the United States 

requires environmental studies, due diligence for financing, 

permit application and approval, and public participation at 

multiple steps in the process, each contributing to the time 

from the start of the process to initial production. This timeframe for permitting encompasses the time needed to discover 

and develop an orebody and to conduct the engineering studies 

necessary to design and build a mine.

Geoenvironmental Models

The environmental footprint of a mine is complex 

and depends upon a number of factors. Mine design, 

environmental protection planning, and closure must all 

be customized to the unique geologic and geographical 

characteristics of the deposit to be mined. In many cases, 

the quest for best environmental practices can obscure the 

important influences that the geologic characteristics of an 

ore deposit have on its environmental behavior. A means of 

linking the geologic characteristics of a mineral deposit to 

its potential environmental effects, either in its undeveloped 

state or as a mine, is through the “geoenvironmental model” 

concept developed by the U.S. Geological Survey (USGS). 

A geoenvironmental model of a mineral deposit is defined 

as “a compilation of geologic, geochemical, geophysical, 

hydrologic, and engineering information pertaining to the 

environmental behavior of geologically similar mineral 

deposits (a) prior to mining, and (b) resulting from mining, 

mineral processing, and smelting” (Plumlee and Nash, 1995, 

p. 5; Seal and Hammarstrom, 2003). Geoenvironmental

models are currently largely descriptive in nature. They

represent empirical compilations of data that provide a

powerful predictive capability of possible ranges of environmental signatures. The models are best used as guidelines for

potential ranges of environmental challenges that may apply

to the site. They provide guidance for the early stages of

mine development (prefeasibility stage) to be supplanted by

site-specific data collected after a resource has been identified.

The geologic characteristics of a mineral deposit type 

influence approaches for mineral production, which, in turn, 

influence the potential environmental impacts. Geologic factors 

include the geographic distribution of mineral deposits; the 

depth of the mineral deposit; the mineralogy of the deposit; the 

grain size of the ore minerals; the ore grade; and the tonnage of 

ore (DeYoung and Singer, 1981). Climate is also an important 

factor in determining the environmental characteristics of a 

mineral deposit. Mineral deposits that are developed in wet 

climates have a greater probability of affecting surface-water 

resources, whereas impacts to groundwater resources are a 

greater concern in arid climates. The depth of the deposit 

dictates whether it will be developed by underground or open 

pit mining methods. Open pit mines require greater volumes of 

waste rock to be removed to access the ore. The mineralogy of 

the ore determines a number of the environmental properties 

of the ore and waste, such as their acid-generating potential. 

Pyritic waste rock requires special management to mitigate acid 

mine drainage. The grain size of the ore influences the distinction between ore and waste because fine grinding may be cost 

prohibitive. More grinding also increases the energy requirements for a mine, and therefore, its carbon footprint. The ore 

grade determines how much of the processed ore becomes 

waste. Further, the environmental footprint of a mineral deposit 

varies with its stage of mineral production; at the exploration 

stage, potential environmental impacts are the lowest, but the 

potential impact increases through the mine-development stage 

to the mineral-production (mining) stage (Eggert, 1994).

The USGS has been developing geoenvironmental 

mineral deposit models for more than two decades. Preliminary 
Summary B13

models were completed in 1995 (du Bray, 1995). Since then, 

individual deposit models have been refined, such as those 

for base- and precious-metal massive sulfide deposits and 

gold deposits (Seal and Hammarstrom, 2003). Furthermore, 

environmental features of mineral deposits have been included 

in updated mineral deposit models for a variety of mineral 

deposits, including Mississippi Valley-type lead-zinc deposits 

(Leach and others, 2010); porphyry copper deposits (John and 

others, 2010); volcanogenic massive sulfide deposits (Shanks 

and Thurston, 2012); porphyry molybdenum deposits (Taylor 

and others, 2012); stratiform chromite deposits (Schulte and 

others, 2012); beryllium deposits (Foley and others, 2012); 

sedimentary-hosted cobalt-copper-gold deposits (Slack, 2013); 

nickel-cobalt laterite deposits (Marsh and others, 2013); 

rare-earth-element deposits (Verplanck and others, 2014); and 

magmatic nickel-copper-(platinum-group-element) deposits 

(Schulz and others, 2014).

Future Trends

Future trends in mining will likely have major implications for environmental protection—some positive and some 

negative. Some of these trends are the result of the increasing 

challenge inherent in making new discoveries coupled with 

increased demand for mineral commodities and the associated 

increase in the price of those commodities. Other trends are 

the result of a desire for more sustainable mining practices. 

These trends include the move toward the mining of lowerore-grade, large-tonnage deposits; increasing awareness of 

the carbon footprint of mining operations; and addressing 

the water requirements for larger scale mining operations in 

a context of predicted global increases in drought conditions 

resulting from climate change.

The trend to mine lower grade, larger tonnage deposits 

is motivated by several factors (Mudd, 2007a, b; 2010). One 

is the depletion of high-grade deposits near Earth’s surface, 

which has driven explorers to seek deposits at greater depth 

and (or) of lower grade. The lower grades necessitate finding 

deposits of larger size to achieve greater economies of scale. 

The economic feasibility of mining lower grade deposits has 

also been made possible by technological breakthroughs, 

such as solvent extraction and electrowinning of low-grade 

copper ore (a heap-leaching technique), in situ leaching for 

uranium, carbon-in-pulp extraction of gold, and other hydrometallurgical techniques. From an environmental perspective, 

these techniques will likely lead to larger waste rock piles 

and tailings storage facilities that will require mitigation and 

long-term management. Other trends that have important 

environmental ramifications include the use of more cyanide 

at gold mines, increased energy use, and increased water use.

Increased energy use is correlated with increased 

greenhouse-gas emissions (Mudd, 2007b, 2010). This correlation has raised concerns about greenhouse-gas emissions 

associated with mining and the carbon footprint of mining in 

general. The desire to lower the carbon footprint of mining has 

motivated recent research in the feasibility of storing carbon 

dioxide in mine waste through mineral-carbonation reactions. 

Mine waste that contains magnesium-silicate minerals appears 

to be the most amenable for carbon sequestration through 

the engineered formation of magnesium-carbonate minerals. 

Candidate mine wastes include those from platinum-groupmetal deposits, magmatic nickel-copper deposits, asbestos 

deposits, and diamond deposits (Rollo and Jamieson, 2006; 

Pronost and others, 2011; Vogeli and others, 2011; Assima and 

others, 2014). The volume of this material may contribute to 

the development of “carbon-neutral” mines in the future.

Water management is an important issue at mines, 

both because of the quantity of water used at mines and the 

potential effect of mining on water quality. Resolving this 

issue is likely to become more pressing in the future if the 

frequency of droughts increases owing to climate change, as 

is widely predicted. A number of factors influence water use 

at mines, including climate, the ore-processing method used, 

mine design (for example open pit, underground, and in situ 

leaching), mine size, and material transport (for example, 

by slurry pipeline or by truck). There is a strong correlation 

between lower grade ores and increased water use for base 

(copper, lead, nickel, and zinc) and precious-metal (gold, 

silver, and platinum-group metal) mines, uranium mines, 

and diamond mines (Mudd, 2008). Thus, the trend toward 

mining lower grade ores is likely to result in increased water 

use at future mine sites and to have greater impacts on water 

availability in areas around mines.

Summary

Although the connection between society and mining 

began in the Stone Age when human ancestors first began 

fashioning tools from stone, policies and laws for managing 

the impact of mining on the environment have been in place 

for only a few decades. The steady and rapid growth of the 

global population since the start of the Industrial Revolution 

means that the amount of land available to provide for all our 

material needs is continuing to decrease on a per capita basis. 

With this increasing population comes increased demand 

for mineral resources, and meeting this demand is likely to 

become more challenging as near-surface high-grade deposits 

are depleted. A future trend with respect to mining will probably be toward the mining of larger deposits with lower grade 

ores. This trend will pose new environmental challenges, 

including how to manage greater volumes of mine waste, how 

to deal with increased greenhouse gas emissions resulting 

from the mining of larger volumes of ore, and how to address 

the increased water-resource needs of larger scale mines while 

still meeting the water requirements of areas near the mine for 

such uses as agriculture, drinking water, and sanitation.
B14 Critical Mineral Resources of the United States—Environmental Considerations Related to Mining of Nonfuel Minerals

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Antimony

By Robert R. Seal II, Klaus J. Schulz, and John H. DeYoung, Jr.

With contributions from David M. Sutphin, Lawrence J. Drew, 

James F. Carlin, Jr., and Byron R. Berger

Chapter C of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–C

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter C

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................C1

Introduction..................................................................................................................................................C1

Uses and Applications ......................................................................................................................C1

Demand, Availability of Supply, and Consumption .......................................................................C3

Strategic and Critical Resource Issues .........................................................................................C4

Geology.........................................................................................................................................................C6

Geochemistry......................................................................................................................................C6

Mineralogy..........................................................................................................................................C6

Deposit Types .....................................................................................................................................C6

Resources and Production ........................................................................................................................C9

Identified Resources .........................................................................................................................C9

Additional Resources ........................................................................................................................C9

Exploration for New Deposits .................................................................................................................C10

Environmental Considerations ................................................................................................................C11

Sources and Fate in the Environment ..........................................................................................C11

Mine Waste Characteristics ..........................................................................................................C11

Human Health Concerns .................................................................................................................C12

Ecological Health Concerns ...........................................................................................................C12

Mine Closure.....................................................................................................................................C13

Problems and Future Research ..............................................................................................................C13

References Cited.......................................................................................................................................C13

Suggested citation:

Seal, R.R., II, Schulz, K.J., and DeYoung, J.H., Jr., with contributions from David M. Sutphin, Lawrence J. Drew, 

James F. Carlin, Jr., and Byron R. Berger, 2017, Antimony, chap. C of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and 

Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and prospects 

for future supply: U.S. Geological Survey Professional Paper 1802, p. C1– C17, https://doi.org/10.3133/pp1802C.
Antimony

By Robert R. Seal II, Klaus J. Schulz, and John H. DeYoung, Jr.

With contributions from David M. Sutphin, Lawrence J. Drew, James F. Carlin, Jr.,1

 and Byron R. Berger1

1

 Deceased.

Abstract

Antimony is an important mineral commodity used 

widely in modern industrialized societies. The element imparts 

strength, hardness, and corrosion resistance to alloys that are 

used in many areas of industry, including in lead-acid storage 

batteries. Antimony’s leading use is as a fire retardant in safety 

equipment and in household goods, such as mattresses. The 

U.S. Government has considered antimony to be a critical 

mineral mainly because of its use in military applications. The 

great majority of the world’s antimony comes from China, 

and much of the remainder is shipped to China for smelting. 

Antimony resources are unevenly distributed around the 

world. China has the bulk of the world’s identified resources; 

other countries that have identified antimony resources 

include Bolivia, Canada, Mexico, Russia, South Africa, 

Tajikistan, and Turkey. Resources in the United States are 

located mainly in Alaska, Idaho, Montana, and Nevada. The 

most significant antimony mineral deposits occur in geologic 

environments with a thick sequence of siliciclastic 

sedimentary rocks in areas with significant fault and fracture 

systems. The most common antimony ore mineral is stibnite 

(Sb2S3 ), but more than 100 other minerals also contain 

antimony. The presence of antimony in surface waters and 

groundwaters results primarily from rock weathering, soil 

runoff, and anthropogenic sources. Global emissions of 

antimony to the atmosphere average 6,100 metric tons per 

year. Empirical data suggest that the acid-generating potential 

of antimony mine waste is low.

Introduction

Antimony is a brittle, silver-white, shiny metal that has a 

specific gravity of 6.68 and a melting point of 630.5 degrees 

Celsius (°C) (Miller, 1973). It is a poor conductor of heat 

and electricity. Antimony metal is too easily broken to be 

used by itself, but it imparts strength, hardness, and corrosion 

resistance to alloys (Miller, 1973). In addition, antimony does 

not readily oxidize and keeps its luster even in moist air and 

at elevated temperatures in the range of 100 to 250 °C. At 

temperatures above its melting point, powdered antimony 

ignites and burns with a white-green flame. Antimony is 

resistant to attack by dilute hydrochloric acid and concentrated 

hydrofluoric acid. Small amounts of antimony (about 0.2 part 

per million [ppm]) are found in Earth’s crust, making antimony about as rare as some of the heavy rare-earth elements 

(Eyi, 2012). Antimony rarely occurs as the native metal 

because of its strong affinity for sulfur and such other metals 

as copper, lead, and silver; it is typically found in sulfides, 

sulfosalts, oxides, antimonates, and antimonites (Boyle and 

Jonasson, 1984).

Identified resources of antimony are unevenly distributed 

around the world. Figure C1 shows the locations of selected 

antimony mines, deposits, and major occurrences. All the 

deposits that produce antimony are hydrothermal deposits. The 

mines that produce antimony fall into two main categories—

those that produce antimony as their primary commodity 

and those that produce antimony as a byproduct of mining for 

other commodities. Antimony production is dominated by one 

country, China.

Uses and Applications

Antimony has been an important mineral through much 

of human history. For example, the ancient Egyptians and 

early Hindus used stibnite (Sb2S3), which is the major ore 

mineral for antimony, to produce black eye makeup as early as 

about 3100 B.C. Medieval alchemists thought that antimony 

could be used to convert lead into gold. Today, antimony 

is used in lead-acid storage batteries for backup power and 

transportation; in chemicals, ceramics, and glass; in flameretardant materials; and in heat stabilizers and plastics (fig. C2; 

Wang, 1919; Gibson, 1998; Carlin, 2011; Anderson, 2012).

Lead-acid storage batteries (the kind commonly used in 

automobiles) contain from 4 to 6 percent antimony alloyed 

with lead; antimony’s resistance to corrosion is why so-called 

“lead” battery terminals are actually made of a lead-antimony 

alloy (Gibson, 1998). Lead-acid storage batteries account for 

more than two-thirds of the use of metallurgical antimony, 
C2 Critical Mineral Resources of the United States—Antimony 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10°

20°30°40°50°60°

80°

70°

70°

80°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Antimony deposits, mines,  and major occurrencesMajor antimony belt and name

EXPLANATION

Antimony Figure C1.

Coeur d’Alene district(Yellow Pine [StibniteGold] U.S. Antimony) McLaughlin

Beaver Brook

Lake George

Becker-Cochrane

San Jose (Sierra de Catorce)

La Cruz

Bolivian antimony belt

Bolivian

antimony

belt

Kharma, Chilcobija

Xikuangshan

South China belt

Yinshan-Tianshan belt

Quinling-Kunlurshan belt

Sarylakh

Sentachan

Olimpiada

Kadamzhai

Kassan

Jijikrut, Chulboi Nichesu, Savoyardy Turhal

Stadtschlaining

Krupanj-Zajaca

Tupiza area

La Lucette

Brioude-Massiac

Krasna Hora

Chatkal Range

Hillgrove

Costerfield

Bawdin

Bau

Blue Spec

Thailand antimony district

Workman’s Bench

Eagle Creek

Alto do Sobrido

San Antonio

Antimony Line(ConsolidatedMurchison)

 resource, whereas World map showing locations of selected antimony deposits, mines, and major occurrences. Some symbols represent a single mine or

o shown are major antimony belts in China

Figure C1.others represent a cluster of deposits in one area. Data are from Laznicka (2010, table 8.4) and Schwarz-Schampera (2014). Als(Wu, 1993) and the Bolivian antimony belt, which has approximately 500 antimony-gold deposits (Arce-Burgoa and Goldfarb, 2009).
Introduction C3

Flame retardants

35%

Transportation

and batteries

29%

Chemicals

16%

Ceramics

and glass

12%

Others

8%

Figure C2.Antimony Figure C2 Pie chart showing major end uses of 

antimony as a percentage of antimony consumption in the 

United States in 2012. Data are from Carlin (2013b).

much of which is obtained from secondary (scrap) sources 

(Chegwidden and Bedder, 2012). Very high purity antimony 

metal (99.999+ percent pure) is used by the semiconductor 

industry in silicon wafers for making infrared detectors, 

diodes, and other devices, as well as for other applications. 

Anti-friction bearings, type metal for mechanical typesetting, solder (as much as 10 percent antimony, but typically 

much less), and decorative castings of Britannia metal 

(5 percent antimony) and pewter (7.5 percent antimony in 

modern products) also contain antimony. Graphite bearings 

are impregnated with antimony to increase heat tolerance. 

Antimony is used in nuclear reactors together with beryllium 

as startup neutron sources. “Antimony black” is finely ground 

metallic antimony used in bronzing for metals and plaster 

casts (Miller, 1973).

Antimony in the form of antimony trioxide (Sb2

O3 ) 

is used in the United States mainly as a flame retardant in 

adhesives, paints, papers, plastics, and sealants. Antimony 

trioxide is also used as a fire retardant backing on rubber 

and textile upholstery, typically with bromine- or chlorinebased halogenated compounds (European Flame Retardants 

Association, 2006). Major markets for flame retardants include 

electronics, plastics, and fabrics used in making children’s 

clothing, aircraft and automobile seat covers, and bedding.

Other antimony compounds have a variety of uses. 

Antimony sulfide will combust in the presence of oxygen. It 

is a key combustion-supporting ingredient in the manufacture 

of ammunition primers, detonators, smoke-screen generators, 

visual range-finding shells, tracer bullets, and the striking 

surface of safety matches; it also provides the “glitter” effects 

in fireworks. The rubber industry uses antimony as a vulcanizing agent (Miller, 1973; Gibson, 1998). Antimony is also 

used in ceramics and glassmaking; for example, with suitable 

stabilizers and coloring additives, antimony trioxide glass can 

be made opaque to all visible light except long-wave infrared 

rays (Miller, 1973). Sodium antimonite (NaO3Sb) is used as 

a flame retardant, as well as for removing bubbles from glass 

(United States Antimony Corp., 2016).

Antimony also has several pharmaceutical uses. Organic 

antimony compounds, such as antiprotozoal drugs, are used 

in the treatment of certain parasitic diseases. Potassium 

antimonyl tartrate (K2Sb2(C4H2O6 )2 ) was formerly used 

as a nauseant and expectorant, although this use has been 

abandoned because of the compound’s toxicity and because 

it is difficult to administer (Abdi and others, 2003).

In an analysis of strategic mineral supplies prepared 

prior to World War II, Roush (1939, p. 238) stated that 

“Antimony has more uses of a direct military character than 

other members of the strategic group and probably more 

important uses than any of the others except mercury.” In 

addition to serving as a combustion-supporting ingredient, 

antimony has important additional uses for the military, some 

of which were first exploited in the Russo-Japanese War of 

1905 and, later, by the United States during World War I 

(Miller, 1973; Anderson, 2012). Antimony is a hardening 

agent in metals used in ball bearings, bullets capable of 

penetrating armor plate, and lead shot. It helps to strengthen 

cable sheaths, chemical pumps, foils, plumbing fixtures and 

pipes, roofing sheets, and tank linings. During World War II, 

the fireproofing compound antimony trichloride (SbCl3) 

saved the lives of many American troops when it was 

applied to tents and vehicle covers. In a fire, antimony and 

chlorine recombine to form unstable compounds that remove 

oxygen from the air, smothering the flames (Gibson, 1998; 

Eyi, 2012).

Demand, Availability of Supply, and Consumption

Demand and production.—Historically, U.S. consumption of antimony has been primarily for automotive and 

military uses. In the United States, stibnite deposits and 

occurrences were first discovered in Nevada in the 1860s, 

and some antimony ore was produced by 1865. Idaho’s 

production started in the early 1890s (Miller, 1973). During 

World War I (fig. C3), the military use of antimony resulted 

in a shortage that was eased by increased domestic production (driven by higher prices) and by imports from Bolivia 

(Miller, 1973). Expanded use of storage batteries in automobiles and trucks beginning in the 1930s further increased 

antimony consumption. In the 1940s, the many military 

applications for antimony during World War II escalated the 

demand for antimony; idle domestic mines were reopened, 

foreign mines expanded production, and secondary antimony 

recovered from scrap became a major supply source (Miller, 

1973). In 1948, domestic mine production of antimony 

reached a record high of 5,890 metric tons, exceeding the 

previous high of 5,040 metric tons produced in 1943, owing 
C4 Critical Mineral Resources of the United States—Antimony

200

150

100

50

0

1900 1910 1920 1930 1940 1950 1960 1970 1980 1990 2000 2010

Year

Antimony, in thousand metric tons

World production

U.S. apparent consumption

U.S. mine production

EXPLANATION

Figure C3

World War I

Great Depression

World War II

Korean war

Vietnam war

Iraq war

Figure C3. Graph showing world production, U.S. apparent consumption, and U.S. mine 

production of antimony from 1900 to 2012. Data are from U.S. Geological Survey (2016).

mainly to production from the Yellow Pine Mine in Idaho 

(Miller, 1973; Butterman and Carlin, 2004). Post-war domestic 

antimony production had returned to a lower level when the 

Korean war caused a temporary spike in production in the 

early 1950s. From the end of the Korean war to 2009, despite 

nearly unrivaled growth in automobile and truck sales and a 

building boom in the latter half of the 20th century, domestic 

mine production of antimony decreased; only in 1970 did it 

equal or exceed 1,000 metric tons (Miller, 1973; Butterman 

and Carlin, 2004; U.S. Geological Survey, 2016).

Since about 1984, domestic antimony mine production 

has been limited to a few companies. In 2001, antimony 

mining in the United States stopped when the Sunshine silver 

mine in Idaho stopped producing antimony as a byproduct 

(Masters, 2007). Domestic antimony mine production was 

reported in only 1 year between 2001 and 2011 (U.S. Geological Survey, 2016).

From 1900 to 2008, world mine production increased 

by a factor of more than 24, from 7,710 metric tons in 1900 

to 185,000 metric tons in 2008. In 2009, world production 

decreased to 158,000 metric tons because of the economic 

recession but increased to 186,000 metric tons in 2011 

(U.S. Geological Survey, 2016).

Supply.—In 2013, China produced about 80 percent of 

the world’s supply of antimony. Its leading role in antimony 

production is greater than it is for any other minor metal 

(Masters, 2007). In 2006, as consumption of antimony 

decreased in most countries, China’s burgeoning economy 

also made it the world’s leading antimony consumer. However, 

a reduction in antimony mining and increased domestic 

consumption led China to look to foreign sources for supply 

of additional antimony ore and concentrate, including to 

Canada and Tajikistan, thus further reducing the production 

of smelted antimony metal and oxide in those countries from 

which China was importing antimony ores (Masters, 2007). 

In the United States, there is one active antimony smelter 

in Thompson Falls, Montana (United States Antimony 

Corp., 2016).

Consumption.—Apparent consumption of antimony 

in the United States is sensitive to fluctuations in the 

economy (fig. C3). In 1997, apparent consumption peaked at 

46,600 metric tons, and it has been below that level ever since. 

In 2009, apparent consumption dropped to 21,200 metric tons, 

which was the lowest amount since 1979 (U.S. Geological 

Survey, 2016). The overall decrease in consumption may 

reflect industry’s adjustment to the increase in the price of 

antimony over time (the price rose from $0.627 per pound 

in 1999 to $2.80 per pound in 2008); however, the low 

demand brought about by economic recession is evident in the 

continued drop in apparent consumption in 2009, despite the 

decrease in price to $2.36 per pound (Carlin, 2013a).

Strategic and Critical Resource Issues

In 2012, the United States obtained its supply of 

antimony from imports, particularly from China. In the period 

2008 through 2011, the United States obtained 67 percent of 

its total imported antimony metal, ores and concentrates, and 

oxide from China; Mexico provided 15 percent; Belgium, 

7 percent; Bolivia, 4 percent; and other countries, 7 percent 

(fig. C4; Carlin, 2013b).

In 2010, two actions by China caused reductions in that 

country’s antimony production. Early in the year, the Government stated that it would not approve any new antimony 

projects until June 30, 2011. At about the same time, the 

Government shut down about 100 antimony smelters in its 

main antimony-producing region—an action that was aimed 
Introduction C5

A. Antimony metal

Italy

45%

Bolivia

26%

China

23%

Other

6%

B. Antimony ore and concentrate

China

Mexico 63%

15%

Belgium

9%

Bolivia

9%

Other

4%

C. Antimony oxide

China

67%

Mexico

15%

Belgium

7%

Bolivia

4%

Other

7%

D. All types of antimony

Antimony Figure C4.

China

74%

Mexico

12%

Peru

3%

Other

11%

Figure C4. Pie charts showing percentages of contained antimony in U.S. imports for the period 2008 –11, by 

source country. Data are from Carlin (2013b). A, Antimony metal; B, Antimony ore and concentrate; C, Antimony oxide; 

and D, All types of antimony. 

at closing illegal mines and curbing pollution (Carlin, 2011). 

As recently as 2003, China had about 400 antimony producers, 

but by the end of 2011, that number had been reduced to about 

14 to 18 qualified companies (Chegwidden and Bedder, 2012). 

In 2011, most ores and concentrates from Australia, Canada, 

Russia, and Tajikistan were exported to smelters in China, and 

Chinese antimony metal production had become increasingly 

dependent on these imports (Chegwidden and Bedder, 2012; 

Connelly and Rosart, undated).

If antimony supplies become limited, then compounds 

of chromium, tin, titanium, zinc, and zirconium could be 

substituted for antimony compounds in paint, pigments, and 

enamels; calcium, copper, selenium, strontium, sulfur, and tin 

can be substituted for antimony to harden lead and to replace 

antimony in lead-acid batteries, with enhanced performance in 

some cases (Eyi, 2012); and a number of organic compounds 

could be used as fire retardants. Such substitutes could make 

changes in production techniques and factory equipment 

necessary, however, and these substitutes have their own 

critical supply issues or could be more expensive to use. 

Barring market manipulation by a few dominant producers, 

recycling, mining, and smelter production are expected to 

meet the demand for antimony and antimony compounds for 

the foreseeable future.
C6 Critical Mineral Resources of the United States—Antimony

Geology

Geochemistry

The name antimony comes from the Greek words anti

(opposed) and monos (solitude), meaning “against being 

alone,” because antimony is most commonly found in nature 

in combination with other elements (Gibson, 1998). Antimony 

is a chalcophile element, meaning that it combines readily 

with sulfur; it also commonly combines with other metals, 

such as copper, lead, and silver, in a class of minerals known 

as sulfosalts (Miller, 1973). Although its concentration in 

continental crust (0.2 ppm) is very low (Taylor and McLennan, 

1995), it occurs in mineral deposits in concentrations ranging 

from trace amounts to major amounts (>10 percent) of 

economic importance. It occurs either as distinct antimonybearing sulfide minerals (for example, stibnite), sulfosalt 

minerals (for example, tetrahedrite), rare antimonides (for 

example, aurostibite), or as a minor or trace constituent in 

sulfides (such as arsenopyrite and pyrite), arsenides, and 

sulfosalts. Antimony occurs in anomalous concentrations in 

the deposits of many metallic minerals, including those of base 

metals (such as copper, lead, and zinc) and precious metals 

(such as gold, silver, and platinum-group elements; Boyle 

and Jonasson, 1984). Because of its association with so many 

types of mineral deposits, antimony is a significant pathfinder 

element in geochemical prospecting surveys (Boyle and 

Jonasson, 1984).

Mineralogy

Antimony is found in more than 100 minerals (Boyle 

and Jonasson, 1984), of which 46 are listed in table C1 

(Eyi, 2012). The most common antimony ore mineral is 

stibnite, which may contain traces of other metals, including 

copper, iron, gold, lead, and silver (Pohl, 2011). Stibnite 

occurs in numerous mineral deposit types (Cox and Singer, 

1986). Antimony ore may also carry undesirable elements, 

such as arsenic and mercury (Pohl, 2011). Other antimonybearing minerals, such as boulangerite (a lead-rich mineral), 

bournonite, gudmundite (an iron-rich mineral), jamesonite, 

polybasite, pyrargyrite, tetrahedrite (a copper-rich mineral), 

and valentinite are or have been of minor economic importance as sources of antimony (Miller, 1973; Eyi, 2012). 

Aurostibite is common in gold deposits enriched in antimony 

(Boyle and Jonasson, 1984). Metallic accessory minerals 

commonly found with primary antimony ores are arsenopyrite, 

chalcopyrite, galena, gold, pyrite, pyrrhotite, sphalerite, and 

silver; common gangue minerals are quartz (predominantly), 

calcite, and barite (Miller, 1973). The most common supergene antimony minerals are bindheimite, kermesite, nadorite, 

senarmontite, and stibiconite (table C1).

Deposit Types

Antimony occurs in a variety of deposits of various ages, 

including epithermal veins, pegmatites, and replacement and 

hot-spring deposits (Miller, 1973). Ore-grade concentrations 

of antimony are not common, but antimony mines can be 

divided into the following two broad categories: primary 

antimony producers and byproduct antimony producers. 

This distinction also corresponds to the empirical differences 

between simple stibnite deposits and complex polymetallic 

deposits (Schwarz-Schampera, 2014). Simple quartz-stibnite 

vein and replacement deposits account for most of the current 

and recent mine production. They can form in several different 

types of hydrothermal systems, including the peripheral parts 

of orogenic gold deposits, intrusion-related gold deposits, 

porphyry copper and molybdenum deposits, polymetallic 

mesothermal vein deposits, and sediment-hosted Carlin-type 

gold deposits (Hofstra and others, 2013). They can also occur 

alone with no apparent association with other mineral deposits.

The most significant simple quartz-stibnite deposits 

include those in Bolivia, Canada, China, Russia, and South 

Africa. Important or representative deposits from these 

countries include the Kharma (Bolivia), Beaver Brook and 

Lake George (Canada), Xikuangshan (China), Sarylakh and 

Sentachan (Russia), and Consolidated Murchison (South 

Africa) deposits. The Yellow Pine deposit in Idaho and the 

U.S. Antimony Mine in Montana are the most important 

deposits that fit within this category in the United States 

(fig. C1).

The Cordillera Oriental in Bolivia contains more than 

500 known antimony-gold deposits, which are typified by 

the Kharma deposit. The Kharma deposit consists of fissure 

veins dominated by stibnite with subordinate amounts of other 

sulfides (pyrite, arsenopyrite, and sphalerite) and sulfosalts 

(plagionite, fülöppite, and veenite), and minor amounts 

of aurostibite (Dill and others, 1995). The gold content is 

generally low (about 1 ppm) and the gold is difficult to 

recover. Quartz is the main gangue mineral, accounting for 

up to 90 percent of the veins, followed by siderite and calcite. 

The vein system is found in a fault zone that cuts Paleozoic 

fine-grained clastic rocks (argillites), which were deformed 

and later intruded by granitic rocks from the Permian to 

Jurassic Periods.

The two prime examples of simple stibnite deposits in 

Canada are the Beaver Brook deposit in central Newfoundland 

(Lake and Wilton, 2006) and the Lake George deposit 

in New Brunswick (Scratch and others, 1984; Seal and 

others, 1988). These two deposits share many geologic 

similarities. Both occur in fractures of fault systems hosted 

by Ordovician to Silurian siliciclastic sedimentary rocks, 

and they are both located near Siluro-Devonian granitic 

intrusions. The Beaver Brook deposit contains resources of 

2.12 million metric tons at an average grade of 4.41 percent 
Geology C7

Table C1. Selected antimony minerals.

[Source: Anderson (2012). *, mineral is mentioned in the text of this chapter]

Mineral name Chemical formula

Andorite AgPbSb3

S6

Annivite Cu12 (Sb,Bi,As)4S13

Arite Ni(As,Sb)

Aurostibite* AuSb2

Berthierite* FeSb2S4

Berthonite Cu7Pb2Sb5S13

Bindheimite* Pb2Sb2O6 (O,OH)

Bolivianite Ag2

Sb12S19

Boulangerite* Pb5Sb4S11

Bournonite* PbCuSbS3

Breithauptite NiSb

Cervantite Sb2

O4

Cylindrite Pb3Sn4Sb2S14

Dyscrasite Ag3Sb

Falkmanite Pb3

Sb2S6

Famatinite Cu3SbS4

Franckeite Pb5Sn3Sb2S14

Freibergite (Cu,Ag)12Sb4S13

Geocronite Pb5(As,Sb)12S8

Gudmundite* FeSbS

Horsfordite Cu6

Sb

Jamesonite* Pb4

FeSb6S14

Kermesite* Sb2S2O

Livingstonite HgSb4

S7

Mineral name Chemical formula

Meneghihite Pb4

Sb2S7

Nadorite PbSbO2Cl

Native antimony Sb

Polybasite* (Ag,Cu)16Sb2

S11

Pyrargyrite* Ag3SbS3

Ramdohrite Ag2Pb3Sb3S9

Romeite (Ca,Fe,Mn,Na)2 (Sb,Ti)2O6 (O,OH,F)

Senarmontite* Sb2

O3

Stenhuggarite CaFeSbAs2O7

Stephanite Ag5SbS4

Stibiconite* Sb3O6(OH)

Stibiobismuthinite (Bi,Sb)4

S7

Stibiocolumbite SbNbO4

Stibiodomeykite Cu3(As,Sb)

Stibioluzonite Cu3

(Sb,As)S4

Stibiotantalite SbTaO4

Stibnite* Sb2S3

Sulfo-antimonite Ag2Pb7Sb8S20

Tellurobismuthite (BiSb)2Te3

Tetrahedrite* Cu12Sb4S13

Ullmannite NiSbS

Valentinite* Sb2

O3

Zinckenite PbSb2S4

antimony (Lake and Wilton, 2006; Tallman and Evans, 1994). 

The Lake George deposit consists of two orebodies totaling 

slightly less than 2 million metric tons at average grades 

ranging between 3 and 4.2 percent antimony (Seal and others, 

1988). The mineralization at both deposits is dominated 

by quartz-stibnite veins with lesser amounts of carbonate 

minerals (calcite or dolomite) and minor amounts of pyrite. 

Arsenopyrite and native antimony are important accessory 

phases at Lake George (Scratch and others, 1984; Seal and 

others, 1988). At Lake George, the veins range in thickness 

from 0.5 to 1.5 meters (m) (Scratch and others, 1984; Seal 

and others, 1988). The predominant alteration assemblages at 

Lake George are siliceous and phyllic; the siliceous alteration 

typically extends to less than 5 centimeters (cm) from the edge 

of the vein, whereas the phyllic alteration can extend to more 

than 10 m from the vein (Seal and others, 1988).
C8 Critical Mineral Resources of the United States—Antimony

The Xikuangshan deposit in Hunan Province, China, is 

China’s leading source of antimony (Panov and No, 1989; 

Wu, 1993; Fan and others, 2004) and has been worked since 

the 16th century (Eyi, 2012). The deposit consists of four 

large orebodies and several smaller ones. The deposit replaces 

Devonian carbonate rocks and consists predominantly of 

stibnite with trace amounts of pyrite, pyrrhotite, and sphalerite. 

The primary gangue minerals are quartz and calcite along with 

some secondary barite and fluorite (Wu, 1993; Fan and others, 

2004). The ore is mostly in two stratiform, 2.5- to 5-m-thick 

mineralized beds or lenses that extend for several hundred to 

thousands of meters within a 45-m-thick carbonate sequence. 

The deposit covers an area of approximately 16 square 

kilometers (km2), has an average grade of 4 percent antimony, 

and contains more than 2 million metric tons of antimony 

(Yang and others, 2006). Quartz-stibnite and calcite-stibnite 

ore assemblages account for more than 90 percent of the ore; 

barite-quartz-stibnite and fluorite-quartz-stibnite assemblages 

contain relatively minor amounts of ore (Fan and others, 

2004). The dominant alteration that accompanied mineralization was silicification; there is also some sericitic, argillic, 

and pyritic alteration (Yang and others, 2006). Igneous rocks 

are notably absent in the vicinity of the Xikuangshan deposit 

(Wu, 1993). In addition to simple quartz-stibnite deposits, 

China also has important antimony-gold-tungsten deposits 

(Wu, 1993).

The largest antimony deposits in Russia are the Sarylakh 

and the Sentachan antimony-gold deposits in the Yakutiya 

region of eastern Russia. Estimates of the resources contained 

in these deposits vary. GeoProMining Ltd. reported that 

these two deposits contain 5 percent of the world’s antimony 

reserves with an average grade of 14 percent antimony 

(Nossoff, 2012). The company also estimated that the 

Sarylakh deposit contains 10 grams per metric ton (g/t) gold 

and that the Sentachan deposit contains 32 g/t gold. Bortnikov 

and others (2010, p. 339) estimated that the Sarylakh deposit 

contains 180,000 metric tons of antimony ore grading 

20 percent antimony along with 40 metric tons of gold ore 

grading 8 g/t gold, and that the Sentachan deposit contains 

110,000 metric tons of antimony ore grading 30 percent 

antimony along with 20 metric tons of gold ore grading 

35 g/t gold. Baltukhaev and Solozhenkin (2009) reported that 

the two deposits contain 200,000 metric tons of antimony 

with grades of 12 to 18 percent antimony at Sarylakh and 

24.6 percent antimony, 38.2 g/t gold, and 13.4 g/t silver at 

Sentachan. Another source (Porter GeoConsultancy Pty Ltd., 

2010) reported ore grades at Sentachan to be 25 percent 

antimony, 37 g/t gold, and 13.4 g/t silver, and resources at 

Sarylakh to be 2.17 million metric tons at grades of 6 percent 

antimony and 6 g/t gold, for a total of 130,000 metric tons of 

contained antimony.

The Sarylakh and Sentachan deposits are hosted by 

deformed Triassic siliciclastic sedimentary rocks (sandstones, 

siltstones, and shales) that later were intruded by granitic 

rocks. The area was also subject to related volcanic activity. 

The deposits, which are separated by 400 kilometers (km), 

are located along a major regional fault zone (Bortnikov and 

others, 2010). The orebodies are controlled by faults and are 

dominated by quartz and stibnite. Important, although less 

abundant, minerals include ankerite, arsenopyrite, berthierite, 

bournonite, chalcopyrite, chalcostibite, galena, native gold, 

pyrite, and sphalerite. The antimony mineralization and 

the gold mineralization are paragenetically distinct from 

one another at these deposits (Bortnikov and others, 2010). 

At Sentachan, the mineralized zones reach a maximum 

thickness of 6 m; at Sarylakh, the thickness varies from 

0.8 to 17 m.

In South Africa, the Consolidated Murchison Mine 

is the most important antimony deposit on South Africa’s 

Antimony Line, and it is representative of mineralization 

along a significant Archean shear zone (Jaguin and others, 

2012). Mineralization is predominantly hosted by quartzmuscovite schists and occurs as quartz-carbonate veins 

bearing stibnite and berthierite with associated pyrite and 

arsenopyrite (Muff, 1978).

The Yellow Pine deposit in Valley County, Idaho, 

contains the largest antimony resource in the United States 

(Miller, 1973) and produced intermittently from 1932 through 

1992. Mining in the area began when gold was discovered 

in 1900; antimony recovery with gold was first recognized 

in 1929 (Palencia and Mishra, 1986, p. 9). The antimony 

occurs partly in small high-grade quartz-stibnite veins, rarely 

more than a foot wide, and in large low-grade zones of 

disseminated stibnite in shear zones in quartz monzonite of the 

Idaho batholith (White, 1940). Two periods of mineralization 

are recognized (White, 1940): the first consists of pyrite, 

arsenopyrite, and gold; the second consists of antimony and 

silver. The Yellow Pine deposit was being re-evaluated in 

2014 as part of the Stibnite Gold project (formerly known as 

Golden Meadows project) (Huss and others, 2014).

The U.S. Antimony Mine in western Montana is the 

second largest simple quartz-stibnite vein deposit in the 

United States. It had significant production in the past, but 

was closed in 1983; the mine’s production and reserves 

are estimated to be nearly 15,400 metric tons of antimony 

(Hofstra and others, 2013). The deposit is located on the north 

side of a major east-west-trending wrench fault system (the 

Lewis and Clark line). The quartz-stibnite veins occur near 

the contact between quartzite and phyllite in the uppermost 

Prichard Formation. The veins can be as large as 1 km long 

and 1 m wide and contain 60 percent stibnite; accessory 

sulfide minerals include arsenopyrite, pyrite, sphalerite, 

and chalcopyrite (Hofstra and others, 2013). Additional 

potentially important byproduct sources of antimony include 

the polymetallic veins at the Sunshine silver mine in the 

Coeur d’Alene mining district, Idaho.
Resources and Production C9

Resources and Production

Identified Resources

In 2013, the largest identified resources of antimony were 

located in Bolivia, China, Mexico, Russia, South Africa, and 

Tajikistan. Resources in the United States are located mainly in 

Alaska, Idaho, Montana, and Nevada. World reserves total about 

1.3 million metric tons of contained antimony (table C2). Onethird of these reserves are located in China (Guberman, 2014).

In a study of large deposits of several mineral commodities,

Laznicka (1999) noted that the world’s 24 giant and supergiant 

antimony deposits contained a total of 6.97 million metric tons 

of antimony. Laznicka estimated the total global endowment 

to be between 7.1 and 7.5 million metric tons of antimony.

There are abundant identified antimony resources 

available, but the bulk of those resources are in a few very 

large deposits that are not evenly distributed across the globe. 

Today’s market favors large deposits that are conducive to 

high-volume bulk-mining techniques. The United States 

and the countries of the Western Hemisphere appear to have 

mostly small deposits that are uneconomic to mine under 

current and foreseeable conditions.

In 2012, Midas Gold Corp. (the owner) contracted for a 

preliminary economic assessment to be done of the indicated 

and inferred mineral resources at the Golden Meadows project 

in Idaho to demonstrate the potential for positive economic 

returns and identify areas requiring additional work. A 

subsequent prefeasibility study of the project (renamed the 

Stibnite Gold project) was completed in 2014. The property 

includes three zones of gold-antimony-silver mineralization 

within an area of significant historical mining activity. Past 

production was reported from the Hanger Flats (1928–38), 

Yellow Pine (1938 –92), and West End (1978–97) deposits 

(Huss and others, 2014, p. 6 -7– 6 -9). Conventional open pit 

methods were being considered for mining the three deposits, 

all of which are located within 3 km of each other and may in 

fact represent one large mineralized zone. The 2014 study estimated indicated and inferred resources to be about 130 million 

metric tons at grades of about 1.6 g/t gold and 0.06 percent 

antimony (Huss and others, 2014, p. 1-10).

Additional Resources

Antimony resources that may be mined in the future are 

likely to be those tied directly to deposits of precious metals, 

copper, lead, and (or) zinc, similar to those from which most 

domestic antimony has historically been recovered as a 

byproduct or coproduct. Gold is an important joint product 

with antimony, but gold-antimony veins are commonly mined 

just for their gold. Because the presence of antimony makes 

gold more difficult and more expensive to process (the antimony interferes with the heap-leaching agent by consuming 

oxygen and hindering the effect of cyanide on the gold ore), 

some amount of gold ore that has a high antimony grade may 

be stockpiled (Eyi, 2012).

Table C2. Estimated world production and reserves of antimony 

in 2013, in metric tons of antimony.

[e, estimated; NA, not available; n.r., not reported]

Country Production1, e Reserves2

Australia 3,275 n.r.

Bolivia 5,081 310,000 

Burma 9,000 n.r.

Canada 76 n.r.

China 120,000 3

460,000

Kyrgyzstan 1,200 n.r.

Mexico 294 n.r.

Russia 8,700 350,000

South Africa 2,400 27,000 

Tajikistan 4,675 50,000

Turkey 4,600 n.r.

Other countries NA 4

150,000

World total 159,000 1,300,000

1

From Guberman (2015).

2

From Guberman (2014) and National Bureau of Statistics of China (2014).

3

In 2009, the Chinese Ministry of Land and Resources reported reserves 

of 2,460,000 metric tons; however, high-grade antimony reserves in 

Lengshuijiang, Hunan Province, China, are near exhaustion (Chegwidden 

and Bedder, 2012, p. 5).

4

Includes estimated reserves for countries listed as “n.r.” above.

A post-World War II evaluation of domestic mined 

resources concluded that “the United States has no deposits, 

from which the ore is mined principally for antimony, that are 

large enough or rich enough to compete with foreign sources 

in normal times” (Shaum and others, 1948, p. 53). This situation has not changed since that report was completed. If the 

United States could no longer import antimony in quantities 

to meet its needs, then additional domestic production may 

be obtainable from base- and precious-metal deposits and 

from areas of identified base-metal, gold, silver, and tungsten 

mineralization (Miller, 1973). These sources may include 

undiscovered resources in parts of the United States and other 

nations of the Western Hemisphere, as well as Australia, 

South Africa, and other trading partners of the United States. 

Table C3 lists some potential near-term sources of antimony 

ore and concentrate (Chegwidden and Bedder, 2012).

Enhanced recovery of antimony from precious-metal 

deposits may represent the most readily available source of 

antimony if demand were to increase rapidly. The Sunshine 

silver mine in Idaho was a significant producer of byproduct 

antimony from a deposit of silver-rich tetrahedrite (a common 

ore mineral in silver deposits). The mine stopped recovering 

antimony before its closure in early 2001 (Carlin, 2003). In 

addition, antimony can be recovered as a byproduct from 

some base-metal refining, for example, from Mississippi 

Valley-type lead deposits in the United States.
C10 Critical Mineral Resources of the United States—Antimony

Table C3. Potential additional sources of antimony ore and concentrate, by country.

[Sources: Chegwidden and Bedder (2012, p. 10) and Huss and others (2014)]

Country Notes

Australia The Hillgrove Mine, which closed in late 2015, might be reopened. Other deposits in New South Wales and

Western Australia are being evaluated.

Bolivia The San Antonia de Turiri Mine may be developed. 

Burma Exports of antimony concentrates increased from 1,200 metric tons in 2002 to 14,100 metric tons in 2011. 

There may be capacity for a further increase in output.

Canada At least three companies have been exploring antimony deposits in Newfoundland and Labrador, the 

Yukon Territory, and New Brunswick.

China Tibet may increase output. 

Mexico U.S. Antimony Corp. expected to increase production from Los Juarez deposit in Coahuila State. 

Russia In the Far East, increased production may come from Sentachanskoe and Ilinskoe; new deposits, such as 

Udereiskoe and Zhipkoshinskoe, were being developed in eastern Siberia.

South Africa There may be recovery of production at the Consolidated Murchison Mine. 

Tajikistan Output reached about 8,000 metric tons per year of concentrate from a plant with a capacity of 30,000 metric tons 

per year; there may be potential for a further increase in output. The mercury content of the ore is a concern.

Turkey Possible new production in projects near Gediz in Kütahya Province as well as Niğde and Izmir Provinces.

United States Midas Gold Corp.’s Stibnite Gold project in central Idaho is a large low-grade gold-antimony property that 

includes the Hanger Flat, Yellow Pine, and West End deposits.

Exploration for New Deposits

Most undiscovered primary antimony deposits will 

likely be located in geologic settings similar to those where 

antimony deposits are known to occur. From what is known 

about where these deposits form, rocks of any age that 

were deposited in the world’s sedimentary basins on active 

continental margins are permissive for the occurrence of 

antimony deposits, especially rocks in those basins with 

known antimony deposits. In the United States, these types 

of sedimentary rocks located in the Cordillera of Alaska, and 

in California, Idaho, Montana, and Nevada would be prime 

targets. Areas within large-scale shear zones and strike-slip 

fault zones, such as those in Nevada, are potentially favorable areas for antimony fissure-vein deposits. In Arkansas, 

California, Idaho, Montana, Nevada, Washington, and other 

States, rocks near to, on trend with, and similar to those with 

known antimony occurrences could be explored for additional 

undiscovered antimony deposits.

The empirical association of hydrothermal antimony 

deposits with fine-grained sedimentary rocks, especially 

carbonaceous shales or their metamorphosed equivalents, 

may be a reflection of the antimony enrichments commonly 

found in these types of rocks. The continental crust averages 

0.2 ppm antimony (Taylor and McLennan, 1995), yet average 

shales contain 1.5 ppm (Li and Schoonmaker, 2003), and 

black shales average 5.0±0.5 ppm (Ketris and Yudovich, 

2009). Although the source, transport, and concentration 

mechanisms will likely vary among deposit types, a fertile 

source region may be essential to the formation of economic 

deposits. Further, Pitcairn and others (2006) demonstrated 

a significant depletion in antimony and other elements 

(arsenic, gold, mercury, molybdenum, silver, and tungsten) 

with increased metamorphic grade in the Alpine and Otago 

schists of New Zealand. These findings suggest that highgrade metamorphism (>400 °C) may mobilize antimony 

into hydrothermal fluids that may form quartz-antimony 

veins at lower temperature under favorable conditions. The 

combination of siliciclastic sedimentary rocks and an egress 

pathway for hydrothermal fluids from high-temperature 

source regions is common to most of the significant simple 

quartz-stibnite deposits described in the Geology section 

above. An association with major structural fault systems 

that could access deeper regional metamorphic fluids is 

found at the Xikuangshan deposit (China), the Sentachan 

and Sarylakh deposits (Russia); the Antimony Line (South 

Africa), and the U.S. Antimony Mine (Montana). In contrast, 

igneous heat sources associated with more local fault and 

fracture zones are found at the deposits in Bolivia and at the 

Lake George and Beaver Brook deposits in Canada.

Historically, very little prospecting has been directed 

solely toward antimony (Miller, 1973). Most exploration 

that resulted in the incidental discovery of an antimony 

deposit has been the result of prospecting for another mineral 

commodity, such as copper, gold, lead, and silver. Exploration 

for antimony deposits usually includes techniques commonly 

employed to find other mineral deposits, such as geochemical 

soil and rock sampling, ground geophysical surveying, 

geologic mapping, and drilling. Surveys of mercury-vapor 

haloes in soil are helpful for identifying sulfide deposits 

of antimony, the base metals, mercury, gold, and tin 

(Fursov, 1990).
Environmental Considerations C11

Environmental Considerations

Sources and Fate in the Environment

Environmental aspects of antimony have been summarized by Rish (2004). The environmental aqueous chemistry of 

antimony has been reviewed by Filella and others (2002a, b). 

Antimony in waters results primarily from rock weathering, 

soil runoff, and anthropogenic sources. Global emissions 

of antimony to the atmosphere average 6,100 metric tons 

per year, which includes 3,500 metric tons per year from 

anthropogenic sources (energy production, mining, smelting, 

and waste incineration) and 2,600 metric tons per year from 

natural sources (including, in order of contribution to atmospheric antimony, dust, volcanoes, sea aerosols, forest fires, 

and biogenic sources; Filella and others, 2002a). Seawater 

contains approximately 0.2 part per billion (ppb) dissolved 

antimony, and uncontaminated surface waters (streams, 

rivers, and lakes) contain less than 1 ppb (Filella and others, 

2002a). In surface water, pentavalent antimony species (Sb5+) 

dominate over trivalent species (Sb3+) (Filella and others, 

2002b). The solubility of antimony in water at room temperature reaches 90 ppm in oxygenated waters, 13 ppm in anoxic 

waters, and 1.2 ppm in anoxic and sulfidic waters (Tourky and 

Mousa, 1948; Krupp, 1988).

Data are available for antimony concentrations in soils 

and plants. In their national geochemical survey of soil in the 

United States, Smith and others (2013) report concentrations 

of antimony that range from less than 0.05 to 630 ppm, 

with a mean value of less than 0.9 ppm. Filella and others 

(2002a) summarized published data for sediments that ranged 

from less than 0.005 ppm in uncontaminated areas to up to 

12,500 ppm in areas around a smelter. Antimony concentrations in plants do not show a strong correlation with soil 

concentrations; plant concentrations range from less than 

0.1 to 500 ppm (Boyle and Jonasson, 1984; Rish, 2004).

Antimony is mined primarily from quartz-stibnite vein 

deposits. Modern environmental baseline characterization 

studies of antimony deposits have not been reported; however, 

Ritchie and others (2013) report background (upstream) 

concentrations of antimony in surface water and sediments 

from mineralized areas of the historic Kantishna Hills 

district in Alaska that range from 2.7 to 4.2 ppb and from 

91 to 968 ppm, respectively. Austria (1971) reported antimony 

concentrations of soils up to 500 ppm in the vicinity of the 

Lake George antimony deposit in New Brunswick, Canada.

Mine Waste Characteristics

Information on antimony mine waste is extremely 

limited. Historically, small, high-grade (about 35 to 70 percent 

antimony) deposits have been mined, but modern production 

comes from larger, lower grade deposits, which have a median 

tonnage of 0.09 million metric tons, a maximum tonnage of 

3 million metric tons, a median grade of 3.6 percent antimony, 

and a maximum grade of more than 7.0 percent antimony 

(Bliss and Orris, 1986a–c). The Xikuangshan deposit in 

Hunan Province, China, which is the world’s largest antimony 

deposit, is an anomaly. The deposit, which has an average 

grade of 4 percent antimony, is estimated to contain more than 

2 million metric tons of antimony (Yang and others, 2006).

The mineralogical character of the ores greatly influences 

the geochemistry of the mine waste. The ores are typically 

predominantly stibnite and quartz, with subordinate amounts 

of calcite, pyrite, arsenopyrite, and native antimony, hosted 

mostly by siliciclastic or carbonate sedimentary rocks (Wu, 

1993). The most common secondary antimony minerals 

in mineral deposits are kermesite (Sb2

S2O), senarmontite 

(Sb2O3), and stibiconite (Sb3O6(OH)). Thus, the elements of 

greatest environmental concern are predominantly antimony 

and arsenic.

The acid-generating potential of mine waste is expressed 

in terms of the amount of calcium carbonate it would take to 

neutralize it; it is measured in kilograms of calcium carbonate 

per metric ton (kg CaCO3 /t) of mine waste (Price, 2009; 

International Network for Acid Prevention, 2011). The 

acid-generating potential resides primarily in pyrite. Mine 

waste can also have acid-neutralizing potential, which resides 

in carbonate minerals, such as calcite, and in some silicate 

minerals, such as feldspars.

Limited data are available on the acid-generating 

potential of antimony mine waste. Klimko and others (2011) 

investigated the acid-generating potential of mill tailings at 

two abandoned stibnite mines in Slovakia. One had a neutral 

pH (7 to 8), whereas the other had an acidic pH (3 to 5). 

Acid-base accounting results indicated that the neutral tailings 

had an acid-generating potential that was negligible as well as 

a modest acid-neutralizing potential (70 to 120 kg CaCO3/t 

of tailings). In contrast, the acidic tailings had a low acidgenerating potential (0.8 to 2.8 kg CaCO3/t) and a low 

acid-neutralizing potential (16 to 18 kg CaCO3/t) that was 

nonetheless higher than the acid-generating potential. The 

acid-neutralizing potential to acid-generating potential ratios 

ranged between 5.7 and 22.6, which suggests a low potential 

for acid generation. The presence of carbonate minerals, such 

as calcite, and only minor amounts of pyrite in the mine waste 

from antimony deposits in general suggest that the acidneutralizing potential is likely to exceed the acid-generating 

potential of the mine wastes (Price, 2009; International 

Network for Acid Prevention, 2011).

Mine-drainage data from abandoned antimony mines 

support this conclusion (Wilson and others, 2004; Casiot and 

others, 2007; Majzlan and others, 2007), although some sites 

can generate acid drainage. Ritchie and others (2013) and 

Eppinger and others (2002) found low pH (2.5 to 5.6) in pore 

water and seeps from tailings piles in the abandoned mines 

in the Kantishna district, Alaska, but the receiving creeks 

immediately downstream of the mines had nearneutral pH (6.1 to 8.3).
C12 Critical Mineral Resources of the United States—Antimony

Human Health Concerns

Belzile and others (2011) have summarized the general 

aspects of antimony exposure to humans. The Agency 

for Toxic Substances and Disease Registry (ATSDR) has 

developed a toxicological profile for antimony (Agency for 

Toxic Substances and Disease Registry, 1992). In general, 

trivalent antimony (Sb3+) is more toxic than pentavalent 

antimony (Sb5+). In humans, antimony can result in diseases 

of the liver, respiratory and cardiovascular systems, and skin 

(Wu and others, 2011). Compared to trivalent arsenic, trivalent 

antimony is 5 times less cytotoxic and 10 times less genotoxic 

(Rish, 2004).

The U.S. Environmental Protection Agency (EPA) (2009) 

has set a maximum contaminant limit of 6 ppb for antimony 

and 10 ppb for arsenic for drinking water. The World Health 

Organization (WHO) drinking water guideline is 20 ppb for 

antimony and 10 ppb for arsenic (World Health Organization, 

2008). In the Xikuangshan mining district, which includes 

three mines that produce from China’s largest antimony 

deposit and an associated smelter, Wu and others (2011) 

found 100 percent of drinking water samples collected in a 

roughly 300-km2

 surrounding area exceeded the maximum 

contaminant limit of 6 ppb for antimony set by the EPA. 

Concentrations ranged from 8.1 to 152 ppb antimony. They 

concluded that mine drainage and smelter wastes were the 

dominant sources. In contrast, the arsenic concentration of 

most samples was below the maximum contaminant limit. 

Wu and others (2011) also investigated the dietary intake 

of antimony by local residents and found that their daily 

intake exceeded WHO guidelines and were several orders of 

magnitude higher than those found in other parts of the world, 

none of which approached the WHO tolerable daily intake 

guideline. The greatest sources in the daily dietary uptake 

were rice (33 percent), vegetables (26 percent), drinking water 

(19 percent), and meat and poultry (19 percent) (Wu and 

others, 2011).

The concentrations of antimony and arsenic in soils in 

the Xikuangshan mining district also exceed EPA regional 

screening levels (U.S. Environmental Protection Agency, 

2015). Wang and others (2010) collected 23 soil samples 

within 15 km of the mines. Antimony concentrations ranged 

from 10 to 2,159 ppm, of which 96 percent exceeded the 

EPA residential soil regional screening level for antimony 

(39 ppm). More than 80 percent of soil samples also exceeded 

EPA residential guidelines for arsenic. Local exceedances 

of the residential guidelines were also found for cadmium, 

chromium, lead, manganese, and mercury.

Ecological Health Concerns

National ambient-water-quality criteria with respect to 

freshwater organisms are not available from the EPA for antimony; however, Suter (1996) presents both acute and chronic 

ecological screening benchmarks for antimony. The acute 

toxicity value is 985 ppb (also expressed as 985 micrograms 

per liter [μg/L]) and the chronic toxicity value is 104 ppb. 

No guidelines are available for sediment.

Information about surface water impacts from antimony 

mines is limited to an active mining district in China and 

several small abandoned mines throughout the world. Several 

sources of mine drainage data are available. In their study 

of the abandoned Goesdorf antimony mine in Luxembourg, 

Filella and others (2009) compiled mine drainage data for 

antimony mines from around the world (Australia, Italy, 

New Zealand, Scotland, and Slovakia). Majzlan and others 

(2007) presented mine-drainage data from the Pezinok Mine in 

Slovakia. Eppinger and others (2002) and Ritchie and others 

(2013) investigated surface water quality in the Kantishna 

district, Alaska. Casiot and others (2007) provide additional 

data from an antimony mine in France. Data are also available 

from New Zealand (Wilson and others, 2004). Collectively, 

the pH of surface water downstream of mines is near neutral 

(6.2 to 8.5), which is a reflection of the low acid-generating 

potential of the ore and mine waste and the significant acidneutralizing potential provided by carbonate minerals in the 

mineralized assemblages.

Ponded water on mine waste, pore water in tailings, and 

seeps from mine waste in the Kantishna district in Alaska have 

been documented as having low pH (2.5 to 5.6) (Eppinger 

and others, 2002; Ritchie and others, 2013). Information on 

the dissolved concentrations of sulfate (<0.5 to 2,115 ppm) 

and iron (<10 to 3,000 ppb) in mine waters and surrounding 

surface waters is limited (Casiot and others, 2007), but their 

concentrations are moderate, which is a reflection of the typically limited pyrite contents of these ores. Dissolved antimony 

concentrations range from less than 0.3 to 55,000 ppb, 

and dissolved arsenic concentrations range from less than 

1 to 3,509 ppb. Thus, mine drainage from antimony mines can 

locally exceed both acute and chronic ecological guidelines 

for antimony and arsenic. Information on other trace elements 

is lacking, with exception of surface waters in the vicinity 

of the Slate Creek deposit in the Kantishna district. All trace 

element concentrations were in the subparts-per-million 

range (<1.0 ppm), except for iron (<170 ppm), manganese 

(<6.5 ppm), aluminum (<3.7 ppm), and zinc (<1.6 ppm), all 

in mine waters (Eppinger and others, 2002).

The geochemistry of mine waters and surface waters in 

the Xikuangshan mining district in China was investigated by 

Liu and others (2010). The samples included stream waters 

in the district, seepage from waste piles, and mine water. 

The pH of the samples was comparable to those from the 

abandoned mines discussed above and ranged from 7.7 to 8.3. 

Dissolved sulfate concentrations in surface water samples 

were significantly higher at Xikuangshan (46 to 1,267 ppm) 

than those at an abandoned mine in France (Casiot and others, 

2007). Dissolved antimony concentrations at Xikuangshan 

(330 to 11,400 ppb) were significantly higher than most values 

from abandoned mines elsewhere in the world.
References Cited C13

Mine Closure

The nature of mine closure for antimony mines depends 

primarily on the method of mining. Most antimony deposits 

are exploited as underground mines. The most common solid 

waste from underground mines is mill tailings, which are 

disposed in some sort of tailings storage facility.

The long-term fate of tailings storage facilities depends 

upon the nature of the tailings and the method by which 

the facility is constructed. Some facilities can be regraded, 

capped, and revegetated. These facilities may have seepage 

and, depending upon the acid-generating potential of the 

material, may require some form of water treatment. Some of 

the tailings could be placed back in the mined-out workings, 

depending upon how the orebody was mined; however, 

the entire volume of tailings cannot be placed back in the 

mine workings because of the volume expansion caused by 

crushing and milling the ore.

Problems and Future Research

A major problem facing the countries that depend upon 

stable and secure supplies of a critical mineral commodity 

such as antimony for their industry and defense needs is 

the uneven distribution of resources and production around 

the world. Finding new resources in countries other than 

China could be a challenge for antimony-consuming nations. 

Knowledge of how quickly the United States and the rest 

of the world could develop additional production by adding 

capacity to existing mines, by reactivating nonoperating 

mines, or by bringing newly discovered deposits online is 

essential to understanding the effect of supply shortfalls 

in the event that China, which is the dominant producer, 

was to reduce antimony exports. Since 2011, production of 

antimony in China has declined and its antimony reserves 

may be declining as well (Guberman, 2015).

Although antimony occurs in several different 

deposit types of various ages, ore-grade concentrations 

of antimony are not common, and economically minable 

deposits of stibnite are generally small and discontinuous. 

This makes exploration for antimony deposits challenging. 

Detailed geologic studies would be required to understand 

the tectonic history of ancient continental margins and 

sedimentary basins as well as the relation of antimony 

deposits to that history in order to enhance the possibility 

of delineating areas that are prospective for undiscovered 

antimony resources.

From an environmental perspective, no clear case study 

of the behavior of antimony and related trace elements in a 

modern mine setting using current best practices exists. The 

toxicity of aqueous antimony species to aquatic organisms 

is a notable gap in knowledge. Knowledge of the toxicity of 

antimony in sediments is also limited.

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mineralogy, geology, metallurgy, uses, preparations, 

analysis, production, and valuation; with complete 

bibliographies (2d ed.): London, United Kingdom, Charles 

Griffin & Company, Ltd., 217 p.

Wang, Xiangqin, He, Mengchang, Xie, Jun, Xi, Jianghong, 

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Bulletin 922–I, p. 247–279.

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Barite (Barium)

By Craig A. Johnson, Nadine M. Piatak, and M. Michael Miller

Chapter D of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–D

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter D

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................D1

Introduction..................................................................................................................................................D1

Geology.........................................................................................................................................................D2

Geochemistry......................................................................................................................................D2

Mineralogy..........................................................................................................................................D3

Deposit Types .....................................................................................................................................D3

Bedded-Sedimentary Deposits ..............................................................................................D3

Bedded-Volcanic Deposits ......................................................................................................D5

Vein, Cavity-Fill, and Metasomatic Deposits ........................................................................D6

Residual Deposits .....................................................................................................................D6

Resources and Production ........................................................................................................................D7

Reserves..............................................................................................................................................D7

Other Identified Resources ..............................................................................................................D7

Undiscovered Resources .................................................................................................................D8

Exploration for New Deposits ...................................................................................................................D8

Environmental Considerations ..................................................................................................................D8

Barium Behavior in the Environment ..............................................................................................D8

Mine Waste Characteristics ............................................................................................................D9

Human Health Concerns .................................................................................................................D10

Ecological Health Concerns ...........................................................................................................D10

Mine Closure.....................................................................................................................................D11

Problems and Future Research ..............................................................................................................D11

References Cited.......................................................................................................................................D12

Suggested citation:

Johnson, C.A., Piatak, N.M., and Miller, M.M., 2017, Barite (Barium), chap. D of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, 

and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and 

prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. D1– D18, http://dx.doi.org/10.3133/

pp1802D.
Barite (Barium)

By Craig A. Johnson, Nadine M. Piatak, and M. Michael Miller

Abstract

Barite (barium sulfate, BaSO4 ) is vital to the oil and 

gas industry because it is a key constituent of the mud used 

to drill oil and gas wells. Elemental barium is an additive 

in optical glass, ceramic glazes, and other products. Within 

the United States, barite is produced mainly from mines in 

Nevada. Imports in 2011 (the latest year for which complete 

data were available) accounted for 78 percent of domestic 

consumption and came mostly from China.

Barite deposits can be divided into the following 

four main types: bedded-sedimentary; bedded-volcanic; 

vein, cavity-fill, and metasomatic; and residual. Beddedsedimentary deposits, which are found in sedimentary rocks 

with characteristics of high biological productivity during 

sediment accumulation, are the major sources of barite 

production and account for the majority of reserves, both 

in the United States and worldwide. In 2013, China and 

India were the leading producers of barite, and they have 

large identified resources that position them to be significant 

producers for the foreseeable future. The potential for 

undiscovered barite resources in the United States and in many 

other countries is considerable, however. The expected tight 

supply and rising costs in the coming years will likely be met 

by increased production from such countries as Kazakhstan, 

Mexico, Morocco, and Vietnam.

Barium has limited mobility in the environment and 

exposed barium in the vicinity of barite mines poses minimal 

risk to human or ecosystem health. Of greater concern is the 

potential for acidic metal-bearing drainage at sites where the 

barite ores or waste rocks contain abundant sulfide minerals. 

This risk is lessened naturally if the host rocks at the site 

are acid-neutralizing, and the risk can also be lessened by 

engineering measures.

Introduction

Barite (barium sulfate, BaSO4 ) is an industrial mineral 

commodity that is used primarily in the drilling of oil and gas 

wells. The mineral commodity is also referred to as baryte 

or barytes. It is a key constituent of drilling mud, which is the 

fluid pumped into the oil or gas well to lubricate the bit and 

drill stem, remove rock chips, prevent collapse of well walls, 

and prevent blowouts if overpressured strata are encountered. 

Barite has an unusual combination of properties—high 

density, softness, and chemical inertness—that make it 

exceptionally well suited for this purpose. Other minor uses of 

the mineral include its use as an additive for friction materials, 

paints, plastics, rubbers, and other products; feedstock for 

chemical manufacturing; and shielding in X-ray and gammaray applications. Barium, which constitutes 59 percent of 

barite by weight, is used in ceramic glazes, enamels, optical 

glass, primers, signal flares, steel hardeners, welding fluxes, 

and a variety of other products. The mineral witherite (barium 

carbonate, BaCO3) is another source of barium that accounts 

for only a small fraction of current production. The barium in 

witherite is more highly concentrated than in barite (barium 

constitutes 70 percent of witherite, by weight). Furthermore, 

barium is more easily liberated from witherite than it is from 

barite because witherite is readily dissolved by acids. On the 

other hand, witherite normally occurs in deposits that are 

smaller and more discontinuous than barite deposits, and they 

are thus more difficult to mine.

Barite is a critical mineral commodity because of its 

uses in the exploration for and development of petroleum and 

natural gas resources. Drilling activity accounts for nearly 

95 percent of domestic consumption and about 90 percent 

of global consumption. Economic deposits of barite are 

relatively common and are found in many countries. Although 

prices have risen substantially in recent years, the commodity 

remains relatively inexpensive. Thus, it is unlikely that the oil 

and gas industry will have a need for substitutes in the short 

term. All the likely substitutes (celestite, hematite, ilmenite, 

and magnetite) present significant drawbacks. The principal 

strategic concern with respect to barite is that adequate 

supplies be available at low cost in the geographic regions that 

are currently being explored for oil and gas resources and in 

geographic regions that will be explored in the future.

Until 1950, output of barite from U.S. mines equaled 

or exceeded domestic consumption. In succeeding years 

(fig. D1), domestic consumption outpaced production from 

U.S. mines by a steadily increasing margin, so that by 1980, 
D2 Critical Mineral Resources of the United States—Barite (Barium)

0

2

4

6

8

10

1950 1960 1970 1980 1990 2000 2010 2020

Barite production and consumption, in million metric tons

Year

Barite, Fig D1.

U.S. production

World

production

U.S. consumption

Oil embargo

of 1973

Economic downturns

of 1982 and 2008

Figure D1. Graph showing barite world production, 

U.S. production, and U.S. consumption from 1950 to 2011. 

Data are from Kelly and Matos (2013).

imports accounted for some 46 percent of U.S. consumption. 

Barite production declined precipitously during the economic 

downturn of 1982. Since then, output from U.S. mines has 

remained relatively low, and increases in demand have been 

satisfied mainly by imported barite. In 2011 (the latest year 

for which complete data were available), import reliance was 

78 percent (Miller, 2013).

Historically, domestic barite has come from mines in 

Arkansas, Georgia, Illinois, Missouri, Tennessee, and Nevada. 

At present, barite is produced primarily from mines in Nevada 

that exploit high-grade deposits; this barite requires only 

crushing, sizing, and jigging to meet the American Petroleum 

Institute (API) specifications for use in the oil industry. The 

API has two specifications for drilling-grade barite that differ 

only in specific gravity (SG) — 4.1 SG and 4.2 SG. Both specifications correspond to a barite content of about 90 weight 

percent. The 4.1 SG specification was introduced in 2010 at 

the request of Nevada barite producers to extend Nevada’s 

barite reserves and avoid the major capital investments that 

would be required to continue producing 4.2 SG material. 

The majority of barite now mined and processed in Nevada 

is 4.1 SG material. Deposits that are lower in grade or that 

contain deleterious gangue minerals can require more involved 

and more expensive beneficiation procedures, including wet 

grinding, magnetic separation, and froth flotation.

The active U.S. mines are open pit operations. They 

produce neither coproducts nor byproducts, so mine output 

depends exclusively on the economics of barite. At some mines 

in other countries, fluorite (CaF2

, also known as fluorspar), 

galena (PbS), sphalerite (ZnS), and (or) gold- or silver-bearing 

minerals are present and are recovered, so production at 

some locations can depend partly on the economics of other 

mineral commodities. Drilling activity in the Great Plains of 

the United States, in the Western United States, and in Western 

Canada are the major markets for barite produced from 

Nevada mines. Drilling activity in the Gulf of Mexico and in 

the eastern and midwestern parts of the United States relies 

heavily on imported barite, mainly from China.

A wealth of information on the geology and geochemistry 

of barite can be found in previous overviews by de Brodtkorb 

(1989), Clark and others (1990), Brobst (1973, 1994), Harben 

and Kuzvart (1997), Hanor (2000), and Jewell (2000). Information on barite as a mineral commodity has been provided by 

Ampian (1985), Coffman and Kilgore (1986), Harben (2002), 

Mills (2006), Roskill Information Services, Ltd. (2006), and 

Miller (2012).

Geology

Geochemistry

Barium occurs in the geologic environment mainly as the 

divalent cation Ba2+. This ion is larger than most other divalent 

cations, and, consequently, barium is not easily accommodated 

in common rock-forming minerals. During fractional crystallization of silicate magmas, barium becomes concentrated 

in the residual silicate liquid. It is also concentrated in 

silicate liquids produced by partial melting. The barium 

content of average upper continental crust is estimated to be 

0.0624 weight percent (Rudnick and Gao, 2003). Granitic 

rocks typically have somewhat higher barium content than 

average continental crust, and basaltic rocks typically have 

somewhat lower barium content. The range of barium content 

of shales spans approximately the same range as the barium 

content of granitic rocks.

Barium substitutes extensively for the nearly-as-large 

Pb2+ and Sr2+ cations in minerals that contain these elements, 

and less extensively for the somewhat smaller Ca2+ and K+

ions; substitution for K+ requires a coupled substitution to 

maintain charge balance. In common igneous rocks, barium is 

present as a trace or minor element in potassium feldspar and 

mica where it substitutes for K+. Less extensive substitution 

is seen for Ca2+ in amphibole, apatite, calcite, plagioclase 

feldspar, and pyroxene. In common sedimentary rocks and 

hydrothermal deposits, barium occurs mainly in barite or 

organic matter. The presence of trace amounts of barite in 

many sedimentary rocks reflects the fact that sulfate (SO4

2–) is 

the stable form of sulfur in most Earth-surface environments, 

and the formation constant for barite (barium sulfate, BaSO4 ) 

is quite large. Although barite is highly stable in oxidizing 

(sulfate-stable) environments, it can be readily dissolved in 

reducing hydrogen sulfide (H2S) stable environments. This 

attribute has important implications for the formation of barite 

deposits and for environmental aspects of barite mining.
Geology D3

Mineralogy

Barite is typically white or colorless. Crystals tend to be 

tabular and have habits that reflect the orthorhombic symmetry 

of the barite lattice. Fibrous, nodular, and massive forms are 

also common. The mineral has a hardness of 3 to 3.5 on the 

Mohs scale, which is about the same hardness as copper, and 

it has a density of 4.48 grams per cubic centimeter (g/cm3), 

which is roughly twice that of common rocks. Barite has 

multiple good cleavages.

Natural barite is relatively pure, generally showing only 

minor replacement of barium by strontium (less than 7 percent) 

or lead. The relative purity is convenient from a resource 

perspective because commercially important properties that 

depend on mineral chemistry, such as density and solubility, 

can be expected to vary little no matter where the barite is 

found. Most barite deposits have low impurity levels because 

(a) deposits typically form as precipitates from aqueous fluids, 

and (b) barium partitions more strongly into the sulfate mineral 

than strontium, lead, and other potential impurities. Calcium 

replacement of barium is also restricted in natural occurrences.

Deposits of barite invariably contain gangue minerals, the 

abundance and identity of which vary from one deposit type to 

another. The most common gangue minerals are quartz or other 

forms of silica (chert, drusy quartz, and jasperoid), sulfide minerals 

(galena, marcasite, pyrite, and sphalerite), clay minerals, carbonates (calcite and siderite), and iron oxides. Organic matter can also 

be present at concentrations up to several weight percent.

Witherite is colorless. Crystals commonly take the form 

of pseudohexagonal dipyramids. Witherite has a hardness of 

3.5 on the Mohs scale and a density of 4.3 g/cm3

. It is readily 

distinguished from barite by the fact that it effervesces in 

cold acids. The mineral typically contains small amounts of 

strontium and calcium substituting for barium. Witherite is 

normally found in veins within sedimentary rocks where it is 

associated with galena and, less commonly, anglesite, barite, 

and barytocalcite. Witherite can also form as an alteration 

product of barite and can itself be altered to barite.

Other barium minerals are known, but none occur in deposits 

as large, concentrated, and widespread as barite. Although 

these are not currently mined, some may merit consideration 

as future sources of barium, particularly where they occur 

in deposits that have other valuable minerals. Noteworthy 

examples are benstonite ((Ba,Sr)6 (Ca,Mn)6 Mg(CO3

)13 ), 

which is a secondary mineral in various barium-rich rocks 

(Hood and Steidl, 1973); sanbornite (BaSi2O5), which 

occurs in metasomatic zones associated with granitic 

rocks in the Big Creek-Rush Creek district of California 

(Alfors and others, 1965); and celsian (BaAl2Si2O8), which 

is locally abundant in a metamorphosed barium deposit 

near Aberfeldy, Scotland (Coats and others, 1980). Other 

minerals that can contain significant barium are todorokite 

((Na,Ca,K,Ba,Sr)0.3 –0.7 (Mn,Mg,Al)6O12 • 3.2– 4.5H2O), which 

is a common constituent of deep-sea ferromanganese nodules, 

and romanechite (BaMn9O16(OH)4), which is a mineral 

currently mined for manganese.

Deposit Types

Barite deposits can be classified into the following four 

groups on the basis of physical characteristics, geochemical 

characteristics, and geologic setting: (1) bedded-sedimentary; 

(2) bedded-volcanic; (3) vein, cavity-fill, and metasomatic; 

and (4) residual. This classification scheme resembles schemes 

that have been used previously (Brobst, 1958; Clark and 

others, 1990), but it has been modified somewhat to accommodate newly discovered deposits and results of more recent 

geologic studies. The locations of deposits or districts that 

are either relatively large or representative of the four groups 

are shown in figure D2. Numerous other deposits, both mined 

and unmined, have also been identified (Roskill Information 

Services, Ltd., 2006).

Bedded-Sedimentary Deposits

Bedded-sedimentary deposits are stratiform bodies of 

massive or near-massive barite within sedimentary successions 

that typically contain organic-rich shale, mudstone, or chert. 

The deposits can be quite large, with resources up to tens 

of millions of tons, and they tend to occur in districts rather 

than as isolated occurrences. The barite beds vary in thickness from centimeters to 100 meters or more and can extend 

for kilometers along strike. The barite can be dark in color, 

reflecting the presence of organic carbon, and it commonly 

occurs in one of the following four forms: massive, laminated, 

nodular, or rosette. Gangue minerals include clays, iron oxides 

(hematite, limonite, or magnetite), quartz, and pyrite or other 

sulfide minerals. Beneficiation can be simpler than for ores 

from other deposit types. Bedded-sedimentary deposits are the 

major source of worldwide barite production, and they also 

account for the majority of world reserves.

The bedded-sedimentary type of barite deposit can be 

associated with stratiform sulfide mineralization of sufficient 

quantity and grade that the districts are better known as 

resources for lead, zinc, or other metals. In principle, sulfide 

mineralization could be employed as an exploration guide 

to undiscovered barite deposits, but the economics normally 

favor the reverse: barite is employed as a guide to sulfide 

deposits. The nature of the association between barite and 

sulfides varies considerably. Barite and sulfide crystals 

can be intergrown, as in the Red Dog deposit in Alaska 

(Kelley and others, 2004); barite and sulfide layers can be 

thinly interbedded, as in the Tom and Jason deposits in the 

Selwyn basin of northwestern Canada (Goodfellow, 2004); 

or discrete barite bodies can lie adjacent to sulfide deposits, 

as in the Meggen and Rammelsberg deposits in Germany 

(Krebs, 1981; Large and Walcher, 1999), the Ballynoe and 

Silvermines deposits in Ireland (Mullane and Kinnaird, 

1998), and the Aggeneys-Gamsberg district in South Africa 

(McClung and others, 2007). Barite deposits can also occur in 

the same district as stratiform sulfide deposits but be separated 

from them by distances of hundreds or thousands of meters, 

as in the Tea, Gary, Moose, Pete, and Oro barite deposits of 
D4 Critical Mineral Resources of the United States—Barite (Barium) 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10° 20° 30°

40°50°

80°

70°

70°

80°

60°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Red Dog Cutaway basin

Nevada

Flagstaff Mountain

Ouachita

Mountains

Selwyn basin

Rammelsberg Meggen

Cuddapah

Qinling

Jiangnan

Sardinia

Aggeneys-Gamsberg

LasbelaKhuzdar

Chiganak

Neuquen-Mendoza

La Helvecia

Camamu

Sonora

NorthernThailand

Gertrudis-Jagua

Barite Hill

Zhayrem

Kuroko

Rosebery

Madneuli

Dorreh Kashan

Mountain Pass

Central

Missouri

Jbel Irhoud

Zelmou

Chordi

Araxá

Flinders Range

Chaillac

Mau Xe

Nghê An

Isparta

Southeast Missouri

Cartersville

Sweetwater

Jameson Land basin

EXPLANATION

Barite deposits

Bedded-sedimentaryBedded-volcanicVein, cavity-fill, metasomaticResidual

Barite, Fig D2.

-coded by deposit type. Where districts contain more than one deposit type, the symbol 

u (Cassedanne, 1989); Chaillac (Sizaret and others,

World map showing locations of selected barite deposits and districts, color Figure D2.

reflects the dominant type. Citations for published descriptions of the deposits are given in the text or are as follows: Camam

inogradov and others, 1978); Dorreh Kashan (Yaghubpur, 2003); Flinders Range (McCallum, 1990); Gertrudis-Jagua (Maynard and others, 1995); Isparta 2009); Chiganak (V

(Kiran Yildirim, 2010); Jameson Land basin (Thomassen, 2007); La Helvecia (de Brodtkorb and others, 1989); Lasbela-Khuzdar (Ahsan and Mallick, 1999); Neuquen-Mendoza 

, 2000); northern Thailand (Hahn and others, 1986), Harben and Kuzvart, 1997); Rosebery (Solomon and others, 1969); (de Brodtkorb and others, 1989); Nghê An (Kusnir

 (Padalino and others, 1989); and Zhayrem (Gauert, 2001). Sardinia
Geology D5

the Selwyn basin (Lydon and others, 1985; Butrenchuk and 

Hancock, 1997) and the Anarraaq, Gull Creek, Moil, and Su 

barite deposits of the Red Dog district (Johnson and others, 

2004). Other bedded-sedimentary barite deposits occur without 

sulfides. Examples of these include the deposits of the Ouachita 

Mountains of Arkansas (Zimmermann and Amstutz, 1989), the 

Nevada barite belt (Papke, 1984), the Cutaway basin in Alaska 

(Schmidt and others, 2009), the Sonora area of Mexico (Poole 

and others, 1991), the Cuddapah district in India (Neelakantam 

and Roy, 1979), and the Qinling and Jiangnan regions in 

southern China (Wang and Li, 1991; Wang, 1996).

Bedded-sedimentary barite associated with sulfide 

mineralization owes its origin to subsea-floor fluid flow 

systems that can span lateral distances of hundreds of kilometers or more; these flow systems are directly or indirectly 

related to the flow systems that give rise to the metal sulfide 

deposits (Lydon and others, 1985; Emsbo, 2000). The fluids 

typically migrate through pore spaces in buried sediments 

and sedimentary rocks toward extensional fault zones located 

along the margins of depositional basins. Barite deposits form 

where the ascending fluids, which carry Ba2+ ions scavenged 

from organic matter or from trace barite in the underlying 

sediments, encounter and mix with sulfate-bearing pore fluids 

or seawater. Where the mixing occurs within the sediment, 

barite precipitates as interstitial cement or replacements of 

preexisting minerals, such as calcite. Where the mixing occurs 

by fluid seepage onto the sea floor, barite particles form in the 

water column and then accumulate on the sediment surface. 

Because barite is not decomposed by seawater, preexisting 

deposits can be reworked by sea-floor erosion, and the barite 

can become reconcentrated in submarine channels.

The origin of bedded-sedimentary barite that lacks sulfide 

mineralization is more controversial. This deposit type has 

been attributed to an oceanographic process that operates 

beneath waters with high biological productivity (Jewell, 

2000). Decomposition of sinking organic matter promotes 

anoxic conditions by consuming dissolved oxygen, and it also 

releases organic-bound barium. The released barium combines 

with seawater sulfate to form barite, which accumulates on 

the sea floor at the interface between anoxic and oxic water 

masses. An alternative proposal (Lydon and others, 1985) 

holds that the deposits form at sea-floor vents for fluids whose 

chemical composition, temperature, or oxidation state are not 

conducive to sulfide precipitation. This proposal has been 

strengthened by the discovery of barite deposits on the modern 

continental margins at locations where sedimentary formation 

waters are emerging from fault-controlled seeps (Torres and 

others, 2003; Clark and others, 2004; Koski and Hein, 2004; 

Johnson and others, 2009).

Bedded-sedimentary barite deposits form in marine 

basins in association with sediments characteristic of high 

biological productivity. The deposits are most common in 

sedimentary strata of early or middle Paleozoic age, but they 

have also been found in strata of Archean, Proterozoic, and 

Mesozoic age. Concentrations of barite nodules have been 

identified in Cenozoic age rocks (Bogoch and others, 1987; 

Stamatakis and Hein, 1993), but there are no identified 

deposits of Cenozoic age that are of sufficient grade and 

size to consider mining. The presence or absence of sulfide 

mineralization has been proposed to reflect different tectonic 

settings for the depositional basins, one of which is underlain 

by continental crust—a source of zinc and lead (Maynard and 

Okita, 1991). This proposal has been controversial (Turner, 

1992), but tectonic setting undoubtedly influences sulfide 

formation in some way because it can affect important parameters of the ore-forming process, including the availability of 

heat to drive fluid movement, the nature of the fault systems 

that serve as channels for flow, and the structure and oxidation 

state of the overlying basin.

Bedded-Volcanic Deposits

Bedded-volcanic deposits are stratiform bodies 

of massive or near-massive barite within stratigraphic 

successions containing igneous rocks—typically both felsic 

and mafic volcanic varieties—and sedimentary rocks. The 

barite can occur in large high-grade deposits (up to 90 percent 

BaSO4) or can be intergrown or interbedded with copper, lead, 

zinc, or gold- or silver-bearing sulfide minerals. The deposits 

tend to occur in districts, and individual barite±sulfide bodies 

can be closely spaced enough that both barite and metals are 

exploited by a single mining operation. Barite resources at 

individual mines can range up to millions of tons. Gangue 

minerals are typically clays, quartz, or sulfides. Whereas 

bedded-volcanic deposits have been significant sources of 

barite in the past (in Japan, for example; Marumo, 1989), 

they are presently a minor source compared with beddedsedimentary deposits.

Examples of bedded-volcanic deposits are the kurokotype deposits of northern Honshu and southern Hokkaido, 

Japan (many of which are better known for their copper, lead, 

precious metals, and zinc; Marumo, 1989), the Barite Hill 

deposit in the Carolina slate belt of South Carolina (Clark 

and others, 1999), and deposits that occur in a belt extending 

westward from the Madneuli deposit in the Republic of 

Georgia through Turkey and Bulgaria to northern Algeria 

(Kekeliya and others, 1984; Migineishvili, 2005).

Bedded-volcanic barite deposits form at submarine 

volcanic centers at divergent, convergent, and, in some cases, 

transform plate boundaries (Marumo, 1989; Herzig and 

others, 1993; Hein and others, 1999). Barium is scavenged 

from the volcanic edifice and underlying rocks by seawater 

that is heated by magmas and circulated by convection. In 

convergent-margin volcanic-arc settings, additional fluid may 

be supplied to the convection system by crystallizing magmas. 

Barite deposits form where the hot barium-bearing fluids are 

expelled onto the sea floor and encounter cold sulfate-bearing 

seawater. Like the bedded-sedimentary deposits, the beddedvolcanic deposits have analogs on the modern sea floor, but 

these are located near active submarine volcanoes mostly 

away from continental margins (Kusakabe and others, 1990; 

Urabe and Kusakabe, 1990).
D6 Critical Mineral Resources of the United States—Barite (Barium)

Vein, Cavity-Fill, and Metasomatic Deposits

Barite deposits included in the vein, cavity-fill, and 

metasomatic classification occupy faults, breccia zones, 

solution-collapse, or other fluid-permeable structures in which 

barite has either filled open spaces or replaced adjacent rock. 

The deposits are commonly tabular and extend along structures, often discontinuously, that can be traced for hundreds or 

thousands of meters. The vein deposits can attain thicknesses 

of several tens of meters. Gangue minerals include carbonates 

(calcite, siderite), fluorite, quartz, and pyrite or other sulfide 

minerals. The deposits tend to occur in districts or in linear 

distributions. This deposit type has been a source of highpurity barite for use in fillers and ceramics, which generally 

requires a grade of at least 95 percent BaSO4. In some mining 

operations, barite is a byproduct or coproduct of fluorite, lead, 

or zinc production. The economic advantage of producing 

barite as a byproduct or coproduct from deposits of this type 

tends, however, to be offset by higher mining costs because 

underground mines or large open pits are required to exploit 

ore zones that commonly have steep dips.

Examples of vein barite deposits are found in the 

Culpeper, Hartford-Deerfield, and Newark basins in the 

Eastern United States (Robinson and Woodruff, 1988), the 

North Pennine and South Pennine orefields of England 

(Ixner and Vaughan, 1993), and the Greater Caucasus in 

the Republic of Georgia (particularly the Chordi deposit) 

(Tvalchrelidze and Shcheglov, 1990). Some vein barite 

deposits are located in districts that are better known for their 

lead-zinc sulfide deposits (Mississippi Valley-type), as in 

the Pennine orefields. A direct association with Mississippi 

Valley-type sulfide mineralization is characteristic of cavityfill barite deposits, such as those of the Central Missouri 

and Southeast Missouri districts, the Sweetwater district of 

Tennessee (Leach, 1980; Kaiser and others, 1987), and the 

Jbel Irhoud and Zelmou deposits of Morocco (Valenza and 

others, 2000). Cavity-fill deposits are more irregular in form; 

they tend to be restricted to a particular sedimentary stratum 

or sequence of strata. District-wide zoning is apparent in 

rare cases, as in the South Pennine orefield where a region of 

barite-dominant deposits is bounded on the east by a region 

of fluorite-dominant deposits and on the west by a region of 

calcite-dominant deposits.

Metasomatic barite deposits included in this grouping 

are associated with alkalic or carbonatite complexes. The 

best example in the United States is found at Mountain Pass 

in California, where a body of rock that has been mined for 

rare-earth elements contains approximately 20 to 25 percent 

barite (Castor and Nason, 2004). The barite is not presently 

recovered. The Araxá deposit, which lies within the alkalineultramafic Barreiro complex in Minas Gerais, Brazil, is 

capped by a 150- to 230-meter-thick regolith that has been 

leached of carbonate minerals and contains approximately 

20 percent barite (Woolley, 1987). The Mau Xe deposit in 

northern Vietnam is another carbonatite-associated deposit 

that is composed of about 51 percent barite (Kusnír, 2000). 

These deposits are resources for other mined commodities 

that are more valuable than barite—rare-earth elements for 

Mountain Pass and both rare-earth elements and niobium for 

Araxá and Mau Xe. Thus, any barite production from them 

would be as a byproduct or coproduct. The same is likely to be 

true for other examples of this deposit type.

In detail, deposits of the vein, cavity-fill, and metasomatic category have diverse origins, but they share a basic 

genetic attribute—the ore constituents are transported by 

one or more high-temperature fluids (100 degrees Celsius 

or hotter) that are channeled along structurally controlled 

pathways. Depending on the specific location, the source of 

ore constituents varies from predominantly sedimentary rocks 

to predominantly igneous rocks, as has been shown for the 

deposits of central Europe (Dill and Nielsen, 1987). Barite 

associated with carbonatite complexes is of igneous origin 

but can be either a precipitate from hydrothermal fluids or a 

crystallization product of a carbonate melt. The mechanism 

by which fluids were heated also varies from one location to 

another. The fluids that formed the U.S. midcontinent deposits 

were heated as they circulated to deep levels in an extensional 

tectonic setting where the geothermal gradient was elevated, 

whereas the fluids that formed the central Europe deposits 

were heated by igneous intrusions (Dill and Nielsen, 1987). 

In carbonatite-associated deposits, heat was supplied by 

intruded carbonatite magma. The tectonic settings are also 

diverse. Many vein deposits are found in narrow rift basins 

that contain red beds and other terrigenous sedimentary rocks, 

whereas cavity-fill deposits are commonly found in carbonate 

platforms that developed along the margins of marine basins. 

Carbonatites are unusual rocks that originate by low degrees 

of partial melting of mantle material beneath continental rifts.

Residual Deposits

Residual barite deposits are poorly consolidated 

materials that form by weathering of preexisting deposits. 

Clay minerals, quartz, and unweathered rock fragments are 

typically present in the deposits. Iron oxides, metal sulfides, 

and sulfide weathering products may also be present. 

Within the United States, residual deposits are found in the 

Cartersville area of Georgia (Reade and others, 1980), the 

Southeast Missouri district (Leach, 1980; Kaiser and others, 

1987), and the Sweetwater district of Tennessee (Kesler, 

1996). These deposits formed by weathering of cavity-fill 

barite mineralization contained in limestone or dolostone 

host rocks. Residual barite deposits can also form from 

weathering of other deposit types. For example, baritehematite gossan at the Mount Lyell Mine in western Tasmania 

was produced by weathering of bedded-volcanic barite and 

sulfide mineralization (Solomon and others, 1969). At Araxá, 

Brazil, the defined ore reserve is a weathered regolith leached 

of carbonate minerals (Woolley, 1987). At Mount Lyell and 

Araxá, weathering resulted in residual enrichment of barite 

and upgrading of the deposit.
Resources and Production D7

Resources and Production

Reserves

Barite reserves tend to be defined only a few years prior to 

production because (a) drilling activity—the major market for 

barite—fluctuates in ways that are difficult to forecast farther 

in advance, and (b) the ease of extraction and ore processing 

allows for faster and less expensive mine startup than for 

many other mined commodities. Consequently, there is a 

general tendency for mined reserves to be replaced regularly 

by new reserves. A boom in oil and gas exploration resulting 

from a combination of increasing global demand for oil and 

gas, high oil prices, and expanded use of hydraulic fracturing 

in development of unconventional oil and gas resources has 

resulted in dramatically increased global consumption of barite. 

Reserves of high-grade barite are being depleted, but they 

exceed demand by a large margin, both in the United States 

and worldwide. Thus, this mineral commodity is unlikely to 

become severely depleted in the foreseeable future.

Problems with costs of current sources of supply are 

evident, however. Consumers have become dependent on 

low-cost barite supplied by China and India, which together 

accounted for about 65 percent of world production in 2011. 

Prices for Chinese and Indian barite have risen substantially 

in recent years, with prices for barite from China increasing 

nearly sixfold in the past 10 years. Consumers are seeking 

alternative sources of low-cost barite to reduce the dependence 

on China and India for supplies. Global barite resources 

are extensive, but the exploration and development of these 

resources may not proceed in direct response to changes in 

the price or supply of the commodity. In recent years, even 

while barite prices rose steeply, little exploration took place. 

Currently, however, indications are that exploration and development work is expanding domestically and internationally in

such countries as Kazakhstan, Liberia, Mexico, and Zimbabwe.

An additional consideration is that transportation is a 

major factor in the total cost of barite (Coffman and Kilgore, 

1986). Long transport distances can add substantially to the 

overall price of the commodity, particularly if shipping is 

by rail or truck rather than by sea. Thus, supplies of lowcost barite for oil and gas producers depends partly on the 

proximity of barite mines to the drilling activity and whether 

transportation can be by low-cost ocean shipping.

Solutions to problems of tight supply and rising costs 

will likely include increased production from such countries 

as Kazakhstan, Mexico, Morocco, and Vietnam, and new 

production from nontraditional sources, such as Greenland and 

Zimbabwe. Other solutions that have been proposed include 

developing application-appropriate specifications and grades 

in order to preserve existing barite resources, using blended 

weighting agents, expanding the use of managed-pressure 

drilling techniques, and implementing recycling (Newcaster, 

2012). Barite consumption is forecast to increase by about 

30 percent by 2016, with likely fluctuations in the availability 

of supply.

Other Identified Resources

Recent production from existing mines, by country, 

is shown in figure D3. Total world production averaged 

7.97 million metric tons per year for the 5-year period starting 

in 2007. China was the leading producer (4,020 thousand metric 

tons per year, or 50 percent of global production), India was 

second (1,190 thousand metric tons per year, or 15 percent), 

Morocco was third (630 thousand metric tons per year, or 

7.9 percent), and the United States was fourth (574 thousand 

metric tons per year, or 7.2 percent). The top eight countries 

accounted for about 92 percent of global production.

In 2011, output from U.S. mines was 710,000 metric 

tons (Miller, 2013). For the foreseeable future, Nevada mines 

are expected to be the major domestic producers. Current 

producers have gathered extensive data on Nevada barite 

resources to plan future mining activities, and at least one 

producer was expanding production capacity. Renewed 

production is possible from Arkansas, which last mined barite 

in the early 1980s and has identified economic deposits in the 

Ouachita National Forest that could be developed. In addition, 

important new resources have been identified during the past 

20 years in northern Alaska. These are bedded-sedimentary 

barite deposits at the Red Dog district in the western Brooks 

Range and in the Cutaway basin in the central Brooks Range. 

The barite endowment of the Red Dog district could be in 

excess of 1 billion metric tons (Kelley and Jennings, 2004). 

These deposits are not currently mined, and most lie far from 

roads or other existing infrastructure; however, they are in 

close proximity to a region of active oil and gas exploration 

on the north slope of Alaska and to possible future exploration 

China

4,020 kt

India

1,190 kt

Morocco

630 kt

United States

574 kt

Iran 317 kt

Turkey 272 kt

Kazakhstan 174 kt

Mexico 155 kt 

All others 640 kt

Barite, Fig D3. Figure D3. Pie chart showing average annual barite production 

for the period 2007–11, by country and amount (in thousand 

metric tons [kt]). World production during this 5-year period 

averaged 7.97 million metric tons per year. Data are from Miller 

(2012).
D8 Critical Mineral Resources of the United States—Barite (Barium)

in Arctic regions of North America, Europe, and Asia. The 

Cutaway basin deposits lie within the National Petroleum 

Reserve–Alaska.

The largest identified resources of barite on a global 

scale are found in the Qinling and Jiangnan regions of 

southern China and the Cuddapah district of India. Tonnage 

estimates have varied widely, but all suggest that the amounts 

of barite available for mining are quite large. For example, 

recent estimates for one of the larger deposits in China, the 

Gangxi deposit in Hunan Province, have ranged from more 

than 80 million metric tons (Roskill Information Services, 

Ltd., 2006) to 453 million metric tons grading between 

63.6 and 77.3 percent BaSO4 (Kamitani and others, 2007). 

The compilation of published data by Kamitani and others 

(2007) lists seven barite deposits in southern China that 

exceed 5 million metric tons. The Mangampeta deposit in 

the Cuddapah district in India has been estimated to contain 

73.4 million metric tons with a specific gravity of 3.8 or 

greater, 50 percent of which has a specific gravity of 4.2 or 

greater (Clark and Basu, 1999).

Undiscovered Resources

The potential for the discovery of new barite resources 

in the United States and in many other locations in the world 

is considerable. Bedded-sedimentary deposits are probably 

the most attractive target for exploration because they have 

the potential to produce mud-grade barite with minimal ore 

processing. In addition, the most common geologic setting 

for this deposit type—organic-rich chert-mudstone-shale 

successions of early to middle Paleozoic age—exists in many 

regions of the world. Barite is easily overlooked or misidentified, so there is good potential for unrecognized deposits in 

recognized districts. Figure D2 provides a guide to a number 

of districts with current or past production from beddedsedimentary deposits, but many others likely occur.

Another potential source is byproduct barite or barium at 

mineral deposits that will be mined in the future primarily to 

recover other mineral commodities. Foremost among these are 

carbonatite deposits, which are potential sources of rare-earth 

elements and several other mineral commodities, and baritedominant hydrothermal deposits in relatively shallow water 

along active volcanic arc systems.

Exploration for New Deposits

Exploration for barite deposits has traditionally been 

carried out by direct examination of outcrop or by chemical 

analysis of stream sediment in areas where the geologic setting 

is favorable. Barium in stream sediment, in both the fine size 

fraction and the heavy-mineral fraction, has proven to be a 

good guide to bedded mineralization (Coats and others, 1981).

Geophysical methods have been used to delineate 

deposits or to identify barite accumulations that are shallowly 

buried. Gravity methods exploit the large density contrast that 

exists between massive barite bodies and most of the rock 

types that enclose them (Uhley and Scharon, 1954; Schmidt 

and others, 2009). Very-low-frequency electromagnetic 

methods have also provided useful information where barite, 

which has low resistivity, is contained in conductive host rocks 

(Coats and others, 1981).

Environmental Considerations

Barium Behavior in the Environment

Barium has limited mobility in the environment because 

it tends to form insoluble compounds under common 

Earth-surface conditions and does not form soluble organic 

complexes. The barium released from minerals during weathering tends to precipitate as barite (barium sulfate, BaSO4), 

witherite (barium carbonate, BaCO3), or other minerals, or 

to adsorb onto clay minerals, oxides, or hydroxides. Barite 

is practically insoluble in oxidizing environments and is 

highly resistant to weathering. The solubility of most barium 

compounds increases as conditions become reducing or acidic 

(Jaritz, 2004).

Barium in the environment comes mostly from 

natural rock weathering, but anthropogenic contributions 

can be important, particularly emissions related to copper 

smelting, steel production, and automobile manufacturing 

(Reimann and de Caritat, 1998). Combustion of coal and 

diesel and incineration of waste both release bariumbearing particulates to the atmosphere (Agency for Toxic 

Substances and Disease Registry, 2007). In the United States, 

atmospheric barium concentrations have been reported to 

range from 0.00015 to 0.95 microgram per cubic meter 

(U.S. Environmental Protection Agency, 1984). Barium 

concentrations in soil usually reflect the composition of the 

underlying bedrock, but dispersal of fly ash or sludge from 

landfills and application of phosphate fertilizers can lead to 

elevated levels of barium in the soil (International Programme 

on Chemical Safety, 2001; Kabata-Pendias and Mukherjee, 

2007). Surface soils (top 5 centimeters [cm]) in the conterminous United States average 518 parts per million (ppm) 

barium with values ranging from less than 5 to 4,770 ppm 

(parts per million equals grams per metric ton; 1 ppm equals 

0.0001 percent; Smith and others, 2013).

The concentration of barium in natural waters is 

controlled by the solubility of barium compounds, the 

tendency of barium to adsorb onto particulates, and the 

availability of dissolved sulfate or carbonate to form insoluble 

salts (Agency for Toxic Substances and Disease Registry, 

2007). The concentration of barium in surface waters ranges 

from 7 to 15,000 parts per billion (ppb) (Jaritz, 2004). 
Environmental Considerations D9

Gaillardet and others (2003) reported a worldwide river water 

average of 23 ppb. Suspended solids and sediments commonly 

contain more barium than the water. The global average for 

barium in suspended river sediment is 522 ppb (Viers and 

others, 2009). A study of the Rhine River near Darmstadt 

showed a barium concentration of 480 ppb in the suspended 

material compared with 39 ppb in the water (Hoffmann and 

Lieser, 1987). Barium in seawater averages 6 ppb, which is 

lower than the concentration in most freshwater because barite 

tends to precipitate as freshwater enters the ocean (Jaritz, 

2004). The barium content in seawater varies among different 

oceans and with latitude and depth (International Programme 

on Chemical Safety, 2001).

The contents of barium and other elements in waters, 

sediments, and soils in the vicinity of barite deposits can be 

above average, depending on the type of deposit, the extent 

of the outcrop of ore and overburden, the climate, and other 

factors. Barium can be leached by groundwaters in some 

locations, such as in regions of Kentucky, northern Illinois, 

New Mexico, and Pennsylvania, where the presence of barite 

in bedrock has given rise to relatively high barium concentrations in groundwater (Agency for Toxic Substances and 

Disease Registry, 2007).

Barite in drilling muds can be recycled in some 

situations. The procedure involves removing barite from the 

mud by centrifuging and returning it to the active drilling 

system; this recycling of barite can help ensure compliance 

with fluid discharge requirements in environmentally sensitive areas. For the most part, barite is recycled only because 

drilling mud is continuously circulated through a reserve pit to 

allow cuttings to settle, and the mud is then reused during the 

drilling of a particular well.

The U.S. Environmental Protection Agency classifies oil 

and gas drilling fluids as “special wastes,” which are exempt 

from amendments to the Resource Conservation and Recovery 

Act (P.L. 94–580) issued by Congress in 1980. As a result, 

the methods for handling and disposal of oilfield waste fluids 

vary from one State to another (McFarland and others, 2009). 

Various methods are employed to dispose of drilling fluids 

containing barite, including onsite burial, landfill disposal, 

land application, salt cavern injection, and disposal at sea. 

The most common practice on land is onsite burial. After 

well completion, the fluid is pumped into the reserve pit, 

allowed to dry, and then either mixed with soil from the pit 

berms or simply covered with soil. At other locations, drilling 

fluids are transported to a landfill or are land-applied, which 

allows naturally occurring microbes in the soil to metabolize, 

transform, and assimilate the waste constituents. Salt cavern 

injection involves pumping waste drilling fluids into cavities 

in underground salt deposits. Specific regulations apply 

to ocean disposal of drilling fluids, particularly for fluids 

that are oil based or synthetic (Drilling Waste Management 

Information System, undated).

Barium in electronic devices and other manufactured 

products is not recovered. Recycling supplies less than 

1 percent of worldwide barium consumption (Graedel and 

others, 2011).

Mine Waste Characteristics

The solid wastes produced at barite mines vary widely in 

volume, mineralogy, and chemistry because of the variation 

in the types and sizes of deposits, in the compositions of the 

host rock and gangue, and in the ore processing methods. 

At bedded deposit mines, barite of high purity can often be 

obtained with minimal processing. Waste from operations of 

this type consists of host rock and gangue minerals. Common 

host rock types include alluvium, black shale, chert, dolostone, 

limestone, mudstone, and siltstone; igneous rocks can be 

important in some locations. Common gangue minerals 

include carbonates, clays, iron oxides, quartz, and minor 

sulfide minerals. Sulfide minerals are typically sparse enough 

that waste rock piles and tailings impoundments have low 

potential for acid drainage, as at the Argenta Mine in Nevada 

(U.S. Bureau of Land Management, 2010). In locations 

where bedded deposits are hosted by dolostone or limestone, 

the potential for acid drainage is lessened by the capacity of 

these rock types to neutralize acidity. This phenomenon may 

be responsible for the lack of acidity in groundwaters in the 

vicinity of the Mangampeta Mine in India (Nagaraju and 

others, 2006a). Waste rock can be of sufficiently high quality 

that the material is suitable for use as aggregate. On the other 

hand, wall rocks exposed in open pits can be acid-generating 

(U.S. Bureau of Land Management, 2010), in which case, pit 

lakes that develop after mine closure warrant remediation. 

In addition, soils developed on waste piles can have elevated 

concentrations of barium and other metals contained in the 

host rocks, as in the region surrounding the Mangampeta Mine 

in India where soils locally contain significant amounts of 

boron, chromium, and nickel (Raghu, 2001).

Vein barite deposits and bedded barite closely associated 

with massive sulfide mineralization normally require froth 

flotation to separate barite from fluorite, galena, pyrite, 

sphalerite, or other minerals. At these operations, waste rock 

and tailings can have higher sulfide contents and correspondingly greater potential for acidic metal-bearing drainage. The 

acid-neutralization capacity of host rocks and gangue minerals 

can be a significant control of acid drainage. Granite host 

rocks, for example, have less acid-neutralizing capacity than 

limestones or dolostones. At mines where carbonate rocks are 

important constituents of the bedrock, acid drainage can be 

lessened naturally, such as at the Magnet Cove copper-leadzinc-gold-barite mine in Nova Scotia, where surface waters are 

acidic only very locally (Whitehead and Macdonald, 1998). 

Unlike copper, iron, lead, manganese, zinc, and many other 

metals, barium is not particularly mobile in acid drainage 

because barite resists dissolution. Dispersal of barium from 

mining operations is mainly by physical erosion and transport 

of barite particles.
D10 Critical Mineral Resources of the United States—Barite (Barium)

Human Health Concerns

No adverse human health effects have been linked to 

barite mining in the United States. In general, the toxicity 

of barium depends on its chemical form. Soluble barium 

compounds, such as barium chloride, barium hydroxide, and 

barium nitrate, can be toxic to humans, animals, and plants, 

whereas barium sulfate (barite) is effectively nontoxic because 

the barium is not bioaccessible. In fact, barium sulfate is 

routinely ingested by patients undergoing gastrointestinal 

X-rays (Jaritz, 2004). Barium carbonate is only slightly 

soluble in water, but it is toxic to humans because it dissolves 

in the gastrointestinal tract (Agency for Toxic Substances 

and Disease Registry, 2007). There is no evidence that 

barium compounds are carcinogenic or cause genotoxicity in 

humans or animals, although the available data are limited 

(International Programme on Chemical Safety, 2001).

The barium content of drinking water, food, and soils is 

rarely high enough to present a human health concern. Dietary 

intake of barium is mostly from drinking water. The maximum 

contaminant level for drinking water in the United States 

is 2.0 ppm (U.S. Environmental Protection Agency, 2009), 

whereas the World Health Organization guideline is 0.7 ppm 

(World Health Organization, 2004). Drinking waters in 

the United States have a median barium concentration of 

0.043 ppm, and more than 90 percent of U.S. cities were 

found to contain less than 0.1 ppm (Dufor and Becker, 1964). 

High concentrations of barium in drinking water can result 

in gastrointestinal disturbances and muscular weakness, and 

lead eventually to high blood pressure and cardiovascular 

disease (Agency for Toxic Substances and Disease Registry, 

2007; Health Canada, 2012). For barium concentrations up 

to 10 ppm, however—which have been observed in certain 

regions of the United States, including Illinois and Iowa 

(Brenniman and others, 1981)—the links to high blood 

pressure and cardiovascular risk factors have been called 

into question (Wones and others, 1990). In some situations, 

food can be a significant source of dietary barium. Dairy and 

meat products tend to be lower in barium than vegetables, 

but Brazil nuts, seaweed, fish, and some plants can be high 

in barium (Jaritz, 2004; Agency for Toxic Substances and 

Disease Registry, 2007). Consumption of vegetables grown 

on barite-contaminated garden soils has been suggested as a 

human exposure pathway near a barite mine in northwestern 

Croatia (Frančišković-Bilinski and others, 2007). For soils, the 

noncancer screening level is 1,500 ppm in residential settings 

and 19,000 ppm in industrial settings (U.S. Environmental 

Protection Agency, 2012). The level for industrial settings 

is many times higher than the 4,770 ppm maximum 

concentration reported in a survey of soils in the conterminous 

United States (Smith and others, 2013), but the level for 

residential settings is exceeded by some of the reported 

soil concentrations.

The individuals at greatest risk of barium exposure are 

workers at mines, ore processing plants, and manufacturing 

facilities for barium-containing products. Epidemiological 

studies have reported cases of baritosis stemming from 

dust exposure during mining and grinding of barium ores 

(Pendergrass and Greening, 1953; Doig, 1976; Seaton and 

others, 1986). Baritosis is a benign lung disease that does 

not result in medical disability and appears to be reversible 

on a time scale of several years (Seaton and others, 1986). 

In the United States, workplace exposure to soluble barium 

compounds is limited to 500 micrograms of barium per cubic 

meter of air per 8-hour period (Occupational Safety and 

Health Administration, 2013). Some barite deposits contain 

enough quartz that respirable silica can be a concern at mine 

and mill sites. Risks can be mitigated, however, by using dust 

masks and water sprayers or other dust suppression techniques 

(Mills, 2006).

Ecological Health Concerns

No adverse toxicological effects of barium on plants or 

wildlife have been reported near barite mines or elsewhere. 

Barium accumulates only slightly from soil to plant and from 

plant to animal (International Programme on Chemical Safety, 

2001); however, high soil barium contents have been shown 

to inhibit growth in some terrestrial plants (International 

Programme on Chemical Safety, 2001). For the United States, 

a soil screening guideline of 500 ppm has been proposed 

for phytotoxicity (Efroymson and others, 1997), but this 

benchmark has only limited data to support it. Some soluble 

barium compounds are toxic to animals; for example, barium 

carbonate—the mineral witherite where it occurs naturally—is 

sufficiently toxic that it is used as a rodenticide (Jaritz, 2004).

In the aquatic environment, barium can have toxic effects 

on some organisms (water fleas, or Daphnia magna), but 

limited data suggest that the risk to fish and aquatic plants is 

likely to be less. Barium accumulates only slightly in aquatic 

life (Reimann and de Caritat, 1998). Suter (1996) proposed 

secondary acute and chronic water-quality benchmarks for 

freshwater biota of 69.1 and 3.8 ppb, respectively, but these 

values were determined using fewer data than normally 

required to establish levels. Lethal concentrations for water 

fleas and freshwater amphipods (Hyalella azteca; 50 percent 

mortality in laboratory tests) are greater than 1,000 ppb, 

which is significantly higher than the proposed benchmark 

(Borgmann and others, 2005).

Ecological health concerns associated with the mining 

and processing of barite ores arise more from disturbance of 

host rocks and associated minerals than from the barite itself. 

The carbonaceous shales or siltstones associated with many 

bedded-sedimentary barite deposits typically contain minor 

amounts of pyrite (1 to 2 weight percent on average) and 
Problems and Future Research D11

significant amounts of some metals (chromium, manganese, 

vanadium, and zinc typically exceed 100 ppm) (QuinbyHunt and others, 1989). Weathering of these lithologies in 

waste rock piles can potentially lead to acidic metal-bearing 

drainage, the release of which can have deleterious effects 

on aquatic life. The same potential would exist at barite 

mining operations that treat sulfide-bearing ores and produce 

sulfide-bearing tailings. As discussed, the presence of acidneutralizing rocks would tend to lessen this problem.

Drilling mud is disposed of at or near the drilling site. 

Dispersal of this material can have deleterious effects on 

plants, including slowed growth of both seeds and plant, and 

reduction in yields, although these effects could also relate 

to other constituents of drilling mud (Ferrante, 1981). Barite 

alone has been shown to cause toxicity to freshwater fish only 

at very high concentrations (Ferrante, 1981). In some jurisdictions, such as Alaska, restrictions are placed on the metal 

content of the barite used in drilling mud (less than 3 ppm 

cadmium and less than 1 ppm mercury; Harben, 2002).

Mine Closure

Most active barite mines worldwide, and all active mines 

in the United States, are open pit operations. After mining 

ceases, open pits can be left open and fenced. A pit lake will 

form if the depth of the excavation exceeds the depth of the 

water table. The lake water may or may not be acidic and 

metal-bearing depending on the abundance of sulfide minerals 

and the acid-neutralizing capacity of rocks exposed in the pit 

walls and immediate vicinity. Poor water quality may require 

that the lake be drained and the water treated, followed by 

backfilling of the pit. Openings to any underground workings 

are typically sealed.

Waste rock piles, mill sites, and other surface disturbances are graded and contoured. The area is typically 

revegetated by seeding or planting. At least one scrub species 

(Prosopis juliflora) grows prolifically and sequesters barium 

and metals (Nagaraju and others, 2006b). For some waste rock 

and tailings, a possible alternative to grading and revegetation is removal and sale of the material for use as aggregate 

or fill. This has been common practice at barite mines in 

France, Germany, Spain, and the United Kingdom (European 

Commission, 2009).

Problems and Future Research

Several unresolved questions regarding the origin of 

bedded-sedimentary barite deposits have implications for 

exploration. Perhaps the most important is whether the barium 

is derived from organic matter that decomposes as it sinks to 

the floor of a sedimentary basin or from fluids ascending faults 

or other structurally controlled pathways toward the basin 

floor. The correct answer to this question may differ from one 

district to another and will dictate whether new deposits are 

likely to be sought in locations where water column oxicanoxic transitions intersected the basin floor or in the vicinity 

of basin-bounding faults.

Other outstanding questions regard the relationship 

of barite deposits to deposits of other useful minerals. The 

reason that there is a close spatial and temporal relationship 

between bedded-sedimentary barite deposits and zinc-leadsilver sulfide deposits in some sedimentary basins but not in 

others is not fully understood. Evidence has been presented 

that bedded-sedimentary barite deposits are one member of 

a large family of mineral deposits in sedimentary basins that 

includes not only lead-zinc-silver deposits but also deposits 

of gold, phosphate rock, vanadium, and several other mineral 

commodities (Emsbo, 2000; Emsbo and others, 2005). If 

predictable relationships are identified among these deposit 

types on the basis of accurate models for their origin, then 

explorers could likely be able to identify stratigraphic intervals 

in sedimentary basins that present the highest probability of 

new discoveries for a variety of mineral commodities.

Finally, if pursued, advances in mineral processing could 

make it possible for barite to be profitably recovered as a 

coproduct or byproduct where ores are being mined for zinc, 

lead, rare-earth elements, or other mineral commodities. These 

advances could help meet the broader goal of increasing the 

efficient use of mined materials and reducing the amount of 

mining and ore processing waste that requires remediation.
D12 Critical Mineral Resources of the United States—Barite (Barium)

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Beryllium

By Nora K. Foley, Brian W. Jaskula, Nadine M. Piatak, and Ruth F. Schulte

Chapter E of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–E

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter E

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................ E1

Introduction.................................................................................................................................................. E1

Strategic and Critical Resource Issues ......................................................................................... E3

Uses and Applications ...................................................................................................................... E3

Geology......................................................................................................................................................... E6

Geochemistry...................................................................................................................................... E6

Mineralogy.......................................................................................................................................... E7

Deposit Types ..................................................................................................................................... E9

Magmatic-Related Beryllium Deposits ................................................................................. E9

Rare-Metal Pegmatite Deposits—Beryllium Type ..................................................... E9

Skarn and Greisen-Hosted Beryllium-Fluorine Deposits ........................................ E13

Carbonate-Hosted Beryllium Deposits ....................................................................... E14

Volcanogenic-Hosted Beryllium Deposits ................................................................. E14

Nonmagmatic-Related Beryllium Deposits ........................................................................ E16

Resources and Production ...................................................................................................................... E16

Domestic Resources ....................................................................................................................... E16

International Resources ................................................................................................................. E18

Production and Supply .................................................................................................................... E18

Exploration for New Deposits ................................................................................................................. E19

Environmental Considerations ................................................................................................................ E20

Sources and Fate in the Environment .......................................................................................... E20

Mine Waste Characteristics .......................................................................................................... E21

Human Health Concerns ................................................................................................................. E23

Ecological Health Concerns ........................................................................................................... E24

Carbon Footprint............................................................................................................................... E24

Mine Closure..................................................................................................................................... E25

Problems and Future Research .............................................................................................................. E25

References Cited....................................................................................................................................... E26

Suggested citation:

Foley, N.K., Jaskula, B.W., Piatak, N.M., and Schulte, R.F., 2017, Beryllium, chap. E of Schulz, K.J., DeYoung, J.H., Jr., 

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. E1– E32, http://dx.doi.org/

10.3133/pp1802E.
Beryllium

By Nora K. Foley, Brian W. Jaskula, Nadine M. Piatak, and Ruth F. Schulte

Abstract

Beryllium is a mineral commodity that is used in a 

variety of industries to make products that are essential for 

the smooth functioning of a modern society. Two minerals, 

bertrandite (which is supplied domestically) and beryl (which 

is currently supplied solely by imports), are necessary to 

ensure a stable supply of high-purity beryllium metal, alloys, 

and metal-matrix composites and beryllium oxide ceramics. 

Although bertrandite is the source mineral for more than 

90 percent of the beryllium produced globally, industrial beryl 

is critical for the production of the very high purity beryllium 

metal needed for some strategic applications. The current 

sole domestic source of beryllium is bertrandite ore from the 

Spor Mountain deposit in Utah; beryl is imported mainly 

from Brazil, China, Madagascar, Mozambique, and Portugal. 

High-purity beryllium metal is classified as a strategic and 

critical material by the Strategic Materials Protection Board of 

the U.S. Department of Defense because it is used in products 

that are vital to national security. Beryllium is maintained in 

the U.S. stockpile of strategic materials in the form of hotpressed beryllium metal powder.

Because of its unique chemical properties, beryllium is 

indispensable for many important industrial products used in 

the aerospace, computer, defense, medical, nuclear, and telecommunications industries. For example, high-performance 

alloys of beryllium are used in many specialized, hightechnology electronics applications, as they are energy efficient and can be used to fabricate miniaturized components. 

Beryllium-copper alloys are used as contacts and connectors, 

switches, relays, and shielding for everything from cell phones 

to thermostats, and beryllium-nickel alloys excel in producing 

wear-resistant and shape-retaining high-temperature springs. 

Beryllium metal composites, which combine the fabrication 

ability of aluminum with the thermal conductivity and highly 

elastic modulus of beryllium, are ideal for producing aircraft 

and satellite structural components that have a high stiffnessto-weight ratio and low surface vibration. Beryllium oxide 

ceramics are used in a wide range of applications, including 

missile guidance systems, radar applications, and cell phone 

transmitters, and they are critical to medical technologies, 

such as magnetic resonance imaging (MRI) machines, medical 

lasers, and portable defibrillators.

The United States is expected to remain self-sufficient 

with respect to most of its beryllium requirements, based on 

information available at the time this chapter was prepared 

(2013). The United States is one of only three countries that 

currently process beryllium ores and concentrate them into 

beryllium products, and these three countries supply most 

of the rest of the world with these products. Exploration for 

new deposits in the United States is limited because domestic 

beryllium production is dominated by a single producer that 

effectively controls the domestic beryllium market, which is 

relatively small and specialized, and the market cannot readily 

accommodate new competition on the raw material supply side.

Introduction

This chapter provides an overview of the main types 

of mineral resources from which beryllium is extracted and 

refined. Beryllium is a mineral commodity that is used in 

a variety of industries to make products that are important 

for modern society. It updates the work of Griffitts (1973) 

on the history and then-current status of domestic resources 

of beryllium. Advances in understanding the geochemistry 

of beryllium (table E1), the mineralogy (table E2), and 

the genesis of current and potential sources of beryllium 

(tables E3 and E4), the global distribution of beryllium 

resources (fig. E1), and the consequences of mining these 

resources all contribute to a better sense of the factors that can 

influence beryllium supply and demand in a global market. 

Two distinct classes of deposits currently account for most 

beryllium ores; they are (a) volcanic and carbonate-hosted 

deposits that contain the mineral bertrandite, and (b) pegmatite 

deposits that have an abundance of the mineral beryl (fig. E2). 

The primary global provider of bertrandite is a single worldclass deposit of the volcanogenic beryllium type located in 

the United States near Spor Mountain, Utah. Beryllium is also 

derived from beryl found in rare-metal pegmatite deposits 
E2 Critical Mineral Resources of the United States—Beryllium 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10°

20°30°40°50°

80°

70°

70°

80°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

60°

Greenbushes

Mount

Cattlin

Aracuai

Minas Gerais

Yellowknife field

Thor Lake

Tanco James Bay (Lithium One) Québec Lithium

Barkam

Jiajika

DEMOCRATIC

REPUBLIC OF

THE CONGO

Manono Kitolo

Ullava Länttä

JAPAN

Donkerhoek

Vishnyakovskoe

Kings Mountain

Bessemer City

Kamativi

Bikita Field

BRAZIL

NAMIBIA

ZIMBABWE

MADAGASCAR

Sahatany Field AUSTRALIA

NORWAY

SWEDEN

Nanping

CHINA

RUSSIA

CANADA

Spor

Mountain, 

Utah

Yermakovskoye

UNITED

STATES

Zambezia

MOZAMBIQUE

Lost River, 

Alaska

Brockman

Kobe

Beryllium deposit

Bertrandite and (or) phenakite Beryl

EXPLANATION

Beryllium fig 1

 : bertrandite and (or) phenakite, and 
Introduction E3

A

B

Fig 2

Figure E2. Photographs showing the minerals bertrandite, 

which can contain up to 42 percent beryllium oxide, and 

industrial beryl, which can contain up to about 5 percent 

beryllium. Photograph A is a nodule of fluorite, opal, and 

quartz that contains about 1 percent beryllium as bertrandite 

(Foley and others, 2012). Photograph B is beryl from a pegmatite 

of the lithium-cesium-tantalum family and shows typical green 

coloration in an assemblage of quartz and mica (U.S. Geological 

Survey archive).

(mainly of the lithium-cesium-tantalum [LCT] family), which 

are widely distributed throughout the world. These pegmatite 

ores provide the beryllium industry with the beryllium metals 

and metal-oxides of various levels of purity that are essential 

in the production of many special alloys and ceramics. In 

addition, some isotopes of beryllium metal (including 10 Be 

and 7 Be) are used by research scientists to understand many 

geologic processes. Some important uses of beryllium isotopes 

are as a petrogenetic index to trace (a) the connection between 

sediment subduction and island arc volcanism to show whether 

sediments are younger than about 3 to 4 months in age, and 

(b) climatic changes as reflected in soils, snow, and ice.

Strategic and Critical Resource Issues

High-purity beryllium metal is classified as a strategic 

and critical material by the Strategic Materials Protection 

Board of the U.S. Department of Defense (DOD) because 

it is used in products that are vital to national security 

(U.S. Department of Defense, 2009). The DOD is required 

by the Board to continue to take special actions to maintain 

a long-term domestic supply because domestic beryllium 

production capabilities have diminished over time. As a result, 

a partnership between the DOD and Materion Corp. of Ohio 

was established to ensure a long-term stable domestic supply 

of primary beryllium metal, and the United States continues 

to maintain a stockpile of beryllium metal for use during a 

national emergency. As of December 31, 2011, the National 

Defense Stockpile goal for hot-pressed beryllium metal 

powder was 45 metric tons. In 2015, the goal, now updated to 

include vacuum-cast beryllium metal, was 47 metric tons. In 

2015, the inventory held 78 metric tons, a surplus of 31 metric 

tons. Although bertrandite (fig. E2A) is the source mineral 

for more than 90 percent of beryllium produced globally, 

industrial beryl (fig. E2B), which has been supplied in recent 

years by imports, is critical in the production of ultra-highpurity beryllium metal. Ultra-high-purity beryllium is made 

exclusively from beryl, as beryl typically has fewer impurities 

(for example, fluorine and uranium) than bertrandite. Berylsourced ultra-high-purity beryllium is most effectively used 

in nuclear applications, where the absence of uranium in the 

beryl allows for safe and timely disposal of nuclear waste, 

and in foil for use as X-ray windows for medical applications. 

Thus, at the present time, both volcanogenic beryllium and 

LCT pegmatite resources are needed to ensure a stable supply 

of beryllium for strategic and critical applications.

Uses and Applications

Because of its chemical properties, beryllium metal is 

indispensable in many industrial applications. Products made 

from beryllium (fig. E3) are used in the aerospace, computer, 

defense, medical, nuclear, and telecommunications industries 

(Jaskula, 2013a).

Complex processing techniques and high costs (the 

price of beryllium generally ranges from $150 to $225 per 

pound) restrict the use of beryllium to applications where 

performance is critical and for which there are no practical 

alternatives. Although it is such a light metal, beryllium is 

strong and refractory; it is also highly resistant to corrosion 

because it readily develops an oxide film, and it has high 

thermal conductivity (table E1). Strategic military applications for high-purity beryllium metal include housings for 

optical targeting systems used for real-time imagery and 

surveillance in both manned helicopters and unmanned aerial 

systems and beryllium mirrors for ensuring vibration-free 
E4 Critical Mineral Resources of the United States—Beryllium

A. James Webb Space Telescope

B. Be-Cu alloy connectors

C. Al-In-Be-P LEDs 

Beryllium Figure 3

beryllium. A, Beryllium metal has been used to make mirrors of 

powerful telescopes. B, Beryllium-copper (Be-Cu) alloys are 

used to make contacts and connectors, switches, relays, and 

shielding for everything from cell phones and computers to 

thermostats, high-definition televisions, and automobiles, as well 

as electrical connectors necessary for next-generation Thermal 

Ionization Mass Spectrometers, and electrical contacts in cell 

phones and computers. C, Metal matrix composites of beryllium 

are used in high-performance electronics for aluminum-indiumberyllium-phosphorus light-emitting diode (LED) pins, which provide

superior performance in high-voltage heat-producing circuits. 

(Photograph A is courtesy of Ball Aerospace; photographs B and C

are courtesy of Spectromat Massenspektrometer GmbH.)

Figure E3. Photographs illustrating some of the many uses of 

paths for targeting and firing controls in battle tanks and 

guided missile defense systems. Beryllium metal is also 

integral to the airborne equipment used to detect and destroy 

improvised explosive devices (IEDs) and tactical mines. 

Because beryllium metal is transparent to X-rays, it is also 

used to produce metallic glasses, thin foils, and mirrors. 

Beryllium glasses and foils are used in medical X-ray imaging 

and detector applications, for beryllium mirrors in satellites, 

for telescope optics, and in optical guidance systems. 

For example, mirror segments for the James Webb Space 

Telescope (fig. E3A) are made from beryllium.

When combined with copper, aluminum, and other 

metals, beryllium is used to produce light and very strong 

alloys needed for aerospace, automobile, and computer technologies; oil and gas drilling equipment; musical instruments; 

medical devices; and telecommunications equipment. Alloys 

are used in many specialized electronics applications, as they 

are energy efficient and can be used to fabricate miniaturized 

components. Beryllium-copper alloys (fig. E3B ) are used to 

make contacts and connectors, switches, relays, and shielding 

to everything from cell phones and computers to thermostats, 

high-definition televisions, and automobiles. Beryllium-nickel 

alloys excel in producing wear-resistant and shape-retaining 

high-temperature springs, contacts, and connectors. Berylliumaluminum alloys are prepared for applications where 

critical weight, stiffness, and thermal specifications must be 

met. Beryllium-metal composites (fig. E3C) combine the 

fabrication ability of aluminum with the thermal conductivity 

and highly elastic modulus of beryllium. These qualities 

are valuable for producing aircraft and satellite structural 

components that have a high stiffness-to-weight ratio and low 

surface vibration.

Beryllium oxide is an excellent refractory material that 

is produced from beryllium metal, and it has some interesting 

uses in ceramics. In high-performance electronics, heat 

dissipation and electrical insulation are of critical importance 

for the electrical circuits to function effectively. Beryllium 

oxide ceramics provide superior thermal management 

performance in heat-producing circuits, such as those carrying 

high currents, where there is little airflow, or where the circuits 

are exposed to high ambient temperatures. These types of 

ceramics are used in a wide range of applications, including 

missile guidance systems, radar applications, and cell phone 

transmitters. Beryllium oxide in ceramic form is used to make 

high-performance semiconductor parts for radio equipment 

because it has good thermal conductivity and is a good 

electrical insulator. Semiconductor devices that use beryllium 

oxide ceramics instead of aluminum oxide between the 

silicon chip and the metal mounting base tend to last longer 

because the beryllium oxide ceramics provide more heat 

resistance. Beryllium ceramics are also vital to critical medical 

technologies, such as magnetic resonance imaging (MRI) 

machines, medical lasers, and portable defibrillators, because 

they are used to build parts of high-performance microwave 

devices, vacuum tubes, magnetrons, and gas lasers that are 

used in such medical devices.
Introduction E5

The uses of materials produced from beryllium in certain 

specialized applications are expected to expand rapidly with 

advances in technology. For example, work is underway to 

develop a viable manufacturing process for nuclear fuel that 

contains both beryllium and uranium oxides because, in spite 

of the high cost of beryllium relative to that of uranium (the 

price of U235 was about $43 per pound in 2013 compared 

with beryllium’s cost of $150 to $225 per pound), beryllium 

fuels may be longer lasting, more efficient, and safer than 

conventional nuclear fuels (Kim and others, 2010). Because 

the cost of beryllium is high compared with the cost of many 

other materials, less expensive materials are substituted where 

possible. For example, in some less-demanding applications, 

copper alloyed with Ni-Si, Sn, Ti, or Sn-P may be substituted 

for high-cost beryllium-copper alloys, and aluminum nitride or 

boron nitride may be substituted for high-cost beryllium oxide 

with no loss in performance. The substitution of high-strength 

grades of aluminum metal, pyrolytic graphite (an ultra-thin 

graphite film with a thermal conductivity up to four times 

greater than that of copper), silicon carbide, steel, or titanium 

metal for beryllium metal or composites, however, can result 

in substantially reduced performance.

A snapshot of uses of beryllium in the United States for 

the main classes of products by market share for 2011 are 

shown in figure E4. Performance alloys include the berylliumcopper and beryllium-aluminum alloys that consume most of 

the beryllium produced globally. Beryllium and composites 

include high-purity metals and metal-matrix composites 

and ceramics. Sales of performance alloys (fig. E4A) are 

about 600 percent greater than sales of the beryllium metal, 

composites, and ceramics sectors combined (fig. E4B). 

For beryllium metal and composites, defense and science 

applications accounted for approximately one-half of company 

sales in 2011, which was an 8 percent decrease compared 

with that of 2010; the decrease was thought to be owing to 

Government funding delays. Sales increased, however, for 

industrial component and commercial aerospace applications 

of both alloys and composites owing to an increase in the 

production of X-ray window assemblies. Sales for energy 

applications of performance alloys increased slightly in 

2011 compared with those of 2010 (Materion Corp., 2012a). 

Sales of ceramic products decreased slightly in 2011 after 

increasing by more than 50 percent in 2010 owing largely 

to increased shipments for applications within the telecommunications infrastructure market (Materion Corp., 2012a). 

Worldwide beryllium consumption is predicted to increase 

to 506 metric tons per year of contained beryllium by 2017 

owing to sustained consumption from the computer and 

telecommunications infrastructure markets and the growing 

automotive electronics market (Global Industry Analysts, Inc., 

2012). Developing markets in Asia and Latin America are also 

expected to increase consumption even though miniaturization 

in component manufacturing is requiring less beryllium than 

before for many uses.

Recycling efforts are important because of the 

huge energy savings that recycled beryllium brings to 

manufacturing. Manufacturing beryllium products from 

recycled sources requires only 20 percent of the energy 

required to manufacture beryllium products from ore. China, 

Kazakhstan, and the United States are the only countries that 

are currently capable of processing recycled scrap. Beryllium 

is recycled from new scrap generated during the manufacture 

of beryllium-containing components as well as from postconsumer scrap collected from end users worldwide. Owing 

to the long lifetime of many beryllium products, the technical 

difficulty of recycling beryllium (because of its use in 

complex alloys), and the low concentration of beryllium 

present in many consumer products, the percentage of scrap 

in the recycling flow was previously estimated to be between 

Table E1. Selected properties of beryllium, which is a Group 2 

alkaline earth metal.

[Source: Lide (2005, p. 14–39). Å, angstrom; Be, beryllium; BeH, beryllium 

hydride; °C, degree Celsius; eV, electronvolt; g/cm3, gram per cubic 

centimeter; GPa, gigapascal; J/mol K, joule per mole kelvin; K, kelvin; 

kJ/mol, kilojoule per mole; Mpa, megapascal; nΩ-m, nano ohm-meter; ppm, 

part per meter; μm/m K, micrometer per meter kelvin; W/m K, watt per 

meter kelvin]

Property Value

Atomic number 4

Atomic weight 9.012182

Isotopes 12 known radioisotopes; 9

Be 

is the only stable isotope; 

10Be has a half-life of

1.39 million years

Density at 293 K (g/cm3

) 1.8477

Melting point (°C) 1,287

Boiling point (°C) 2,469

Heat of fusion (kJ/mol) 12.2

Heat of vaporization (kJ/mol) 297

Molar heat capacity (J/mol K) 16.443

Hardness (Mohs scale) 5.5

Vickers hardness (Mpa) 1,670

Electrical resistivity at 20 °C (nΩ-m) 36

Thermal conductivity (W/m K) 200

Thermal expansion at 25 °C (μm/m K) 11.3

Young’s modulus (GPa) 287

Crystal structure Hexagonal

Magnetic ordering Diamagnetic

Electron configuration 1s2

2s2

Ionic radius (Å) 0.27

First ionization potential (eV) 9.32

Second ionization potential (eV) 18.21

Common valance states +2, rarely +1 

(BeH molecule)

Abundance, upper continental crust

 (ppm)

About 1.9 

(Hu and Gao, 2008)
E6 Critical Mineral Resources of the United States—Beryllium

A. Performance alloys B. Beryllium and composites

Industrial components 

and commercial

aerospace

32%

Consumer

electronics

20%

Automotive

electronics

14%

Energy

12%

Telecommunications

infrastructure

12%

Appliances, defense,

science, and medical

10%

Defense

and

science

52%

Industrial components 

and commercial

aerospace

26%

Medical

8%

Other

7% Telecommunications

infrastructure

7%

Figure E4. Pie charts showing reported end uses of beryllium consumed in the United States in 2011 for the two main classes of products—

A, performance alloys, and B, beryllium and composites (metals and ceramics). Usage is shown as a percentage of market share 

(Materion Corp., 2012a).

10 and 25 percent, based on data collected for 2000 through 

2005 (Graedel and others, 2011). Recent information from 

Materion, however, shows that an extensive worldwide 

recycling program for beryllium products is currently 

achieving a 40 percent recovery rate of beryllium from both 

of its new and postconsumer scrap sources (Jaskula, 2013a).

Geology

Geochemistry

Beryllium (table E1) is a member of Group 2 of the 

periodic table of elements; however, it has distinct differences in its crystal chemistry from other elements in the 

group, and these differences contribute to beryllium’s 

fundamentally unique attributes. Beryllium has 12 known 

isotopes and only 1, 9Be, is stable. Of the radiogenic isotopes 

of beryllium, the most stable are 10Be, which has a half-life 

of 1.39 million years, and 7Be, which has a half-life of about 

53 days; most of the others decay in milliseconds. Beryllium 

is the lightest of the alkaline earth elements, and its low 

atomic number (4) results in high first and second ionization 

potentials (9.32 and 18.21 electronvolts, respectively). 

As a consequence, beryllium can readily strip (or ionize) 

electrons from other uncharged atoms, and, owing to its 

small size (ionic radius of about 0.27 angstroms (Shannon, 

1976), Be 2+ has a strong tendency to polarize the electron 

clouds of the ligands to which it binds, forming bonds of 

shared electrons in compounds (Wuensch, 1972). These are 

strong and short chemical bonds that generate compounds 

displaying great hardness and high melting temperatures. 

Beryllium generally behaves as a hard acid (Pearson, 1963), 

preferentially forming complexes with hard ligands, such as 

carbonate, fluoride, and hydroxide. Because of its extremely 

small size, beryllium is not easily displaced by other elements 

in crystals and tends to form distinct minerals. Studies of the 

speciation and solubility of beryllium minerals at temperatures 

to 300 degrees Celsius (°C) for Be 2+ (its only known oxidation 

state in aqueous solution) suggest that beryllium is most likely 

to be transported in low-temperature hydrothermal fluid as 

a fluoride or fluoride-carbonate complex (Wood, 1992, and 

references therein). Over a wide range of temperatures and pH, 

the solubility of beryllium as a simple ion and its hydrolysis 

products are much less than 1 part per million (ppm) in 

solution. Thus, once formed, most beryllium minerals persist 

in surficial environments.

Beryllium is the 47th most abundant element found in 

Earth’s crust (Emsley, 2001), and it is widely distributed in 

many rocks types (Hörmann, 1978). Beryllium behaves as 

a lithophile element in that it combines readily with oxygen 

to form light beryllium oxide compounds, which leads to its 

enrichment in Earth’s crust rather than in the core. The concentration in primitive mantle is inferred to be about 60 parts per 

billion (ppb) beryllium (Taylor and McLennan, 1985, 1995), 

whereas estimates of the concentration of beryllium in the 

upper continental crust range from 1.9 to 3.1 ppm beryllium 

(table E1), which is an increase of more than fortyfold. 

This also accounts for the notable increases in beryllium 

concentrations in highly evolved volcanic and plutonic rocks, 

in granitic pegmatites, and in alkalic rocks compared with 

mantle compositions. In granite, the averages vary from 

5 ppm beryllium (in biotite granite) to 10 ppm beryllium (in 

muscovite and two-mica granite) (Hörmann, 1978; Taylor and 
Geology E7

McLennan, 1985, 1995). Nepheline syenite and fluorine-rich 

topaz-bearing rhyolites are even more enriched in beryllium 

than common felsic rocks. For example, alkalic rocks from 

Lovozero, Murmansk Oblast, Russia, contain beryllium in the 

range of about 3 to 24 ppm beryllium, and some fluorine-rich 

topaz-bearing volcanic rocks in the Western United States 

have average contents of 50 ppm beryllium (Hörmann, 

1978). Extreme enrichments are encountered in some unique 

rocks; for example, the Ilímaussaq complex in southern 

Greenland has peralkaline nepheline syenites that contain 

from 7 to 44 ppm beryllium —including some unusual rocks 

that have between 330 and 1,970 ppm beryllium (Engell 

and others, 1971) — and the Cornudas Mountains deposit in 

New Mexico, which has nepheline syenites that contain up to 

150 ppm beryllium (McLemore and Guilinger, 1993). Most 

beryllium-rich pegmatites have bulk beryllium contents of 

less than 250 to 420 ppm beryllium (London and Evensen, 

2002), although some albitic pegmatites have bulk contents 

in the range of from 360 to 720 ppm beryllium, which can be 

highly concentrated in specific zones within zoned granitic 

pegmatites (Černý, 2002). Hydrothermal processes can further 

concentrate beryllium in both pegmatite and volcanogenic 

systems. For example, altered rock (griesen) around a granite 

pluton at the Boomer Mine in Colorado contains up to 

2.3 percent beryllium, by weight, and volcanic tuff altered 

by heated circulating waters at Spor Mountain, Utah, may 

contain up to 3,600 ppm beryllium (Foley and others, 2012). 

Weathering of beryllium-enriched igneous rocks results in 

clay-bearing sedimentary rocks that serve as an important 

reservoir of beryllium in Earth’s crust. Metamorphism of such 

rocks as crustal plates subducted at convergent boundaries 

plays a critical role in the recycling of beryllium. Anatectic 

processes in the subduction zone cause partial melting of 

the crust, which releases the lithophile component in the 

rocks and generates beryllium-enriched magmas that are the 

source for new granitic rocks, thus completing the beryllium 

geochemical cycle.

Mineralogy

More than 100 distinct minerals that contain beryllium 

as an essential component of the mineral structure are known 

to occur in nature; these include arsenate, borate, hydroxide, 

oxide, phosphate, and silicate types. The relative stabilities of 

the major beryllium minerals (table E2) and their occurrence 

in diverse geologic environments are well established (Beus, 

1966; Barton and Young, 2002; Černý, 2002).

Beryl is the main beryllium ore mineral in plutonic 

systems and metamorphic environments, whereas bertrandite 

and phenakite are the predominant ore minerals in nearsurface volcanogenic and carbonate-hosted deposits. A number 

of other beryllium-bearing minerals, such as chrysoberyl and 

helvite, have reportedly been mined in small amounts from 

some deposits.

Beryl (fig. E2) is the primary beryllium silicate mineral 

that forms in rocks at high temperatures, and it has relatively 

widespread natural abundance (Barton, 1986; Barton and 

Young, 2002, and references therein). Beryllium saturation, 

and thus beryl solubility, varies as a function of the activities 

of beryllium, aluminum, and silicon. Beryl saturation can 

occur in granitic melts containing about 50 ppm beryllium 

at temperatures of 400 °C (London, 2005). Beryl has the 

ability to incorporate high and variable amounts of cesium, 

sodium, and water in large channels in its crystal structure 

and may show isomorphic substitutions of lithium for beryllium and magnesium, and iron and scandium for aluminum. 

These factors result in beryl being stable over a large range 

of temperature and pressure conditions and account for 

its presence in diverse geologic environments. The lower 

stability limit of beryl occurrence varies from 200 to 350 °C, 

depending on the compositions of coexisting minerals.

Beryl is the predominant beryllium phase in zoned 

granitic rare-element pegmatite environments where it occurs 

with other rare minerals of commercial importance, such 

as cesium, lithium, and tantalum, as well as clays, feldspar, 

muscovite, and high-purity quartz (London and Kontak, 

2012; Glover and others, 2012). In general, beryl is the only 

source of beryllium found in pegmatite deposits, although 

chrysoberyl is locally abundant in some pegmatites (table E2) 

(Linnen and others, 2012). Other occurrences of beryl include 

its presence in miarolitic cavities in volcanic rocks related to 

topaz-bearing igneous systems; for example, red beryl in the 

Wah Wah Mountains of Utah (Keith and others, 1994).

Bertrandite is the stable mineral in surface environments, 

as it forms at or below a temperature of 250 °C, and phenakite 

is stable at temperatures greater than 250 °C. Phenakite 

hydrates to bertrandite at lower temperatures (Barton, 1986). 

Bertrandite and phenakite are commonly found with beryl in 

carbonate-hosted skarn deposits and greisen associated with 

granite intrusions. Bertrandite is the predominant host for 

beryllium in volcanogenic deposits and occurs with phenakite 

in veins in carbonate replacement deposits. Precipitation in 

epithermal environments can result in cryptocrystalline and 

possibly colloidal bertrandite having variable amounts of 

molecular water that is readily lost on heating. The variety of 

bertrandite at Spor Mountain, Utah, reportedly has chemical 

and physical properties intermediate between colloidal 

bertrandite and bertrandite proper (Montoya and others, 1962).

The beryllium content in minerals in which it occurs 

as a substitute for another element, such as the important 

rock-forming minerals, is typically 10 ppm or less and rarely 

exceeds 100 ppm (London and Evensen, 2002). Several 

minerals have been reported to incorporate beryllium in 

higher amounts (that is, > 0.1 weight percent beryllium 

oxide, or 350 ppm beryllium). These include clay minerals 

of the smectite group that reportedly can contain up to 

2.8 weight percent beryllium oxide as either crystallographic 

substitution or ionic adsorption on clay surfaces (Montoya 

and others, 1962).
E8 Critical Mineral Resources of the United States—Beryllium

Table E2. Selected beryllium minerals.

[Sources: Barton and Young (2002), Grew (2002), Foley and others (2010), and Jaskula (2013b). BeO, beryllium oxide; —, limited or no commercial use]

Mineral name Chemical formula BeO

(percent) Use Deposit model,occurrence type

Barylite BaBe2

Si2O7 16 — Alkalic pegmatite, skarn, 

greisen

Bazzite Be3

(Sc,Al)2Si6O18 13.29 — Alkalic and granitic 

pegmatite, vein

Behoite Be(OH)2 58.13 — Alkalic pegmatite, skarn, 

rhyolite

Bertrandite Be4

Si2O7(OH)2 42 Industrial Greisen, skarn, vein, rhyolite

Beryl Be3

Al2Si6O18 13.96 Industrial, gem Granitic pegmatite, greisen, 

skarn, vein, rhyolite

Beryl, variety emerald Be3

(Al,Cr,V)2Si6O18 13.96 Gem Granitic pegmatite and 

metamorphosed 

equivalents, vein

Beryl, variety aquamarine Be3

(Fe,Al)2Si6O18 13.96 Gem Granitic pegmatite and 

metamorphosed 

equivalents, vein

Beryllite Be3

SiO4(OH)2•(H2O) 43.84 Industrial? Rocks of the Ilimaussaq 

intrusion

Chrysoberyl BeAl2

O4 19.7 Industrial, gem Granitic pegmatite, skarn, 

vein

Danalite Fe4

Be3Si3O12S 13.43 — Skarn, granitic pegmatite

Epididymite Na2

Be2Si6O15•H2O 10.2 — Alkalic pegmatite, skarn

Euclase BeAlSiO4

(OH) 17.24 Gem Granitic pegmatite, greisen, 

skarn, vein

Gadolinite-(Y),-(Ce) Be2

Fe(Y,REE)2Si2O10 10.69 — Alkaline pegmatite and 

granite, vein, greisen

Hambergite Be2

(OH,F)BO3 53.31 — Alkaline and granitic 

pegmatite

Helvite Mn4

Be3Si3O12S 13.52 Industrial Veins, skarn, greisen, alkaline 

and granitic pegmatite

Herderite CaBePO4

(F,OH) 15.34 — Greisen, granitic pegmatite

Hurlbutite CaBe2

(PO4)2 20.17 — Granitic pegmatite, vein

Leucophanite CaNaBeSi2

O6F 10.6 — Alkaline pegmatite, skarn

Meliphanite Ca4

(Na,Ca)4Be4 AlSi 7O24(F,O)4 10.37 — Alkaline pegmatite, skarn

Milarite K(H2

O,Na)2(Ca,Y,REE)2

(Be,Al)3Si12O30

5.05 — Alkaline and granitic 

pegmatite, skarn, vein

Phenakite Be2

SiO4 45.43 Industrial, gem Alkaline and granitic 

pegmatite, skarn, 

greisen, vein

Tugtupite Na4

AlBeSi4O12Cl 5.35 Gem, industrial? Rocks of the Ilimaussaq 

intrusion
Geology E9

Deposit Types

Magmatic-Related Beryllium Deposits

Beryllium deposits are most commonly associated 

with geologic environments that show a clear link to 

magmatic processes (Barton and Young, 2002; Černý, 2002). 

Magmatic-related beryllium deposits can be characterized in 

terms of (a) the petrologic association of the igneous rocks, 

(b) the type of associated country rock, and (c) the nature of 

the depositional environment (table E3).

Magmatic-related beryllium deposits and occurrences 

are found in two general rock types—alkaline to peralkaline 

and metaluminous to peraluminous—which are distinguished 

on the basis of whole-rock chemistry. The deposit types can 

be divided into those associated with rare-metal enriched 

felsic intrusions and related volcanic sequences and pegmatite 

bodies, and those that form where rare-metal enriched magmas 

interact with adjacent country rocks. Thus, rare-metal plutons 

and related pegmatite deposits containing beryl; skarn and 

greisen deposits containing bertrandite, beryl, and phenakite; 

and volcanogenic and carbonate-hosted deposits of mainly 

bertrandite are the types associated with large-scale 

magmatic systems (for example, fig. E5). Of these, lithiumrich pegmatites that contain beryl in high-grade pockets 

(Bradley and McCauley, 2013) and large-tonnage, high-grade 

volcanogenic beryllium deposits (Foley and others, 2012) 

are economically important at present (table E3, shown in 

shading; modified from Barton and Young, 2002).

Rare-Metal Pegmatite Deposits—Beryllium Type

Many plutons of gabbro, granite, and syenite composition contain minor amounts of beryl; however, in general, 

only rare-metal granitic pegmatites contain beryl in sufficient 

quantities to mine. Although granitic pegmatites are widespread and relatively common, rare-metal pegmatites make 

up only about 0.1 percent of the total fraction (Laznicka, 

2006). Rare-metal pegmatites (table E3; fig. E5) that 

contain relatively abundant beryl include those of the rare 

element class, lithium subclass, beryl type (LCT family) 

and those of the miarolitic class, REE subclass, topaz-beryl 

type (niobium-yttrium-fluorine [NYF] family), and of the 

miarolitic class, lithium subclass, beryl-topaz type (LCT 

family) (Černý and others, 2012, table 1). Of these, the most 

important sources of industrial beryl are granitic pegmatites 

of the rare-element LCT family. Miarolitic LCT-family 

and NYF-family pegmatites (fig. E6) are mainly sources of 

gemstone beryl. Industrial beryl is also produced from more 

complex types of pegmatite; for example, Anjanabonoina 

pegmatites in Antananarivo Province, Madagascar, which 

have a mixed NYF–LCT signature (Martin and De Vito, 

2005). Many pegmatite districts that are of interest now for 

lithium and niobium originally attracted attention for other 

rare metals, especially beryllium, tantalum, and tin, and 

industrial minerals, such as feldspar, micas, and high-purity 

silica (Glover and others, 2012; Kesler and others, 2012).

Most granitic pegmatites of the LCT family show 

peraluminous compositions (table E3; fig. E5) that are 

characteristic of S (or sedimentary)-type granites, which form 

from anatexis, or melting, of metamorphosed pelitic rocks 

owing to the crustal thickening that is typically associated with 

subduction and continental collision (Černý and others, 2012). 

LCT pegmatites have been linked to subduction zones because 

of the high abundance of granitic pegmatites in orogenic 

belts (Martin and De Vito, 2005), although others affiliate 

them with a post-tectonic phase of evolution in continentalcollision belts (Tkachev, 2011). Nonetheless, the geochemical 

character of LCT pegmatites depends on their derivation from 

previously unmelted, mica-rich metamorphic rocks because 

muscovite-paragonite-phengite and biotite group minerals 

carry the distinguishing rare elements (Černý and others, 

2012). The presence of diverse fluxing components, such as 

boron, flourine, and phosphorus, in addition to water, in the 

metamorphosed juvenile sediments that make up the peraluminous sedimentary-type source materials are thought to have an 

essential role in the formation of LCT pegmatites because they 

mobilize the rare metals, including Be, Cs, Nb, Rb, Sn, and Ta, 

into the fluid phase (Linnen and others, 2012).

Rare-metal pegmatites typically occur as halos that 

encircle large leucocratic plutons or as swarms that extend 

into country rock adjacent to these intrusions. The best quality 

deposits are found in zoned pegmatites where beryl is mined 

from certain, typically highly fractionated, pegmatite zones; 

from massive vein and fracture fillings; and from replacement 

bodies in complexly heterogeneous pegmatites. Important 

pegmatite districts for beryl mining (fig. E1) occur in Brazil 

(in the Rio Doce Valley and adjacent areas in the State of 

Minas Gerais; the city of Campina Grande in the State of Rio 

Grande do Norte; Cristais-Berilandia in the State of Ceara; and 

Rio Jequitinhonha [Jequitinhonha River] in the State of Minas 

Gerais); Portugal (in the Covas do Barroso pegmatite district, 

Viseu); Mozambique (at Alto Ligonhain in the Province of 

Zambezia); Canada (at the Bernic Lake Mine — commonly 

known as the Tanco Mine — on the shore of Bernic Lake in 

the Province of Manitoba; and at the James Bay (Lithium 

One) Mine on the shore of James Bay in the Province of 

Quebec); Namibia (at Karibib); Zimbabwe (at the Bikita Field 

in the Province of Masvingo); and China (Kesler and others, 

2012). Pegmatite districts located in weathered granites 

of southeastern China—for example, the city of Nanping 

in Fujian Province—may supply some of the reported 

Chinese production, although little information is available 

on resource types for beryllium from China. Subeconomic 

occurrences in the United States that were mined historically 

include pegmatite bodies of North Carolina (at the Kings 

Mountain crushed rock quarry near Bessemer City), South 

Dakota (at Tin Mountain Mine in the Black Hills; Sabey, 

2006), and New Hampshire. For example, pegmatites at the 

Palermo Mine on Bald Mountain near North Groton, New 

Hampshire, and at Beryl Mountain near South Acworth, 

New Hampshire, produced much of the beryllium for the 

U.S. Manhattan Project.
E10 Critical Mineral Resources of the United States—Beryllium

Table E3. Global and domestic types of magmatic-related beryllium resources.

[Modified from Barton and Young (2002). Shading indicates information for deposits that are economically important at present. For each petrologic association listed, the following is included: 1, frequency of known occurrences (abundant, common, uncommon, rare) and end use for which the beryllium is most 

valuable (gems, industrial beryllium); 2, the main beryllium-bearing phase(s); and 3, one or more example localities. “Transbaikalia” is the conventional spelling of the “Zabaykal’ye” region. LCT, lithium-cesium-tantalum; NYF, niobium-yttrium-fluorine. Locality abbreviations: Colo., Colorado; Ill., Illinois; N.C., 

North Carolina; N.H., New Hampshire; N. Mex., New Mexico; Nev., Nevada; Tex., Texas; UK, United Kingdom; Va., Virginia; Mt., Mount; Mtn., mountain; 

—, not applicable] 

Petrologic

association

Host rock (type of mineralization)

Igneous

(pegmatitic, late-stage

plutonic, volcanic glasses)

Aluminosilicate

(greisen, veins)

Carbonate

(skarn, replacement)

Mafic/ultramafic

(blackwall, vein)

Metaluminous to peraluminous

Granite

1 Common?/Industrial-beryllium 

resource?

Abundant/gems ± industrialberyllium resource?Abundant/industrialberyllium resourceCommon/gems, 

emerald

2 Lithium micas, beryl Beryl, phenakite Phenakite, chrysoberyl, 

bertrandite, helvite 

Beryl, chrysoberyl,

phenakite

3 Beauvoir, France; 

Sheeprock, Utah

Sherlova Gora, Russia;

Mt. Antero, Colo.; 

Aqshatau, Kazakhstan;

Erzgebirge, Germany; 

Cornwall, UK; Boomer 

Mine, Pikes Peak, Colo.

Lost River, Alaska;

Mt. Wheeler, Nev.; 

Mt. Bischoff, Australia; 

McCullough Butte, 

Nev.

Reft River, Russia;

Khaltro, Pakistan;

Carnaiba, Brazil

Granitic 

pegmatite 

(LCT-type)

1 Abundant/gems and major 

industrial-beryllium resource

— — —

2 Beryl

3 Minas Gerais, Brazil; 

Bernic Lake, Canada; 

Transbaikalia, Russia; 

Greenbushes, Australia; 

Bikita, Zimbabwe;

Donkethoek, Namibia;

Sahatany Valley, Madagascar; 

King’s Mountain, N.C.

Granitic 

pegmatite 

(NYF-type)

1 Abundant/gems and major 

industrial-beryllium resource

— — —

2 Beryl

3 Mt. Antero, Colo.; Ledu, Quebec; 

Transbaikalia, Russia; 

Klein Spitzkoppe, Namibia; 

Amherst Co., Va.; 

Sawtooth batholith, Idaho; 

Kitsamby, Madagascar

Rhyolite

1 Common/no economic value? Rare/gems, red beryl Uncommon/principal 

industrial-beryllium 

resource

2 Beryllium in glass or micas Beryl Bertrandite —

3 Macusani, Peru; Spor Mtn., Utah; 

topaz rhyolites, Western United 

States; Warm Springs, N. Mex.

Wah Wah Mtns., Utah; 

Black Range, N. Mex.; 

Spor Mtn., Utah

Spor Mtn., Utah;

Aguachile, Mexico;

Sierra Blanca, Tex.
Geology E11

Table E3. Global and domestic types of magmatic-related beryllium resources.—Continued

[Modified from Barton and Young (2002). Shading indicates information for deposits that are economically important at present. For each petrologic association listed, the following is included: 1, frequency of known occurrences (abundant, common, uncommon, rare) and end use for which the beryllium is most 

valuable (gems, industrial beryllium); 2, the main beryllium-bearing phase(s); and 3, one or more example localities. “Transbaikalia” is the conventional spelling of the “Zabaykal’ye” region. LCT, lithium-cesium-tantalum; NYF, niobium-yttrium-fluorine. Locality abbreviations: Colo., Colorado; Ill., Illinois; N.C., 

North Carolina; N.H., New Hampshire; N. Mex., New Mexico; Nev., Nevada; Tex., Texas; UK, United Kingdom; Va., Virginia; Mt., Mount; Mtn., mountain; 

—, not applicable] 

Petrologic

association

Host rock (type of mineralization)

Igneous

(pegmatitic, late-stage

plutonic, volcanic glasses)

Aluminosilicate

(greisen, veins)

Carbonate

(skarn, replacement)

Mafic/ultramafic

(blackwall, vein)

Alkaline to peralkaline

Granite

1 Rare?/no economic value? Rare/industrial-beryllium 

resource

Rare/industrial-berylliumresource

2 — Gadolinite group? Phenakite, helvite, 

gadolinite

Phenakite, bertrandite, 

leucophanite

3 Khaldzan-Burgtey, Mongolia Verknee Espee, Kazakhstan Yermakovskoye, Russia; 

Iron Mtn., N. Mex.

Granitic 

pegmatite 

(NYF-type)

1 Uncommon/industrial-beryllium 

resource

— — —

2 Gadolinite, phenakite

3 Strange Lake, Canada; 

Pikes Peak, Colo.; 

Sawtooth batholith, Idaho;

South Baldface Mtn., N.H.

Granitic 

pegmatite 

(LCT-type)

1 Uncommon/industrial-beryllium 

resource

— — — 2 Beryl

3 Ivisaartoq, Greenland

Rhyolite

1 Rare?/industrial-beryllium

resource?

—

Rare/industrial-berylliumresource

2 — Bertrandite, beryllium in glass, 

feldspar, micas, helvite

Bertrandite

3 Brockman, Australia; 

Shixi, China

Aguachile, Coahuila, 

Mexico

Syenitic

pegmatite 

(both LCT 

and NYTtypes?)

1 Common/no economic value?

— — —

2 Epididymite, eudidymite,

chkalovite

3 Lovozera, Russia; 

Ilímaussaq, Greenland; 

Oslo, Norway

Syenite

1 Uncommon/industrial-beryllium 

resource?

Rare/industrial-beryllium 

resource

Rare/industrial-beryllium 

resource

— 2 Chkalovite, epididymite Barylite, eudidymite Bertrandite

3 Ilímaussaq, Greenland;

Wind Mtn., N. Mex.

Seal Lake, Canada; 

Thor Lake, Canada

Hicks Dome, Ill.;

Thor Lake, Canada 

(skarn)
E12 Critical Mineral Resources of the United States—Beryllium

Beryllium environments associated with metaluminous and weakly peraluminous systems

Be-enriched magma

(Thomas Range, Utah)

Replacement Be minerals in

carbonate (Sierra Blanca, Texas)

Replacement Be minerals in

tuff (Spor Mountain, Utah)

Greisen Sn (-Ta-Nb-Zn)

(Younger Granites, 

Nigeria and Niger)

Red beryl in gas cavities and cooling joints 

(Wah Wah, Utah; Black Range, New Mexico)

Miarolitic pegmatites and 

greisen quartz veins

(Mt. Antero, Colorado; 

Sheeprock, Utah)

Fe-rich skarns and replacements ± porphyrystyle Mo (-W) (Iron Mountain, New Mexico; 

Oslo Graben, Norway; Ermakovka, Siberia) 

Rhyolite lavas and pyroclastic

tuff deposits

Subvolcanic intrusions, 

domes, caldera structures

Granites, syenites, porphyrystyle intrusives

Paleozoic sedimentary sequences, 

platform carbonates 

Magmatic enrichments (dispersed in minerals or volcanic glass)

Miarolitic and pegmatitic zones (Alpine-type clefts or pegmatites 

in metamorphic rocks)

High-temperature (>300 °C) veins, individual, sheeted 

and stockwork

Low-temperature (<300 °C) veins, mostly individual or sheeted

Skarns, typically F-rich with variable Al-Fe-metal contents

Carbonate replacement deposits, typically F-rich

Rock types Mineralization styles

EXPLANATION

Beryllium Figure 5.

NOT TO SCALE

Figure E5. A generalized cross section showing the geologic setting of and some example deposits for the major 

types of beryllium resources associated with rare-metal magma systems. The cross section also shows the relation of 

volcanogenic beryllium deposits, such as the one at Spor Mountain, Utah, to other types of deposits that are associated 

with magmas of this general composition. Modified from Barton and Young (2002). °C, degree Celsius. Elements: Al, 

aluminum; Be, beryllium; F, fluorine; Fe, iron; Mo, molybdenum; Nb, niobium; Sn, tin; Ta, tantalum; W, tungsten; Zn, zinc
Geology E13

Beryllium Figure 6

BRITISH 

COLUMBIA

YUKON

ALASKA

UNITED STATES

Rare-metal plutons

EXPLANATION

Rare-metal deposits 

 and occurrences

Lost River

Kougarok

Kalzas

Why are some rare metal plutons labeled and not all?

Seagull

batholith

Surprise Lake 

 batholith

Parlet 

Creek 

pluton

Hellroaring Creek prospect

Area of map B

PACIFIC

OCEAN

A B

Be

Be

Be

Be Be

Be

Be

Be

Be

Be

Be

Be,Ta, Nb

CANADA

Be

Granite 

Pegmatite

Moyie sill (gabbro)

Creston and Aldridge Formations

Fault

Mineral occurrence

EXPLANATION

2500 500 METERS

0 1,000 FEET

N

Figure E6. Location map and simplified geologic map of the Hellroaring Creek prospect, which is located west of Kimberley, 

British Columbia, Canada. Location map A shows the distribution of potential rare-metal-bearing granite and pegmatite deposits 

of the Canadian and Alaskan Cordillera and of rare-metal deposits associated with the plutons. Geologic map B shows the 

pegmatite and granite-related beryllium-bearing rare-metal mineralization at the Hellroaring Creek prospect. Maps modified 

from Soloviev (2011). Be, beryllium; Nb, niobium; Ta, tantalum

Most pegmatite deposits consist of a few thousand 

tons of rock, although the largest contain as much as 

1 million metric tons. For example, the Tanco deposit in 

Manitoba, Canada, is a large, mineralogically complex, and 

strongly zoned LCT pegmatite that contains an estimated 

900,000 metric tons of beryllium ore. The Tanco pegmatite is 

part of the rare-metal-bearing Bernic Lake pegmatite group 

(Černý, 2005) and is currently being mined for cesium, 

rubidium, and spodumene (a lithium-bearing mineral). Ores of 

the Tanco pegmatite contain greater than 350 ppm beryllium, 

which is an enrichment factor of 180 over bulk continental 

crust, in addition to 13,900 ppm lithium, 236,000 ppm 

cesium, 28,900 ppm rubidium, and greater than 1,200 ppm 

tantalum (Stilling and others, 2006). Beryl at Tanco and in 

many other large complex pegmatite deposits occurs both as 

fine-grained rinds and as crystal up to 20 centimeters (cm) 

in length in pockets in albitic aplite, which is a fine-grained 

granitic rock. In contrast, the Hellroaring Creek prospect, 

which is located west of the city of Kimberley in the Province 

of British Columbia, Canada (fig. E6A), is a chemically 

and mineralogically simple lithium-pegmatite body that has 

indicated resources of 500,000 metric tons grading 0.1 percent 

beryllium oxide (Soloviev, 2011). In this case, the pegmatite 

intrudes an overlying granitic stock, and pockets of beryl 

are distributed in both granite and pegmatite (fig. E6B). The 

pockets have a relatively simple mineralogy of beryl (with 

crystals up to greater than 30 cm in length), muscovite, 

and quartz. The prospect is located in an extensive trend of 

rare-metal granites, pegmatites, and related deposits of the 

Cordilleras of North America (fig. E6A). These range from 

small and uneconomic beryl pegmatite deposits of Idaho and 

Montana (Stoll, 1950) to significant beryllium deposits 

of Seward Peninsula, Alaska (for example, the Lost River 

skarn deposit).

Skarn and Greisen-Hosted Beryllium-Fluorine Deposits

Beryllium-bearing skarn and greisen deposits occur 

in association with many large rare-metal-enriched granitic 

and syenitic plutons. These deposit types, which form by 

metasomatic replacement processes in the mineralized 

contact zone between the intruding pluton and the country 

rock, typically have phenakite, bertrandite, and helvite-group 

minerals. Skarns form by high-temperature alteration of 
E14 Critical Mineral Resources of the United States—Beryllium

carbonate country rocks, and greisens form by alteration of 

rocks of alumino-silicate compositions. They often display 

a distinctive texture of rhythmically banded replacements 

containing alternating light and dark layers of beryllium 

minerals, fluorite, magnetite, and silicates. Fluorine-beryllium 

skarn and greisen deposits associated with, for example, 

tin deposits of the Seward Peninsula, Alaska, such as the 

Lost River beryllium skarn deposit (Alexsandrov, 2010), and 

those in Russia’s Far East, such as the Yermakovskoye deposit 

in the Kizhinginskiy Khrebet [Kizhinginskiy Ridge] area of 

the Republic of Buryatia, are localized in the contact aureoles 

of carbonate rocks intruded by high-silicon (leucocratic), 

fluorine-rich granites (Reyf, 2008). The beryllium is contained 

in veins and vugs in fluorite and calc-silicate altered limestone 

and argillized granite. Beryl, helvite-group minerals, and 

phenakite occur along with calcite, fluorite, garnet, quartz, 

topaz, tourmaline, zinnwaldite (a lithium-mica), and sulfide 

minerals. Skarn deposits near Changning City in Hunan 

Province, China (Lin and others, 1995), may supply some of 

the beryl currently mined in China. Beryllium-fluorine skarn 

and greisen deposits of this type are deep, high-pressure 

analogs of carbonate-hosted beryllium deposits, and, where 

present, may represent the roots of volcanogenic beryllium 

systems (fig. E6).

Carbonate-Hosted Beryllium Deposits

Carbonate-hosted beryllium deposits form in carbonate 

rock sequences where rare-metal and lithophile-enriched 

volcanic rocks intrude and mineralize carbonate rocks, 

depositing beryllium minerals in replacement fronts (fig. E6). 

These types of beryllium deposits occur near Sierra Blanca, 

Texas, and Aguachile, Mexico. At Sierra Blanca, beryllium-rich 

fluorite deposits occur in association with peralkalinemetaluminous volcanic rock sequences. The deposits occur 

within Cretaceous limestones beneath an intrusive contact with 

an overlying rhyolite laccolith. The rhyolite laccoliths of Round 

Top Mountain near Sierra Blanca are strongly peraluminous 

and have anomalous beryllium and fluorine contents (Rubin 

and others, 1987, 1988, 1990). The rhyolites also contain 

elevated amounts of Li, Nb, Rb, rare-earth elements (REEs), 

Th, Y, Zn, and cryolite. The beryllium mineralization at 

Aguachile occurs as replacements and fillings in limestone 

that was brecciated when it was intruded by a rhyolite dike 

near the ring fault of a caldera; the rhyolite is peraluminous, 

although related igneous rocks are peralkaline (McLemore, 

2010a). Bertrandite is present with calcite and quartz in 

fluorite-rich deposits in limestone, and the principal orebody 

has approximately 0.3 percent beryllium oxide.

At the present time, no domestic carbonate-hosted 

beryllium deposits are known to be economic; however, 

they can be important indicators of the potential presence of 

volcanogenic beryllium deposits. For example, a number of 

small and irregularly shaped beryllium-rich zones are present 

in carbonate rocks located stratigraphically beneath the bedded 

tuff that hosts world-class volcanogenic beryllium ores at 

Spor Mountain (McLemore, 2010a; Materion Corp., 2012a, b).

Volcanogenic-Hosted Beryllium Deposits

Beryllium deposits of the volcanogenic type (Barton and 

Young, 2002; Foley and others, 2012, and references therein) 

form by replacement of carbonate fragments found in lithic 

tuff, which is a rock composed of compacted volcanic ash, 

glass, and rock fragments. Host rocks are typically rare-metalenriched volcanic rhyolite lavas and related tuffs and ash 

flows that are metaluminous to peraluminous in composition 

and enriched in fluorine and other lithophile-elements (Be, 

Ce, Li, Rb, REEs, Sn, Th, Tl, and U) (fig. E7A). The classic 

tectonic setting is an extensional environment where young 

igneous rock sequences overlie older dolomite, limestone, 

quartzite, and shale sequences. The deposits form as the result 

of magmatic processes that produce fluorine-, beryllium-, 

and gas-rich rhyolite melts. Degassing of the melts causes 

explosive volcanism that rips up carbonate and another rock 

fragments as magmas vent to the surface through fractures 

in the limestone sequences. The hot magma interacts with 

groundwater, which causes a hydrovolcanic eruption that 

dumps fine ash, beryllium-rich volcanic glasses, and fragments 

of carbonate and other rock in wedged-shaped base surge 

deposits that thin away from the vent structures (fig. E7B; Burt 

and others, 1982). This mix of magmatic and hydrothermal 

fluids is capable of leaching beryllium from the volcanic glass 

and precipitating it as bertrandite when the hydrothermal 

fluid encounters carbonate fragments in the tuff. Beryllium 

is removed from devitrifying glass as a fluoride complex in 

the cooling fluid. Upon encountering carbonate rock, calcium 

from the rock fragment and fluorine from the fluid combine 

to form the mineral fluorite, and the beryllium reacts with 

silica in the fluid to make bertrandite; this leads to essentially 

simultaneous deposition of fluorite and bertrandite. These 

post-magmatic processes play an important role in the development of economic volcanogenic beryllium deposits because 

they concentrate the beryllium into high-grade ore zones. The 

size and grade of a volcanogenic beryllium deposit is thought 

to be related to the distribution of carbonate lithic-rich zones 

in the host beryllium tuff and the duration and efficiency of the 

mineralizing hydrothermal system.

Volcanogenic ores of beryllium are produced at Spor 

Mountain, Utah, from elongate deposits in Miocene-aged 

tuffs composed of glass, ash, carbonate fragments, and 

other rock fragments (fig. E7B). The bertrandite occurs as 

submicroscopic grains mixed with fine-grained fluorite, opal, 

and calcite in layered nodules that replace carbonate in the 

tuff (fig. E7C; Lindsey, 1998). The median concentration of 

unaltered glass melt (59 ppm beryllium) in Spor Mountain 

tuff (Adams and others, 2009) is about eight times that left 

in intensely altered glass (7 ppm beryllium), which suggests 

that almost 90 percent of the original beryllium in the glassy 

tuff matrix was redeposited in the mineralized carbonate 

rock fragments.

Only a small number of volcanic rocks and beryllium 

deposits of this type are known, and most are subeconomic 

(table E3; fig. E1). These include domestic rare-metal-bearing 

volcanic occurrences at The Honeycombs (more commonly 
Geology E15

Roadside

Monitor

Rainbow

Blue Chalk

Hogsback 

South prospect

Wind

Taurus

Sigma

Emma

Fluro

Camp

Fluorite pipes 

 and veins

Fluorite pipes 

at depth? Rhyolite

Vent, tuff

C D

A

B

Rhyolite dome, cap rock

Uranium, in tuffaceous 

sandstone and conglomerate, 

disseminated and in channels

Beryllium ore, in 

lithic-rich tuff

Paleozoic rocks

Older volcanics

Rhyolite unit, porphyritic 

Beryllium tuff unit, vitric 

 and lithic-rich

EXPLANATION

Berylllium Figure 7.

Volcanic rocks and alluvium,

 west of caldera margin

Beryllium deposit, surface 

 projection of bertrandite-

 fluorite-silica deposits 

 in Miocene lithic tuffs 

Fault

Relative movement

Rhyolite vent

Fluorite pipe

Fluorite vein

Uranium deposit

NOT TO SCALE

0 0.5 1 KILOMETER

0 0.25 0.5 MILE

UTAH

Spor Mountain, 

Joab County 

N

Figure E7. A, Generalized geologic cross section showing the setting of volcanogenic beryllium deposits and related deposit 

types relative to volcanic vent structures, host tuff, and overlying cap rocks. B, Gelogic map of Spor Mountain area, Utah, showing 

the distribution of beryllium deposits, fluorite pipes, carbonate rocks, and volcanic rocks. Modified from Lindsey (1998) and Foley 

and others (2012). C, Photograph of beryllium tuff at Spor Mountain showing dolomite clasts (white) replaced by fluorite, clay, 

chalcedonic quartz, and manganese oxide. D, Photograph of nodule from the Spor Mountain tuff showing an interior of opal and 

purple fluorite and an outer zone of opal and fluorite; nodules can contain as much as a few percent beryllium oxide (BeO) as 

bertrandite. The nodule shown is about 20 centimeters wide. Photographs C and D are by D.A. Lindsey, U.S. Geological Survey.
E16 Critical Mineral Resources of the United States—Beryllium

known as Honeycomb Hills), Utah, and at the Apache Warm 

Springs volcanogenic beryllium deposit at the Ojo Caliente 

No. 2 mining district, New Mexico. The Honeycombs are 

two small eroded domes of topaz-bearing alkali rhyolite and 

tuff located about 30 kilometers (km) (20 miles) west of Spor 

Mountain; the prospect contains small amounts of beryllium 

and fluorite. Similar prospects throughout New Mexico and 

Utah are also small and of low grade (McLemore, 2010a). 

Globally, rare-metal metallogenic provinces favorable for 

volcanogenic beryllium deposits are restricted to a small 

number of locations, including the peralkaline Niobium tuff of 

the State of Western Australia, which contains the Brockman 

deposit (table E3) (Ramsden and others, 1993; Taylor and 

others, 1995a, b); the Macusani tuff of the Peruvian Altiplano 

(Pichavant and others, 1988a, b); and peraluminous to 

peralkaline volcanic rocks of the Teg-Ula deposit in GoviAltay Province, Mongolia (Kovalenko and Yarmolyuk, 1995), 

and in Shixi, Zhejiang Province, China (Lin, 1985). At the 

Brockman deposit, beryllium occurs in low-grade lenses of 

Nb-REE-Ta-Zr mineralization associated with the pyroclastic 

eruption of a Paleoproterozoic magma enriched in volatiles. 

The tuff hosts rare-metal deposits with proven reserves of 

about 4.3 million metric tons containing 0.44 percent Nb2O5

and 50 to 1,500 ppm beryllium. Beryllium-bearing ash-flow 

tuffs also occur in the vicinity of the town of Macusani in 

the Puno Region of southeastern Peru (Pichavant and others, 

1988a, b). The highly peraluminous rocks that compose the 

Neogene volcanic field near Muscani may be a northernmost 

expression of the Bolivian tin belt. The tuffs contain elevated 

concentrations of beryllium (about 15 to 37 ppm) but no 

known beryllium deposits. Deposits and occurrences of 

beryllium in Mongolia and Siberia, Russia, reportedly are 

hosted by fluorine-rich igneous suites, such as those typically 

associated with volcanogenic beryllium deposits. The deposits 

are mostly associated with granite and are of the skarn or 

greisen type. In Mongolia, however, the central Gobi volcanic 

zone contains fluorine-bearing rhyolite, beryllium-rich tuff, 

and REE-rich alkaline volcano-plutonic rocks that compose 

part of a late Mesozoic rare-metal metallogenic belt within the 

Transbaikal-Mongolian rare-metal province (Kovalenko and 

Yarmolyuk, 1995).

Nonmagmatic-Related Beryllium Deposits

Beryllium minerals also occur in a variety of nonmagmatic hydrothermal, metamorphic, and sedimentary environments; in most cases, these occurrences are subeconomic. 

Shear-zone deposits form in metamorphic terranes by local 

redistribution of materials during deformation. Metamorphic 

host rocks may include felsic metavolcanics, metasedimentary 

rocks, marbles and metabasites, and, thus, shear-zone 

deposits are found in blackwall (mafic rock), greisen, skarn, 

and vein settings. They are uncommon occurrences of beryl 

and phenakite and are typically mined as gem emerald 

(Grundmann and Morteani, 1989). Basinal brine deposits 

consist of greisens and veins formed by replacement of 

aluminosilicate host rock by low-temperature basinal brines 

with unsubstantiated or unknown connections to magmatic 

systems. Classic examples include the emerald deposits of 

Muzo and Chivor, in the Department of Boyacá, Colombia, 

where beryl and euclase are exclusively mined as gemstones. 

Placer deposits containing beryl are common and historically 

have provided both industrial beryllium ore and gemstones 

(Sinkankas, 1981). Deposits of this type are known in the State 

of Minas Gerais, Brazil, and in Madagascar and Sri Lanka. 

Nonmagmatic deposits can be a source of industrial beryl; 

however, most are mined as gemstones of beryl-type minerals, 

such as aquamarine, chrysoberyl (variety alexandrite), 

and emerald, as well as euclase, phenakite, and tugtupite 

(table E2). Gemstones are a distinct mineral commodity and 

are not covered in this chapter, but they are covered elsewhere 

(see, for example, Olson, 2016).

Resources and Production

Identified resources of beryllium worldwide have been 

estimated to be more than 80,000 metric tons (Jaskula, 2013a). 

About 65 percent of these are in U.S. resources of bertrandite, 

and the rest are in resources of beryl in other countries. 

Beryllium grades and resource tonnages for volcanogenic 

beryllium deposits, LCT pegmatites, and selected other 

deposit types are shown for the purpose of general comparison 

in figure E8. World beryllium reserves are not sufficiently well 

delineated to report consistent figures for all countries.

Domestic Resources

Domestic carbonate-hosted and volcanogenic deposits of 

beryllium occur in New Mexico, Texas, and Utah. The deposits 

range in size from small (Iron Mountain in New Mexico) to 

world class (Spor Mountain in Utah) (table E4; fig. E8).

Exploration sites thought to host the large deposits are 

described below. It is unlikely that other domestic beryllium deposits will be mined in the near future because the 

Spor Mountain deposit contains sufficient beryllium reserves 

to meet expected demand.

Spor Mountain, Utah.—At yearend 2011, Materion Corp. 

(2011) reported proven domestic reserves of bertrandite 

at Spor Mountain of 5.75 million dry tons at an average 

grade of 0.265 percent beryllium (about 0.8 percent beryllium oxide). The bertrandite contains about 15,200 metric 

tons of beryllium. Figure E8 shows ore grade and tonnage 

estimates for Spor Mountain based on a cutoff grade of 

0.23 percent beryllium.

Lost River, Alaska.—The Lost River Mines, which are 

located in the western part of Seward Peninsula, were mined 

intermittently for tin from 1904 through 1955. Near the mines, 

a limestone of Early Ordovician age is partly replaced by 

silicate minerals and cut by numerous veinlets containing 

silicates, ore minerals, and fluorite. A granite pluton and dikes 
Resources and Production E17

Beryllium Fig 9

Apache Warm Springs, NM 

Spor Mountain, UT 

Aguachile, Mexico

Lost River, AK

Tanco, Canada

Brockman, Australia

Boomer, CO

McCullough

Butte, NV

Thor Lake, Canada

Strange

Lake, 

Canada

Atlantic 

Shield,

Brazil 

Ilimaussaq,

Greenland

Black Hills, SD

Aqshatau,

Kazakhastan

Hellroaring Creek, AK 

Seal Lake, Canada

Iron Mountain, NM

Victorio Mountains, NM

SE-U.S. tin

spodumene

 belt Vozneskoe Siberia, Russia

Perga, Ukraine

Sierra Blanca, TX

10 t BeO

1,000 t BeO

10,000 t BeO

100,000 t BeO

100 t BeO

Yermakovskoye, Russia

0.001 0.01 0.1 1.0 10 100 1,000

0.001

0.01

0.1

1.0

10

Grade, in weight percent BeO

Deposit tonnage, in million metric tons

0.0001

Carbonate-hosted 

 epithermal beryllium

Deposit type

Volcanogenic beryllium

Pegmatite/granite

Skarn

Greisen

Peralkaline/alkaline,

 general 

EXPLANATION

Figure E8. Graph showing general estimates (using data from table E4) of grade and tonnage for a variety of beryllium 

(Be) deposits, districts, and belts that have a potential for minable beryllium resources. Diagonal lines are isolines of 

contained beryllium oxide, in metric tons (t). The available data for deposits of these types are sparse; some of the deposit 

information has been compiled using poorly documented past production, reserve, and resource information—data types 

that are typically not combined for use in a single plot of grade versus tonnage. BeO, beryllium oxide. Locality: AK, Alaska; 

CO, Colorado; NM, New Mexico; NV, Nevada; SD, South Dakota; SE-U.S., Southeastern United States; TX, Texas; UT, Utah

that underlie the deposit are greisenized. The fine-grained 

ore of diaspore, euclase, fluorite, mica, tourmaline, and white 

chrysoberyl forms replacement veins, pipes, and stringer lodes 

in limestone. The beryllium content of the bulk ore ranges 

from 0.11 to 0.54 percent (0.31 to 1.50 percent beryllium 

oxide); high-grade nodules of chrysoberyl, diaspore, fluorite, 

and mica contain as much as 6 percent beryllium oxide 

(Sainsbury, 1963, 1964a, b; Dobson, 1982).

Apache Warm Springs, New Mexico.—Exploration for 

volcanic-hosted beryllium deposits in New Mexico has 

focused on the Sierra Cuchillo and San Mateo Mountains, 

which extend across Socorro and Sierra Counties (McLemore, 

2010b, undated). Bertrandite at the Apache Warm Springs 

property at Sierra Cuchillo occurs in small quartz veins and 

stringers in fractures and disseminations in a 28-millionyear-old rhyolite and ash-flow tuff showing intense alteration 

to clay minerals. Drilling by BE Resources Inc. of Toronto, 

Ontario, Canada, identified mineralized zones in altered 

rhyolite; however, assays from 13 exploration boreholes 

revealed grades of beryllium oxide well below the cutoff 

grades determined from the company’s beryllium oxide 

benchmarking project (BE Resources Inc., 2011). In addition, 

beryl has been produced from New Mexico pegmatites in 

Grant, Mora, Rio Arriba, San Miguel, and Taos Counties, and 

the majority of the beryl production (about 770 metric tons of 

beryl) came from the Harding pegmatite in Taos County.

Sierra Blanca, Texas.—Beryllium-fluorine mineralization at 

Round Top Mountain, Sierra Blanca, is found with fluorite 

and replaces limestone adjacent to rhyolite (Rubin and others, 

1988). The Sierra Blanca rhyolite is one of six laccoliths of 

similar age (36.2 ±  0.6 mega-annum [Ma]) (Henry, 1992). 

Cyprus Beryllium Corp. reported resources and reserves of 

850,000 metric tons averaging 1.5 percent beryllium oxide 

(0.5 percent beryllium), although the reported data were not 

compliant with National Instrument (NI) 43–101 guidelines 

(Rubin and others, 1990).

Tin-Spodumene Belt, North Carolina and South Carolina.—

Pegmatite dikes in the tin-spodumene belt of North Carolina 

and South Carolina contain an average of about 0.05 percent 

beryllium oxide (Griffitts, 1973). The pegmatites also 

contain cassiterite, columbite, feldspar, mica, and spodumene. 

Resource estimates for the entire tin-spodumene belt suggest 

747,000 metric tons of beryl (111,400 metric tons of beryllium 

oxide) (Griffitts, 1954, p. 10). Individual pegmatite districts (for 

example, Bessemer City and Kings Mountain, North Carolina) 

have in the range of 36,000 metric tons of beryl (equivalent to 

about 5,400 metric tons of beryllium oxide). Little beryl is found 

in the Piedmont province outside of the tin-spodumene belt.
E18 Critical Mineral Resources of the United States—Beryllium

Black Hills, South Dakota.—The Tin Mountain district 

in the Black Hills includes more than 24,000 pegmatites 

associated with the peraluminous Harney Peak granite of 

Paleoproterozoic age. Only about 2 percent of pegmatites 

in the district are zoned, and these have been mined for 

muscovite, potassium-feldspar, and rare metals, such as 

Be, Cs, Li, Nb, Ta, and Sn. The most highly differentiated 

pegmatites, such as at the Tin Mountain and the Bob Ingersol 

Mines, are dikes with lepidolite, petalite, and spodumene 

mineralization (Krogstad and others, 1993; Walker and 

others, 1986; Sirbescu and Nabelek, 2003).

International Resources

Pegmatite beryl deposits are found throughout the 

world (figs. E1 and E8). The more significant global 

exploration sites for nonpegmatite beryllium resources are 

as follows:

Yermakovskoye, Russia.—The Yermakovskoye beryllium 

deposit occurs in a large block of carbonate rock with a 

sedimentary sequence preserved in a field of gabbroic 

plutons, alkali granite, and leucogranite. The phenakitemicrocline-fluorite ores formed at about 225 Ma by 

replacement of limestone (Kislov and others, 2010; Lykhin 

and others, 2010). The deposit has an average grade of 

1.3 percent beryllium oxide (about 0.5 percent beryllium) 

and may be the largest beryllium deposit in Russia (Reyf, 

2008), although as of 2011, no beryllium was being 

produced from this deposit. East Siberian Metals Corp. (a 

subsidiary of Metropol Group) and JSC Ulba Metallurgical 

Plant of Kazakhstan reportedly had completed a predesign 

phase for resuming mining and the technical design for a 

processing plant. The new plant would produce beryllium 

hydroxide for delivery to China, Japan, and Kazakhstan for 

metal and alloy processing. Russia is not known to produce 

any domestic beryllium metal currently; imports from Russia 

are most likely sourced from Kazakhstan (Jaskula, 2013b).

Thor Lake, Northwest Territories, Canada.—The Thor Lake

rare-metals group of deposits, and specifically the North T

deposit, reportedly host one of the highest grade beryllium 

resources known in the world. An NI 43 –101- compliant 

resource estimate prepared by Wardrop Engineering Inc. of 

Canada in 2007 confirmed an indicated resource in one subzone of the North T deposit of greater than 450,000 metric 

tons grading 0.98 percent beryllium oxide. The primary 

beryllium ore mineral in the North T deposit is phenakite. 

The owner, Avalon Rare Metals Inc., saw no near-term 

economic potential for the North T beryllium deposit and 

had no immediate plans to initiate development (Avalon 

Rare Metals Inc., 2013).

Production and Supply

The recent annual worldwide production of beryllium

(mainly from China, Mozambique, Portugal, and the 

United States) was estimated to be 260 metric tons in 2011. 

Of that output, more than 235 metric tons of contained 

beryllium was derived from deposits at Spor Mountain 

(Jaskula, 2013a, b). In 2011, pegmatite beryllium deposits 

accounted for about 25 metric tons of contained beryllium. 

Global trends show an increase in estimated production of 

beryllium for the United States and China for the period 

2000 through 2012 (fig. E9). During pegmatite mining for 

other minerals, beryl is frequently stockpiled for processing at 

a later time. World production and the United States share of 

world production have a high degree of uncertainty because 

China, which is known to be a significant producer, does not 

report its beryl production.

Countries that produce beryl ore include Brazil, China, 

Madagascar, Mozambique, and Portugal. Nigeria, Kazakhstan, 

and Uganda may also produce beryl ore, but information 

is inadequate to make reliable estimates of output. Nations 

that produced gemstone beryl may also have produced 

some beryl ore for industrial use. The United States is the 

world’s leading producer of bertrandite ore, and Materion 

Corp., which was the sole domestic producer, accounted for 

91 percent of estimated world beryllium ore production in 

2011. China, Kazakhstan, and the United States are the only 

countries known to process beryllium ores and concentrates 

into beryllium products. Materion Corp. has manufactured 

products in the form of foil, rods, sheets, and tubes, and 

a variety of other customized shapes at plants in Elmore, 

Ohio, and Fremont, California, and beryllium oxide ceramic 

products at a plant in Tucson, Arizona.

Figure E9. Graph showing estimated global production of 

beryllium for 2000–11. The layers of the graph are placed 

one above the other, forming a cumulative total. Data are 

from U.S. Geological Survey (2007–15).

Beryllium content, in metric tons

United States

China

300

250

200

100

150

0

50

2000 2001 2002 2003 2004 2005 2006 2007 2008 2009 2010 2011 2012

China 

United States

Other

Year

Beryllium figure 10
Exploration for New Deposits E19

In the United States, Materion converts bertrandite from 

open pit mines in Utah and imports beryl ore (at a ratio of 

90 percent bertrandite to 10 percent beryl) into beryllium 

hydroxide at operations near Delta, Utah, in an effort to 

extend its domestic bertrandite reserves. Because beryl has 

fewer impurities than bertrandite, Materion also processes 

imported beryl alone as a means of producing very high 

purity beryllium (Stephen Freeman, President, International 

Business Development, Materion Corp., oral commun., 

January 10, 2013). Some of the beryllium hydroxide is 

shipped to Elmore, Ohio, where Materion converts it into the 

primary feed material used to make beryllium metal products, 

including beryllium-copper master alloy (BCMA), metal, or 

oxide; in the past, some of the primary feed material was sold 

to NGK Insulators, Ltd. of Japan. Materion ceased producing 

beryllium metal in 2000 when its facility was closed owing 

to obsolescence. Since that time, the company has obtained 

beryllium metal from the National Defense Stockpile and 

through imports from Kazakhstan.

In 2011, Materion built a new primary beryllium facility 

in partnership with the DOD to ensure the availability of 

high-quality domestic beryllium metal to meet critical 

defense needs. Under its Defense Production Act Title III 

Program, the DOD invested in a public-private partnership 

with Materion to build a new $90.3 million primary beryllium 

facility in Elmore, Ohio. The construction was completed in 

2011. Materion provided assets, research and development, 

and technology valued at approximately $23.2 million to the 

project; the balance was funded by the DOD. Plant capacity 

was reported to be 73 metric tons per year of high-purity 

beryllium metal; approximately two-thirds of the facility’s 

output is to be allocated for defense and Government-related 

end uses, and the remaining output is to go to the private 

sector (Metal Bulletin, 2010; Materion Corp., 2012a).

Kazakhstan’s Ulba Metallurgical Plant, which was a 

subsidiary of JSC National Atomic Company “Kazatomprom,” 

reported that it supplied about 26 percent of the world’s beryllium products in 2011 compared with 3 percent in 1999. The 

Ulba Metallurgical Plant reportedly produced these products 

from stockpiled beryllium concentrate imported mainly from 

Russia. The concentrate stockpile, which was built up before 

the breakup of the Soviet Union, was forecast in 2003 to be 

sufficient to support production for about 30 years (Metal 

Bulletin, 2003; McNeil, 2006; Kazatomprom JSC, 2012).

More recent reports (2015) indicated that only a few 

years of supply remain in the concentrate stockpile. In 2015, 

it was estimated that the Ulba Metallurgical Plant produced 

about 90 metric tons of beryllium contained in berylliumcopper alloys, beryllium oxide ceramics, and beryllium metal. 

Owing to anticipated demand for beryllium, Russia is in the 

process of resuming mining and constructing a processing 

plant at the Yermakovskoye deposit in the Siberian Republic 

of Buryatiya. The new plant will produce beryllium hydroxide 

for export to the Ulba Metallurgical Plant in Kazakhstan, 

as well as to China and Japan for processing into beryllium 

metal and beryllium alloys. The annual mining and processing 

capacities are estimated at 25,000 metric tons per year of 

bertrandite ore and 130 metric tons per year of beryllium 

hydroxide. In 2015, Russia announced it would produce its 

own beryllium metal on a commercial scale by 2020.

China’s Shuikoushan Non-Ferrous Metals Co., Ltd. 

processed beryllium ores and produced beryllium products 

in various forms at its 6th Smelting Plant in Songbai, 

Changning City, Hunan Province. The plant had an estimated 

production capacity of 150 metric tons per year of beryllium 

oxide and 1,500 metric tons per year of BCMA (McNeil, 

2006). Fuyun Hengsheng Beryllium Industry Co., Ltd.’s 

refinery (a joint venture among Xinjiang Nonferrous Metals 

Industry Group, Xinjiang Henghe Investment Co., Ltd., 

and Xinjiang Nonferrous Metals Industry Group Nonferrous 

Metals Co.) was designed to produce 100 metric tons per year 

of beryllium oxide and 800 metric tons per year of BCMA. 

The refinery is located in Fuyun County in northwestern 

Xinjiang Uygur Autonomous Region (China Non-Ferrous 

Metal Import & Export Xinjiang Corp., undated). In 2015, it 

was estimated that China produced about 65 metric tons of 

beryllium contained in beryllium-copper alloys, beryllium 

oxide ceramics, and beryllium metal at these two plants. 

Approximately 20 metric tons of the beryllium was sourced 

from domestic ore and 45 metric tons was obtained from 

ore derived from foreign sources, most likely Kazakhstan’s 

Ulba Metallurgical Plant.

Exploration for New Deposits

Early exploration methods to map the occurrence of 

deposits in highly evolved felsic rocks and fluorine-rich 

hydrothermal systems used a neutron-sourced gamma ray 

spectrometer capable of rapid semiquantitative field analysis 

of the beryllium contents of rocks (Meeves, 1966). Current 

exploration and assessment methods are based on identifying 

favorable geotectonic settings, beryllium-bearing magmatic 

rock suites, and local and regional alteration patterns that 

might target economic mineralization (Foley and others, 2012). 

Magnetic, gravity, and seismic data are useful in identifying 

large-scale igneous suites in favorable geotectonic settings that 

have a potential to contain volcanogenic beryllium deposits. 

Rhyolitic rocks rich in fluorine and beryllium that show 

evidence of explosive volcanism, such as unsorted carbonate 

rock fragments in breccias and vent structures capable of 

tapping the sources of rhyolitic magma and mineralizing 

fluids at depth, are a favorable volcanic host rock. The richest 

beryllium ore zones may be associated with mineralized 

carbonate-fragment-rich deposits of tuff. The presence of 

fluorite and uranium deposits suggests the presence of fluids 

capable of transporting sufficient amounts of beryllium to form 

a deposit (Lindsey, 1975). Hydrothermal alteration minerals, 

especially the presence of abundant lithium-smectite, and 

geochemical halos of elements, including Be, Ce, F, Ga, Li, 

Nb, and Y, occur in rocks surrounding deposits (Lindsey and 

others, 1973). Economic deposits are likely to be made up of 
E20 Critical Mineral Resources of the United States—Beryllium

many small mineralized lenses that, as a whole, may cover 

50 square kilometers or more but that individually may cover 

only hundreds of square meters. For example, the surface 

expression of the Apache Warm Springs deposit is approximately 10 meters by 40 meters — comparable in size to one 

of the smaller individual deposits at Spor Mountain.

Exploration for beryl-rich LCT pegmatites is typically 

carried out within major known pegmatite fields where field 

relations are clear and discoveries of new pegmatite bodies 

may be readily mined (Bradley and McCauley, 2013). In areas 

where bedrock is well exposed, pegmatites stand out because 

they are light-colored, have large crystals, and are resistant to 

weathering relative to surrounding country rocks. An initial 

exploration step is the identification of appropriate granitic 

rock types, although, because pegmatites can invade any rock 

type, the absence of granite in surface exposures does not 

eliminate the potential occurrence of LCT pegmatites. In areas 

where exploration has been minimal, first-order criteria, as 

established in Bradley and McCauley (2013), are an orogenic 

hinterland (mountain-building) setting, rocks having metamorphic grades suggesting conditions of medium pressure and 

medium to high temperatures, and the presence of evolved 

granites and common granitic pegmatites. Pegmatites that 

show a regional mineralogical and geochemical zoning pattern 

concentric about the inferred parental granite are favorable 

for beryl. Exploration and assessment for LCT pegmatites 

are guided by a number of important field and laboratory 

observations. Because beryl is associated with the highly 

fractionated pegmatite zones and because these may constitute 

only 1 to 2 percent of a field, mineralogical and geochemical 

variations in granite-pegmatite zones are important (Selway 

and others, 2005). For example, beryl reportedly shows color 

changes from green-brown to pink-white, with increasing fractionation in more evolved bodies (Trueman and Černý, 1982). 

Also, mapping of chemical changes in specific minerals, 

such as increasing rubidium in potassium feldspar, increasing 

lithium in white mica, increasing manganese in garnet, and 

increasing tantalum and manganese in columbite-tantalite, 

may be used to locate LCT pegmatite bodies within larger 

fields of common pegmatite. Many LCT pegmatites show 

haloes of alkali elements (cesium, lithium, and rubidium) in 

surrounding wall rocks and soils that have been used to locate 

buried rare-metal pegmatites (Galeschuk and Vanstone, 2005, 

2007). In weathered environments, anomalies of As, Be, Sb, 

and Sn and resistant minerals, such as cassiterite, in soils may 

also indicate LCT pegmatite (Smith and others, 1987).

Exploration for new deposits is limited in North America 

because domestic beryllium production is dominated by a 

single producer (Materion Corp.) that effectively controls the 

relatively small and specialized domestic beryllium market—

a market that cannot readily accommodate new competition 

on the raw material supply side. Many new discoveries of 

volcanic-hosted bertrandite are extensions of known deposits. 

Further exploration and mining of beryl from pegmatite is 

unlikely in the United States given the current high costs 

associated with labor-intensive beryl extraction.

Environmental Considerations

Sources and Fate in the Environment

The behavior of beryllium in the environment is dominated by its low solubility in water and that it is commonly 

found in solid form. Natural beryllium in the atmosphere 

is mostly from dust sources, including windblown dust 

and volcanic eruptions. The main anthropogenic source of 

beryllium in the atmosphere is from the combustion of coal 

and fuel oil, followed by beryllium processing and municipal 

waste combustion (National Research Council, Committee on 

Beryllium Alloy Exposures, 2007). Globally, the atmospheric 

content of beryllium varies between 0.3 and 300 parts per 

trillion (ppt) (0.1 to 100 nanograms per cubic meter [ng/m3]) 

(Kabata-Pendias and Mukherjee, 2007). Beryllium concentrations in air in rural areas of the United States ranged from 

0.08 to 0.2 ppt (0.03 to 0.06 ng/m3

) and in urban areas ranged 

from less than 0.3 to 18 ppt (0.1 to 6.7 ng/m3

); up to 224 ppt 

(82.7 ng/m3) was reported in the vicinity of a beryllium 

processing plant in Pennsylvania (Bruce and Odin, 2001). 

The average upper crustal abundance of beryllium is 1.9 ppm 

(Hu and Gao, 2008) and natural soils are reported to contain 

between 0.48 to 3.52 ppm beryllium (Kabata-Pendias and 

Mukherjee, 2007). In the conterminous United States, the 

concentration in surface soils (0 to 5 cm depth) averages 

1.3 ppm and ranges from less than 0.1 to 17.3 ppm (Smith 

and others, 2013).

In waterways, the natural weathering of rocks and 

soils, deposition of atmospheric beryllium, and industrial 

wastewater effluents are the main sources of beryllium. The 

mobility of beryllium in the aqueous environment is limited 

because of the poor solubility of most beryllium compounds 

and its affinity to bind with clay minerals and organic 

compounds; beryllium commonly resides in sediment or soils. 

Beryllium mobility is enhanced under acidic water conditions 

and may be found as a dissolved cation (beryllium2+) or may 

form soluble fluorine and organic complexes (Kabata-Pendias 

and Mukherjee, 2007). At near-neutral pH, which is typical of 

most surface and groundwaters, beryllium is controlled by the 

formation of beryllium hydroxide, which is a low solubility 

phase. Globally, rivers commonly contain between less than 

0.008 and 0.6 ppb beryllium and have an average of 0.009 ppb 

beryllium (Gaillardet and others, 2003). Surface waters in the 

United States have a mean concentration of 0.07 ppb beryllium; groundwaters contain comparably low concentrations, 

and seawater contains even lower concentrations (Eckel and 

Jacob, 1988; Bruce and Odin, 2001). Beryllium concentrations 

in industrial effluents are several orders of magnitude higher 

and range from 30 to 170 ppb (Bruce and Odin, 2001).

Pre-mining baseline geochemical data for beryllium 

deposits are sparse. LCT pegmatite deposits may exhibit 

enrichments in B, Cs, F, Li, and Rb in surrounding soils and 

waters. These enrichments form as surrounding country rock 

is hydrothermally altered as a result of pegmatite emplacement 

(Bradley and McCauley, 2013), and they can persist in soils 
Environmental Considerations E21

that develop when the rocks are weathered. For example, the 

Tanco pegmatite in Canada has alteration halos of lithium 

and rubidium over the buried deposit (Trueman and Černý, 

1982). In general, volcanogenic beryllium deposits may have 

enriched halos of F, Li, Mo, Nb, REEs, Sn, Ta, Th, U, and W 

in surrounding volcanic rocks and mineralized tuff; therefore, 

soils and waters in the vicinity may contain anomalous 

concentrations of the elements (Foley and others, 2012). The 

stream waters near the Yermakovskoye deposit in Russia 

contain naturally elevated concentrations of Cu, F, Fe, Hg, 

Mn, Mo, and Zn (Kislov and others, 2010). The solubility 

of beryllium minerals is low under most naturally occurring 

conditions; thus, beryllium contents are not expected to be 

anomalous in waters draining beryllium deposits, although 

soils may be enriched.

Mine Waste Characteristics

Mine waste is material that is generated during ore 

extraction and processing at the mine site as part of the 

mining operation; it has no economic value. The volume and 

characteristics of the mine waste depend on the deposit type 

exploited and its size, whether multiple mineral commodities 

are being recovered, and the mining and processing methods 

used. Mine waste also includes waste generated in the 

chemical conversion of concentrated beryllium ores to 

beryllium hydroxide at the mine site. Additional processing 

steps to convert beryllium hydroxide to beryllium metal, 

oxides, or alloys also create waste; this type of waste is 

not discussed in this paper because these operations are 

commonly conducted at processing plants distant from the 

mine sites. In general, the mining of beryl and bertrandite ores 

creates mine waste that is similar in composition to the rocks 

that host the beryllium-containing minerals because the ores 

grades are low (table E4).

The ore from the bertrandite deposit at Spor Mountain, 

Utah, contains about 1 percent or less bertrandite; therefore, 

nearly all of the mined and processed ore becomes waste. 

Mining at Spor Mountain occurs with a stripping ratio of 

overburden to ore of 23:1, which indicates that significant 

amounts of overburden also have to be disposed of (Hawkins, 

2001). At Spor Mountain, the composition of the mineralized 

tuff and the overlying rhyolite is likely representative of 

the bulk geochemical characteristics of the waste material. 

Quaternary alluvium overlying the mineralized tuff and 

rhyolite is mainly Pleistocene sediments of Lake Bonneville. 

The geochemical characteristics of the mineralized tuff 

at Spor Mountain have been summarized by Lindsey 

(1975), and, for the most part, reflect the rhyolitic protolith 

(Si>Al>K>Ca≈Na>Fe≈Mg) with trace-element enrichments in beryllium, fluorine, lithium, thorium, and uranium 

relative to unmineralized counterparts. The mineralogy of the 

tuff is dominated by clays, cristobalite, potassium-feldspar, 

and quartz; calcite and dolomite clasts are locally important. 

Fluorite and manganese oxides are locally abundant in the ore 

assemblages, and uranium is thought to be bound in fluorite 

and opal (Lindsey, 1981). The lack of sulfide minerals and the 

presence of carbonate minerals as clasts suggest that the waste 

material is net-alkaline (has a pH greater than 7); this measure 

indicates that the rocks will not contribute acid to surrounding 

bodies of water as they weather.

At the Yermakovskoye Mine site in Russia (fig. E1), 

future mining and processing of the ore reserves is expected 

to extract 8.5 million cubic meters of rock from an open pit 

operation and to produce 830,000 cubic meters of tailings 

during its approximate 20-year mine life (Kislov and others, 

2010). The ore minerals of fluorite (about 20 percent fluorite), 

bertrandite, and phenakite are associated with albite, calcite, 

dolomite (about 12 percent carbonate), microcline, and quartz 

with minor amounts of sulfides, including galena, pyrite, 

and sphalerite. Mine tailings from production that ended in 

1989 contains up to 241 ppm beryllium, which is well above 

background concentrations in the area, which are between 

1.5 and 9.9 ppm beryllium (Kislov and others, 2010). The 

reopening of this mine site as proposed would include a 

processing facility that would use flotation to separate the 

sulfides, which would create sulfide-rich tailings, followed by 

beryllium and fluorite flotation, which would create additional 

tailings. The sulfur content of the ore is up to 0.37 percent 

sulfur, including the contribution from pyrite, which can 

generate acid upon weathering. Sulfide tailings could be 

potentially acid generating, but the high carbonate content of 

other mine waste may help neutralize the acid generated.

Beryl in pegmatites can be associated with a range of 

host rock types, but, in general, processing of pegmatite 

ores generates mill tailings of relatively simple mineralogy. 

Enriched zones of beryl (approximately 2 percent) usually are 

associated with muscovite, plagioclase, potassium-feldspar, 

and quartz and may be from zones in which amblygonite, 

feldspar, and spodumene are recovered (Sabey, 2006). Beryl 

has also been found in quartz veins that contain cassiterite, 

molybdenite, wolframite, and other minerals. Enriched-beryl 

zones commonly consist of a few thousand tons of rock, with 

the largest containing up to 1 million metric tons. When it 

was in operation, the Tanco Mine generated tailings that were 

predominantly feldspar with some mica, quartz, pollucite, 

and spodumene when the ore was processed to recover 

spodumene and tantalum (Solylo, 2012). Tailings produced 

after beryl separation would likely be compositionally similar. 

Additional insight into the mineralogy of mine waste from 

pegmatite deposits was gained from experiments conducted 

on recovering beryl from several small operations that were 

mining spodumene in Arizona, Colorado, New Mexico, and 

Canada. Quartz and feldspar tailings from gangue as well as 

mica-schist tailings from overburden were generated. Also, 

in a later processing step to concentrate the beryl, apatite and 

iron-bearing gangue minerals, including garnet, pyroxene, 

and tourmaline, were separated with a magnetic separator, 

which generated additional waste (Bhappu and Fuerstenau, 

1964). Other research on recovering beryl from the former 

Foote Mine near Kings Mountain, North Carolina, generated 

predominantly feldspar and quartz tailings along with separate 
E22 Critical Mineral Resources of the United States—Beryllium

Table E4. Location, grade, tonnage, and other data for selected beryllium deposits of the world.

[Compiled from Griffitts (1954, 1973), Barton and Young (2002), Sabey (2006), Secher and Appel (2007), Brush Engineered Materials, Inc. (2009), and 

McLemore (2010a). Be, beryllium; BeO, beryllium oxide; %, percent; —, no data]

Deposit Location Country Deposit

type

Resource

(metric tons)

Contained 

BeO

(metric tons)

Grade 

(% BeO)

Grade1

(% Be)

Aguachile Coahuila Mexico Carbonatehosted17,000,000 — 0.1 0.036

Apache Warm

Springs

New Mexico United 

States

Volcanogenic 43,060 — 0.26 0.094

Aqshatau Aqshatau Kazakhstan Greisen 16,000 — 0.03 to 0.07 0.02

Atlantic shield Brazil Brazil Pegmatite/

granite

106,000,000 — 0.04 0.015

Black Hills South Dakota United 

States

Pegmatite/

granite

— 13,000 0.05 0.02

Boomer, Lake George Colorado United 

States

Greisen <1,000 to 3,000 — 2.0 to 11.2 2.3

Brockman Western Australia Australia Volcanogenic 4,300,000 — 0.08 0.03

Gold Hill Utah United 

States

Veins — 5,000 — 0.5

Hellroaring Creek British Columbia Canada Pegmatite — 1,000 0.1 0.036

Ilimaussaq (general) Ilimaussaq Greenland Peralkaline 

intrusion

— 20,000 — 0.18

Iron Mountain New Mexico United 

States

Carbonatehosted1,000 — 0.2 to 0.7 0.18

Kvanefjeld Ilimaussaq Greenland Peralkaline 

intrusion

180,000 — — 0.1

Lost River Alaska United 

States

Skarn 3,000,000 10,000 0.3 to 1.75 0.36

McCullough Butte Nevada United 

States

Carbonatehosted175,000,000 47,000 0.027 0.01

Mt. Wheeler Nevada United 

States

Veins — 1,000 0.75 0.27

Rodenhouse Wash Utah United 

States

Pegmatite/

granite

1,000,000 — 0.5 0.18

Seal Lake Northwest

Territories

Canada Peralkaline 

intrusion

— 6,800 0.35 to 0.4 0.14

Sheeprock Utah United 

States

Greisen 1,000,000 — 0.01 to 0.1 0.02

Sierra Blanca Texas United 

States

Carbonatehosted— 11,300 0.5 to 1.9 0.36

Spor Mountain Utah United 

States

Volcanogenic 6,425,000 — 0.266 to 0.720 0.18 to 0.26

Strange Lake Quebec, and 

Newfoundland 

and Labrador

Canada Peralkaline 

intrusion

— 42,000 0.08 0.03

Tanco2 Manitoba Canada Pegmatite/

granite

900,000 1,800 0.2 0.07

Thor Lake Northwest

Territories

Canada Peralkaline 

intrusion

1,600,000 13,300 0.76 0.28

Tin-spodumene belt, 

North Carolina and 

South Carolina 

North Carolina

and South 

Carolina

United 

States

Pegmatite/

granite

— 111,000 0.02 to 0.1 0.031

Ukrainian Shield Ukraine Ukraine Greisen 20,000 — 0.4 0.11

Victorio Mountains New Mexico United 

States

Carbonatehosted11,900,000 — 0.023 0.01

Vozneskoye Siberia Russia Skarn 10,000 — 0.06 0.02

Yermakovskoye Siberian Republic,

Buryatiya

Russia Skarn — 10,000 1.3 0.47

1

Grade used for constructing figure E9.

2

Also known as Bernic Lake.
Environmental Considerations E23

beryl, mica, and spodumene concentrates (Browning, 1961). 

In addition, discarded spodumene tailings at the mine, 

which consisted of feldspar and quartz with minor amounts 

of mica, spodumene (approximately 5 percent), and beryl 

(approximately 0.5 percent) were reprocessed to recover 

beryl, spodumene, and, sometimes, feldspar (Browning and 

others, 1964).

In addition to the waste rock, overburden, and mill tailings 

generated at beryllium mine sites, converting the concentrated 

beryllium ores to beryllium sulfate and subsequently to 

beryllium hydroxide produces additional waste. To remove 

impurities and waste minerals, the beryllium is extracted, 

purified, and precipitated using organic or other solutions. 

The waste solutions and precipitates from various steps are 

discarded in waste ponds and may be acidic. Waste ponds may 

be neutralized and diluted to offset high corrosivity and reduce 

the overall concentrations of Be, Cr, Cu, Se, and cyanide. 

Waste solutions containing uranium are commonly transferred 

to separate storage ponds and shipped elsewhere for U-metal 

recovery. Waters in the ponds are commonly evaporated to 

generate a solid waste product (U.S. Environmental Protection 

Agency, 1989). One study that examined the characteristics 

of red mud solid waste resulting from the fluoride separation 

process for beryl ore found that it was composed mainly 

of aluminum, iron, and silicon, with significant amounts of 

beryllium (Bhat and others, 2002). For example, red mud 

at a processing facility in India contained 0.4 to 0.6 percent 

beryllium, and approximately one-half of the beryllium was 

leached using water (based on laboratory experiments), which 

suggests that much of the beryllium in red mud is mobile in 

surficial environments (Bhat and others, 2002).

Human Health Concerns

Beryllium is toxic and a human carcinogen that can 

impair the function of the blood vessels, endocrine glands, 

heart, and liver (Taylor and others, 2003; Kabata-Pendias 

and Mukherjee, 2007). The general population is exposed 

to beryllium mainly from ingestion of beryllium-containing 

food and water, with smaller contributions from inhalation of 

dust (Bruce and Odin, 2001). Ingested beryllium is less likely 

than inhaled beryllium to cause health concerns because most 

beryllium that is ingested passes through the body without 

entering the bloodstream. In contrast, beryllium particles 

inhaled can dissolve in the lungs and move to the bloodstream or remain in the lungs for up to years after exposure 

has ceased.

An international guideline value for beryllium in drinking 

water has not been established because beryllium is “rarely 

found in drinking water at concentrations of health concern,” 

according to the World Health Organization (2011). For the 

United States, however, the maximum contaminant level for 

beryllium in drinking water is 4 ppb (U.S. Environmental 

Protection Agency, 2009). Beryllium concentrations in excess 

of this drinking water standard are unlikely because most 

beryllium compounds and minerals are poorly soluble in 

water. According to the U.S. Environmental Protection Agency 

(EPA) (2009), consumption of waters exceeding action levels 

may lead to intestinal lesions. The average concentration of 

beryllium in drinking water in the United States is 0.19 ppb 

(which is significantly less than the guidelines), and ranges 

from 0.01 to 1.22 ppb (Agency for Toxic Substances and 

Disease Registry, 2002; Sabey, 2006). As for soils, screening 

levels based on human health risk in the United States have 

been established to determine whether levels of contamination 

at hazardous waste sites warrant further investigation or site 

cleanup. The noncancer screening level for beryllium in residential soils is 160 ppm, and in industrial soils, it is 2,000 ppm 

(U.S. Environmental Protection Agency, 2013). As previously 

stated, the range in beryllium concentrations in surficial 

soils of the United States are from less than 0.1 to 17.3 ppm 

(Smith and others, 2013), although orders of magnitude higher 

concentrations can be found in soils, such as those formed by 

weathering of beryllium-rich rocks.

Human health concerns from exposure to beryllium are 

mostly associated with inhaling airborne particles during the 

mining and processing of beryllium ores and the manufacturing of beryllium-containing products. The occupational 

threshold exposure level value for beryllium in the atmosphere 

is 5 ppb (2 milligrams per cubic meter [mg/m3]) for an 8-hour 

exposure period and 68 ppb (25 mg/m3) for a 30-minute 

exposure period (National Institute for Occupational Safety 

and Health, 2005). The level of occupational exposure to 

beryllium inhalation has decreased from early years of 

beryllium production in the 1930s to modern workday 

environments (National Institute for Occupational Safety and 

Health, 2005). Processing facilities implement safety procedures to minimize exposure that include regularly washing 

hands and laundering clothes; personal safety equipment, such 

as respirators; and the collection of dusts, fumes, and mists 

(Petkof, 1985).

Beryllium dust and fumes have been recognized to 

cause chronic beryllium disease, which is a debilitating and 

potentially lethal lung disease, and its precursor, beryllium 

sensitization, which is an immune system disorder. The 

human health risk is variable among the different sectors of 

the beryllium industry, and only mining and ore processing 

that occur at the mine site are considered here. A study of 

workers from the Spor Mountain Mine and mill reported that 

beryllium exposure levels during mining and milling were 

generally less than the threshold exposure level of 5 ppb 

(2 micrograms per cubic meter [μg/m3]) (Deubner and others, 

2001). Another study at the site by Stefaniak and others 

(2008) reported that concentrations in the air among various 

work areas ranged from 0.003 to 6 ppb (0.001 to 2.1 μg/m3). 

There may be a lower risk of sensitization and of developing 

chronic beryllium disease from beryllium during mining 

and ore (both bertrandite and beryl) processing than during 

other phases of beryllium manufacturing, including the 

production of beryllium oxide, beryllium ceramics, and 

copper-beryllium alloys (Deubner and others, 2001, 2011). 
E24 Critical Mineral Resources of the United States—Beryllium

Deubner and others (2011) hypothesize that the decreased risk 

is because the beryllium-containing materials found at the 

Spor Mountain facility dissolve faster in lung fluid than do 

other beryllium-containing materials. The physicochemical 

properties of beryllium, such as chemical composition, particle 

size, and specific surface area, influence bioavailability and 

may be more important than the exposure concentration 

in determining risk of developing beryllium sensitization 

or chronic beryllium disease (National Research Council, 

Committee on Beryllium Alloy Exposures, 2007; Stefaniak 

and others, 2008). In addition, genetic susceptibility is likely 

an additional risk factor (Taylor, Ding, and others, 2003).

Although inhalation is the main industrial exposure 

pathway of concern, skin exposure to beryllium may result in 

it entering the body, contributing to sensitization and causing 

irritation, lacerations, and chemical ulcers, and granulomas 

(Epstein, 1991; Day and others, 2006; National Research 

Council, Committee on Beryllium Alloy Exposures, 2007; 

Duling and others, 2012). Duling and others (2012) found the 

beryllium from bertrandite ores from Spor Mountain dissolves 

in artificial sweat as well as synthetic lung fluids.

Radiation exposure from uranium or thorium associated 

with volcanogenic beryllium ore may be a human health 

concern during mining and processing operations, although 

specific case studies could not be identified. Nonetheless, 

anomalous amounts of uranium (up to 200 ppm) and thorium 

(up to 150 ppm) are associated with the beryllium ores at 

Spor Mountain (Lindsey, 1981). If uranium is concentrated in 

waste generated from converting ore to beryllium hydroxide, 

precautions need to be taken when handling the waste.

Ecological Health Concerns

Beryllium is toxic to plants and animals; however, similar 

to human health concerns, ecological issues are associated 

mainly with the mining and manufacturing of beryllium. 

In uncontaminated areas, the concentration of beryllium in 

plants is commonly low (less than 1 ppm), and soils tend 

to contain beryllium in forms that are unavailable to plants 

(Bruce and Odin, 2001; Rossman, 2004). In some cases, 

beryllium is present as soluble salts in soils (for example, 

beryllium chloride [BeCl2 ] and beryllium sulfate [BeSO4 ]) 

that are bioavailable and potentially may cause toxic effects, 

including reduced growth and degradation of some proteins 

and enzymes, and may inhibit the uptake of calcium, magnesium, and phosphorus. Toxicity of beryllium to plants has been 

observed for soil concentration ranges of from 10 to 50 ppm 

(Kabata-Pendias and Mukherjee, 2007). Additionally, soils 

with concentrations of beryllium between 20 and 40 ppm have 

been shown to affect the survival and reproduction of soil 

invertebrates (Kuperman and others, 2006). Also, beryllium, 

both inhaled and ingested, has been shown to be acutely toxic 

to mammals in laboratory tests, but information regarding 

toxicity to wildlife was not found (Bruce and Odin, 2001).

Beryllium in water and stream or lake sediment is 

not significantly concentrated by aquatic organisms or 

accumulated within the food chain (Bruce and Odin, 2001). 

In laboratory tests, beryllium (administered as soluble salts) 

is acutely toxic to a variety of freshwater fish and water fleas 

(Daphnia magna), but toxicity varied widely depending on 

the species studied and the test conditions; soft and acidic 

waters had the highest toxicity associated with the lowest 

concentrations of beryllium (Bruce and Odin, 2001, and 

references therein). Water-quality criteria and stream-sediment 

guidelines for the protection of aquatic life are generally 

lacking; however, the EPA does provide lowest observable 

effect limits (LOELs) as guidance for acute and chronic 

toxicity to aquatic life for beryllium of 130 ppb and 5.3 ppb, 

respectively (U.S. Environmental Protection Agency, 1980). 

Also, Suter (1996) suggests secondary acute and chronic 

aquatic ecosystem guidelines for surface waters of 271 ppb 

and 5.09 ppb beryllium, respectively. These guidelines are 

generally higher than the mean concentration of 0.07 ppb 

beryllium reported for surface waters in the United States 

(Eckel and Jacob, 1988). In general, the concentrations needed 

to cause toxicity to aquatic organisms from the dissolution of 

bertrandite or beryl would likely occur only in soft waters with 

acidic pH values.

The mining and processing of beryllium ores may 

enrich waters and soils in potentially toxic trace elements and 

contaminate the ecosystem at some mine sites. For example, 

water from the open pit at a previously mined, skarn-related, 

beryllium-fluorite deposit in Russia (the Yermakovskoye 

deposit) contained elevated concentrations of some elements, 

including Cu, Fe, Mn, Mo, Mg, Pb, and Zn, that exceeded 

local standards for water bodies used for fishery purposes. 

Additionally, sediment at the bottom of the pit was contaminated with As, Be, Cd, Pb, and Zn, and the concentrations of 

lead and zinc exceeded values above which effects to aquatic 

biota are likely. Mine waste piles contained concentrations of 

beryllium that were above the maximum allowable concentrations (Kislov and others, 2010). The natural background 

contents of some of the elements, such as Cu, Fe, Mn, Mo, 

and Zn, in local streams and melt waters were also elevated 

and above environmental guidelines, which made it difficult 

to distinguish the contamination from mining compared to 

pre-mining concentrations associated with the mineralized 

area. According to Kislov and others (2010), the pasture 

vegetation in the local area may contain high amounts of 

beryllium from the deposition of contaminated dust, and the 

agricultural uses of the land should be limited.

In contrast to the mine site in Russia, the Spor Mountain 

deposit is characterized by an arid climate, a lack of surface 

water, and low concentrations of beryllium and uranium in 

both the volcanic host rock and ore materials. Ecosystem 

issues related to beryllium mining are not known in the 

vicinity of the Spor Mountain deposit.

Carbon Footprint

Beryllium has the potential to reduce the carbon footprint 

from energy production because of its potential for use in wind 
Problems and Future Research E25

energy and hydrogen storage and fuel cell applications. IBC 

Advanced Alloys Corp. of Canada and a Washington, D.C.-

based clean-energy consulting company are working to 

commercialize wind energy applications of beryllium. The 

use of beryllium is expected to increase wear resistance 

in wind turbines, and thus reduce the cost of operations 

(IBC Advanced Alloys Corp., 2013; Jaskula, 2013a). IBC 

has also collaborated with Hydrogen Link Inc. of Canada to 

advance research on the use of beryllium in hydrogen fuel 

storage. Lithium-beryllium hydrides have the highest viable 

hydrogen capacity at reasonable temperatures and can be 

used for small-scale applications, such as laptops and other 

portable devices (IBC Advanced Alloys Corp., 2010).

Mine Closure

The post-mining landscape at beryllium mines depends 

on the deposit type and size, the mining and processing 

methods, the climate, and regulations regarding mine 

closure. Open pits, shafts, and adits from underground 

operations, processing facilities, and mine and processing 

wastes may need to be addressed for mine closure; some 

mines are not reclaimed. For example, remediation efforts 

were not undertaken when production ceased in 1989 

at the Yermakovskoye deposit in Russia (Kislov and 

others, 2010). The shallow open pit covered a surface 

area of 277,000 square meters. The pit floor was at least 

70 m deep at the deepest part, and the pit contained a few 

meters of water (Kislov and others, 2010). Future plans 

for the site included resuming mining and building a 

processing plant. Construction was expected to begin in 

2013 (Jaskula, 2013b). The life of the mine was expected 

to be 17 to 20 years (Kislov and others, 2010). Eventually, 

mine closure will involve stabilizing disturbed surfaces, 

reclaiming soils, waters, and stream sediments at and near 

the mine site, and addressing any issues associated with the 

open pit, waste-rock piles, and tailings.

Similar remediation efforts may also be undertaken 

at the Spor Mountain Mine after mining there ceases. The 

mining operations thus far have created several elongated 

open pits created to exploit the large sheet-like orebodies. 

Stockpiles of ore and waste piles of overburden material 

cover areas that are hundreds of meters in diameter.

Beryl is typically mined from small-scale operations 

using minimal mining equipment, commonly as a byproduct 

of feldspar and mica mining. Mining of ore is typically in 

quantities ranging from a few kilograms to hundreds of tons 

(Sabey, 2006). These types of operations commonly do not 

have mine closure plans. Some operations where beryl is 

mined as a byproduct are larger operations, and they leave 

a landscape of waste rock, tailings, and open pits after 

mining has ceased. For example, the flooded open pit at the 

Foote Mine in the Kings Mountain district, North Carolina, 

is several hundreds of meters in diameter, and waste 

rock and tailings cover an even larger area that is devoid 

of vegetation.

Problems and Future Research

Research efforts into understanding the factors that 

may affect future supplies of beryllium include developing 

assessment models for new deposits, conducting experimental 

studies on beryllium at conditions applicable to biochemical 

processes occurring in surface environments, and devising 

new and more efficient extraction technologies for both 

raw ore and recycled products. These are described in more 

detail below.

Assessment models.—Assessing for beryllium in volcanogenic systems is problematic at present because there is 

only one known and essentially unique economic deposit. 

Comprehensive modern geologic and geochemical studies 

of the exemplary deposit at Spor Mountain and comparative studies of additional regions and related deposit types 

(carbonate-hosted and fluorite deposits)—for example, the 

central Gobi volcanic zone in Mongolia, and the REE-rich 

alkaline volcano-plutonic rocks within the TransbaikalMongolian rare-metal province, Siberia—are needed to 

provide for more widely applicable deposit models and 

assessment methods. New approaches to evaluating the 

geologic availability of beryllium are also needed to assess 

pegmatite resources and other rare-metal systems, such as 

rare-metal granites, pegmatites, and skarn and greisen rocks 

of Alaska, where economics may require extraction and 

recovery of byproducts and coproducts (such as Be, Cs, F, Li, 

Nb, Si, Ta, and so forth).

Experimental studies.—Thermodynamic data on beryllium 

in various fluid-melt systems are needed to gain a better 

understanding of transport mechanisms for beryllium in a 

magmatic-hydrothermal environment and in fluorine-rich 

magmatic systems having rare-metal granites and associated 

beryllium-enriched pegmatites. These data have practical 

implications for resource and environmental assessment 

because they are needed to model these specialized 

geologic environments.

Environmental aspects.—The advent of new technologies 

and new products that use beryllium in conjunction with 

other elements, such as beryllium oxide enhanced nuclear 

fuel pellets, will likely require further study into the possible 

effects of use and disposal of complex and hazardous alloys, 

manufactured products, and waste materials.

Extraction technologies.—Improved extraction technologies may eventually be devised to handle materials that are 

currently below cutoff grades for volcanogenic deposits. 

Better refining methods may also be developed to separate 

contaminants (for example, Al, BeO, C, Fe, Mg, and Si) from 

raw beryllium minerals so that very high purity beryllium 

metal can be made from any beryllium source mineral.

Recycling.—As production levels and uses increase, the 

percentage of recycled beryllium from both new and old 

scrap may be expected to increase if separation technologies 

are improved.
E26 Critical Mineral Resources of the United States—Beryllium

New markets.—Materials producers will continue to work 

closely with users, especially the aerospace and defense 

industries, to meet advances in end-use products and to ensure 

that new materials are developed to meet high-performance 

requirements. According to some industry experts, the 

global beryllium market forecast is for consumption to 

reach greater than 500 metric tons per year within 5 years 

and for the growth in consumption to be shaped by upward 

trends in some important end-use application areas (Global 

Industry Analysts, Inc., 2012). In particular, growing security 

concerns are proposed to increase the demand for beryllium 

in defense applications. Further, the advent of new alloys 

of beryllium with aluminum is expected to drive the beryllium market in the United States and other countries as 

new products are developed. If successful, the beryllium 

oxide enhanced nuclear fuel pellet could increase demand 

for beryllium substantially.

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(GTOPO30): Reston, Va., U.S Geological Survey dataset 

(digital elevation model), accessed February 12, 2014, 

at https://lta.cr.usgs.gov/GTOPO30.

U.S. Geological Survey, 2007–15, Beryllium, in Metals and 

minerals: U.S. Geological Survey Minerals Yearbook 

2004–2013, v. I, variously paged.

Walker, R.J., Hanson, G.N., Papike, J.J., and O’Neil, J.R., 

1986, Nd, O, and Sr isotopic constraints on the origin of 

Precambrian rocks, southern Black Hills, South Dakota: 

Geochimica et Cosmochimica Acta, v. 50, no. 12, 

p. 2833–2846.

Wood, S.A., 1992, Theoretical prediction of speciation and 

solubility of beryllium in hydrothermal solution to 300 °C 

at saturated vapor pressure—Application to bertrandite/

phenakite deposits: Ore Geology Reviews, v. 7, no. 4, 

p. 249–278.

World Health Organization, 2011, Guidelines for drinkingwater quality (4th ed): Geneva, Switzerland, World 

Health Organization, 541 p., accessed June 27, 2013,

at http://www.who.int/water_sanitation_health/

publications/2011/dwq_guidelines/en/index.html.

Wuensch, B.J., 1972, Beryllium, in Wedepohl, K.H., ed., 

Handbook of geochemistry, v. 2, pt. 1: Berlin, West 

Germany, Springer-Verlag, p. 4–A–8, 1–2.

E32 Critical Mineral Resources of the United States—Beryllium
Cobalt

By John F. Slack, Bryn E. Kimball, and Kim B. Shedd

Chapter F of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–F

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter F

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract.........................................................................................................................................................F1

Introduction...................................................................................................................................................F1

Cobalt Uses, Demand, and Availability of Supply ..........................................................................F1

Strategic and Critical Resource Issues ..........................................................................................F3

Geology..........................................................................................................................................................F4

Geochemistry.......................................................................................................................................F4

Mineralogy...........................................................................................................................................F4

Principal Deposit Types .....................................................................................................................F4

Stratiform Sediment-Hosted Cu-Co Deposits .......................................................................F4

Ni-Co Laterite Deposits .............................................................................................................F4

Magmatic Ni-Cu(-Co-PGE) Sulfide Deposits .........................................................................F5

Other Deposit Types ...........................................................................................................................F5

Black-Shale-Hosted Ni-Cu-Zn-Co Deposits ..........................................................................F5

Fe-Cu-Co Skarn and Replacement Deposits .........................................................................F5

Iron Oxide-Cu-Au(-Ag-U-REE-Co-Ni) Deposits .....................................................................F8

Metasedimentary-Rock-Hosted Co-Cu-Au Deposits ...........................................................F8

Mississippi Valley-Type Zn-Pb(-Co-Ni) Sulfide Deposits ....................................................F8

Polymetallic (Ag-Ni-Co-As-Bi) and Other Cobalt-Rich Vein Deposits ..............................F8

Volcanogenic Cu(-Zn-Co-Ag-Au) Massive Sulfide Deposits ..............................................F8

Sea-Floor Fe-Mn(-Ni-Cu-Co-Mo) Nodules ............................................................................F9

Sea-Floor Fe-Mn(-Co-Mo-REE) Crusts ...................................................................................F9

Sea-Floor Volcanogenic Cu(-Zn-Co-Ag-Au) Massive Sulfide Deposits ...........................F9

Resources and Production .......................................................................................................................F10

Production and Identified Resources ............................................................................................F10

Grade-Tonnage Characteristics ............................................................................................F10

Mining Methods .......................................................................................................................F13

Typical Mine Life ......................................................................................................................F14

Ore-Processing Methods .......................................................................................................F14

Undiscovered Resources ................................................................................................................F14

Unconventional Resources .............................................................................................................F14

Metamorphosed Black Shales ..............................................................................................F14

Modern Sea-Floor Volcanogenic Massive Sulfide Deposits ............................................F15

Ancient Ni-Co Laterites ..........................................................................................................F15

Ancient Fe-Mn(-Cu-Co-Ni) Nodules .....................................................................................F15

Ancient Cobalt-Rich Manganese Deposits .........................................................................F15

Ultramafic Igneous Rocks ......................................................................................................F15

Exploration for New Deposits ..................................................................................................................F15

Environmental Considerations .................................................................................................................F16

Sources and Fate in the Environment ...........................................................................................F16

Mine Waste Characteristics ...........................................................................................................F16

Human Health Concerns ..................................................................................................................F17

Ecological Health Concerns ............................................................................................................F18

Mine Closure......................................................................................................................................F19

Problems and Future Research ...............................................................................................................F19

Acknowledgments.....................................................................................................................................F20

References Cited........................................................................................................................................F20

Suggested citation:

Slack, J.F., Kimball, B.E., and Shedd, K.B., 2017, Cobalt, chap. F of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and 

Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and prospects 

for future supply: U.S. Geological Survey Professional Paper 1802, p. F1– F40, http://dx.doi.org/10.3133/pp1802F.
Cobalt

By John F. Slack, Bryn E. Kimball, and Kim B. Shedd

Abstract

Cobalt is a silvery gray metal that has diverse uses based 

on certain key properties, including ferromagnetism, hardness 

and wear-resistance when alloyed with other metals, low 

thermal and electrical conductivity, high melting point, multiple 

valences, and production of intense blue colors when combined 

with silica. Cobalt is used mostly in cathodes in rechargeable 

batteries and in superalloys for turbine engines in jet aircraft. 

Annual global cobalt consumption was approximately 

75,000 metric tons in 2011; China, Japan, and the United States 

(in order of consumption amount) were the top three cobaltconsuming countries. In 2011, approximately 109,000 metric 

tons of recoverable cobalt was produced in ores, concentrates, 

and intermediate products from cobalt, copper, nickel, platinumgroup-element (PGE), and zinc operations. The Democratic 

Republic of the Congo (Congo [Kinshasa]) was the principal 

source of mined cobalt globally (55 percent). The United States 

produced a negligible amount of byproduct cobalt as an 

intermediate product from a PGE mining and refining operation in southeastern Montana; no U.S. production was from 

mines in which cobalt was the principal commodity. China 

was the leading refiner of cobalt, and much of its production 

came from cobalt ores, concentrates, and partially refined 

materials imported from Congo (Kinshasa).

The mineralogy of cobalt deposits is diverse and includes 

both primary (hypogene) and secondary (supergene) phases. 

Principal terrestrial (land-based) deposit types, which represent most of world’s cobalt mine production, include primary 

magmatic Ni-Cu(-Co-PGE) sulfides, primary and secondary 

stratiform sediment-hosted Cu-Co sulfides and oxides, and 

secondary Ni-Co laterites. Seven additional terrestrial deposit 

types are described in this chapter. The total terrestrial 

cobalt resource (reserves plus other resources) plus past 

production, where available, is calculated to be 25.5 million 

metric tons. Additional resources of cobalt are known to 

occur on the modern sea floor in aerially extensive deposits 

of Fe-Mn(-Ni-Cu-Co-Mo) nodules and Fe-Mn(-Co-Mo-rareearth-element) crusts. Legal, economic, and technological 

barriers have prevented exploitation of these cobalt resources, 

which lie at water depths of as great as 6,000 meters, although 

advances in technology may soon allow production of these 

resources to be economically viable.

Environmental issues related to cobalt mining concern 

mainly the elevated cobalt contents in soils and waters. 

Although at low levels cobalt is essential to human health 

(it is the central atom in the critical nutrient vitamin B12), 

overexposure to high levels of cobalt may cause lung and heart 

dysfunction, as well as dermatitis. The ecological impacts of 

cobalt vary widely and can be severe for some species of fish 

and plants, depending on various environmental factors.

Introduction

Cobalt Uses, Demand, and Availability of Supply

Cobalt is a technologically important metal that has 

many diverse uses, including in batteries, superalloys, and 

cemented carbides and diamond tools (fig. F1). Globally, the 

leading use is in the manufacture of cathode materials for 

rechargeable batteries, primarily lithium-ion, nickel-cadmium, 

and nickel-metal-hydride batteries, which are used in 

consumer electronics, electric and hybrid-electric vehicles, 

energy storage units, and power tools. Superalloys are alloys 

developed for high-temperature service where relatively high 

mechanical stress is encountered and where surface stability 

is frequently required. The principal use for superalloys is in 

parts for turbine engines for jet aircraft and terrestrial energy 

generation. Cemented carbides, which are composite materials 

made of cobalt and tungsten carbide, are used as cutting tools 

and wear-resistant components by the metalworking, mining, 

oil and gas drilling, and construction industries. Diamond 

tools are similar to cemented carbides in that cobalt is the 

matrix that binds the wear-resistant particles together. Cobalt 

is used to make permanent and soft magnetic alloys. Cobaltbearing steels include high-speed steels for cutting tools 

and maraging steels, which are characterized by their great 

strength, toughness, and workability. Other cobalt-bearing 

alloys are characterized by their resistance to corrosion and 

(or) wear or by their particular thermal expansion properties. 

Additional chemical applications for cobalt include animal 
F2 Critical Mineral Resources of the United States—Cobalt

feed additives; bonding agents in steel-belted radial tires; 

catalysts for chemical, petroleum, and other industries; drying 

agents for paint; glass decolorizers; ground coat frits for 

porcelain enamels; humidity indicators; magnetic recording 

media; pigments; and vitamin B12.

In recent years, annual global cobalt consumption (shown 

as “apparent consumption” in fig. F2) has generally trended 

upward to approximately 75,000 metric tons. World apparent 

consumption is the sum of apparent consumption calculated 

for individual countries; the calculation for each country 

uses reported data for production, trade, and stock variations, 

where available. China, Japan, and the United States (in order 

of consumption amount) are the top three cobalt-consuming 

countries (Cobalt Development Institute and World Bureau of 

Metal Statistics, 2011; Cobalt Development Institute, 2012).

From 1950 to 2011, global mine production of cobalt 

increased by more than an order of magnitude (fig. F3). In 

2011, approximately 109,000 metric tons of recoverable 

cobalt was produced in ores, concentrates, and intermediate 

products from cobalt, copper, nickel, platinum-group-element 

(PGE), and zinc operations (fig. F2). More than one-half of 

the world’s cobalt was mined in the Democratic Republic 

of the Congo (Congo [Kinshasa]). Other leading countries 

where cobalt was mined are shown in figure F4 (Shedd, 

2013a). Trends in cobalt mineral exploration and supply are 

discussed in Wilburn (2012), Menzie and others (2013), and 

Shedd (2013a). Deposit types from which cobalt currently is 

mined are discussed in the Resources and Production section 

of this chapter. Table F1 (back of chapter) gives locations and 

grade-tonnage data for significant cobalt deposits of the world.

In 2011, the United States produced a negligible amount 

of byproduct cobalt as an intermediate product from Stillwater 

Mining Co.’s PGE mining and refining operation in southeastern Montana. Since then, minor amounts of byproduct 

cobalt in nickel concentrate have been produced from Lundin 

Mining Corp.’s underground Eagle Mine, which is a highgrade magmatic Ni-Cu(-Co-PGE) sulfide orebody located 

northwest of Marquette, Michigan, where mining began in 

2014. No U.S. production was from mines in which cobalt 

was the principal commodity. Other projects in the feasibility 

and development stages include Formation Metals Inc.’s 

underground cobalt mine to be constructed in a strata-bound 

Co-Cu-Au deposit in the Blackbird district (part of the Idaho 

cobalt belt) in Lemhi County and PolyMet Mining Corp.’s 

open pit mine to be constructed in the NorthMet Cu-Ni-CoPGE deposit in the Duluth Complex, which is a large mafic 

intrusive complex in northeastern Minnesota. Cobalt would be 

produced as a byproduct from the NorthMet project.

China was the leading refiner of cobalt in 2011, with 

much of its production coming from cobalt ores, concentrates, 

and partially refined materials imported from Congo 

(Kinshasa). Other significant sources of refined cobalt were 

Australia, Belgium, Canada, Congo (Kinshasa), Finland, 

Norway, and Zambia. Refined cobalt from Belgium and 

Finland was wholly or mainly produced from imported 

material, respectively (Shedd, 2013a).

Batteries

30%

Superalloys

19%

Cemented carbides

and diamond tools

13%

Catalysts

9%

Pigments and

other ceramic uses

9%

Magnetic alloys

7%

Hardfacing, steels,

and other alloys

5%

Drying agents for paints,

rubber adhesion promoters,

other carboxylate uses

5% Other chemical uses

3%

Cobalt Figure 1

Figure F1. Pie chart showing major end uses of cobalt as a 

percentage of consumption worldwide in 2011. Data are from the 

Cobalt Development Institute (2012).

0

50

100

150

2007 2008 2009

Year

2010 2011

Cobalt content, in thousand metric tons 

Mine production

Refinery production

Apparent consumption

Cobalt Figure 2

EXPLANATION

Figure F2. Bar chart showing world cobalt mine and refinery 

production and apparent consumption from 2007 to 2011. Mine 

production consists of the recoverable cobalt content of ores, 

concentrates, and intermediate products from cobalt, copper, 

nickel, platinum-group-element, and zinc operations. World 

apparent consumption is the sum of apparent consumption 

calculated for individual countries. The calculation for each 

country uses reported data for production, trade, and stock 

variations, where available. Apparent consumption data are from 

the Cobalt Development Institute and World Bureau of Metal 

Statistics (2011) and the Cobalt Development Institute (2012); 

production data are from Shedd (2013a).
Introduction F3

0 

20 

40 

60 

80 

100 

120 

1950 1960 1970 1980 

Years

1990 2000 2010 

Cobalt content, in thousand metric tons 

Cobalt Figure 3

Figure F3. Graph showing world cobalt mine production 

from 1950 to 2011. Production consists of the recoverable 

cobalt content of ores, concentrates, and intermediate 

products from cobalt, copper, nickel, platinum-groupelement, and zinc operations. Data are from U.S. Geological 

Survey (2013).

Cobalt Figure 4

Congo

(Kinshasa)

55%

Canada

6%

China

6%

Russia

6%

Zambia

5%

Cuba

4%

Australia

4%

Brazil

3%

New Caledonia

3% Morocco

2%

Other countries

6%

Figure F4. Pie chart showing percentage of world cobalt 

mine production in 2011, by country. The sources of production 

are cobalt, copper, nickel, platinum-group-element, and zinc 

operations. Data are from Shedd (2013a). Congo (Kinshasa) is a 

short-form name for Democratic Republic of the Congo.

Strategic and Critical Resource Issues

Cobalt is generally critical to the performance of the 

products in which it is used. Past periods of high prices and 

concern about availability have resulted in various efforts to 

conserve, reduce the use of, or substitute for cobalt. In many 

applications, further substitution would result in a loss in 

product performance and (or) increased cost. Cobalt’s diverse 

uses—particularly in parts for aircraft turbine engines; in 

numerous magnet applications, including marine propulsion 

systems, missile guidance systems, sensors, and radar; and 

in machine tools—make it important to the U.S. military and 

civilian industries.

That nearly all cobalt is mined as a byproduct of other, 

more abundant metals, such as nickel or copper, means that 

production is driven primarily by the markets for the principal 

metals, not by the need for cobalt. This situation limits 

producers’ flexibility in adjusting the amount of cobalt mined 

in response to changes in demand and can result in periods 

of oversupply or shortage. From 2009 to 2015, global cobalt 

production was higher than consumption, as production from 

new projects and from expansions to existing operations 

added to global supply. This trend led to a market surplus and 

downward pressure on prices (Shedd, 2013a, b; Searle, 2015; 

Spencer, 2016).

More than one-half of the world’s cobalt supply is 

mined in Congo (Kinshasa). This country has a high risk 

index for doing business owing to poor infrastructure, resource 

nationalism, a high perception of corruption, and a lack of 

transparency as well as wars during the 1990s to early 2000s, 

persistent tension in the eastern part of the country with 

substantial risk of civil war, and conflicts with neighboring 

Rwanda (Bedder, 2013; The Coface Group, undated). 

Although civil unrest in the eastern part of Congo (Kinshasa) 

has not affected the cobalt-producing areas, problems with 

infrastructure (particularly power and transportation) and 

reviews of and changes to mining contracts have slowed some 

of the potential growth in mine development and production. 

In spite of these issues, the copper-cobalt mining industry in 

Congo (Kinshasa) has significantly recovered from the collapse 

in production that took place in the 1990s (Shedd, 2013a).

The United States is highly reliant on imports for its 

cobalt needs. In recent years, approximately 75 to 80 percent 

of the U.S. cobalt supply has come from imports and 

releases from the National Defense Stockpile; the remaining 

20 to 25 percent has been from recycled scrap. This high 

reliance on imports increases the potential for supply 

disruption and high prices during supply shortfalls. The 

leading source of refined cobalt for the United States is

China (Shedd, 2013a, b).
F4 Critical Mineral Resources of the United States—Cobalt

Geology

Geochemistry

Cobalt is a silvery gray metal with diverse uses based 

on certain key properties, including hardness, wear-resistance 

when alloyed with other metals, low thermal and electrical 

conductivity, high melting point, multiple valences, and 

production of intense blue colors when combined with silica. 

Cobalt also is ferromagnetic, and it retains this property at 

the highest temperature of any metal. Although pure cobalt is 

not found in nature, cobalt-bearing minerals and compounds 

are numerous and widespread. The ionic radius of cobalt is 

0.72 angstroms (Å) for Co2+ and 0.63 Å for Co3+, both of 

which are similar to the ionic radii of Mg2+, Mn4+, Fe2+, Fe3+, 

and Ni2+, thus allowing substitution under favorable conditions for these elements by cobalt within many minerals and 

other phases. Among common rock types, the highest average 

cobalt content occurs in ultramafic igneous rocks, such as 

dunite and serpentinite, which contain, on average, 109 and 

115 parts per million (ppm) cobalt, respectively (table F2); 

mafic igneous rocks such as basalt contain, on average, about 

47 ppm cobalt. For comparison, Earth’s bulk continental crust, 

including all rock types, contains 29 ppm cobalt, on average. 

Lower cobalt contents characterize shale (which averages 

19 ppm), as well as granite and related felsic igneous rocks 

(which average 2 to 3 ppm) (table F2).

Mineralogy

The mineralogy of cobalt deposits is diverse and includes 

both primary (hypogene) and secondary (supergene) phases. 

In primary deposits, most cobalt is recovered from sulfide 

minerals, such as carrollite (Cu(Co,Ni)2S4), pentlandite 

((Fe,Ni,Co)9S8), linnaeite (Co3S4), and siegenite ((Co,Ni)3S4 ); 

arsenide minerals, such as skutterudite ((Co,Fe,Ni)As2-3) and 

safflorite ((Co,Fe)As2 ); and sulfarsenide minerals, including 

cobaltite (CoAsS) and glaucodot ((Co,Fe)AsS). Cobaltiferous

pyrite and pyrrhotite—(Fe,Co)S2

 and (Fe,Co)1-xS, respectively—

are mined in some deposits. Among secondary cobalt-rich 

phases, which form during surficial weathering, one of the 

most important historically was erythrite (Co3(AsO4)2 • 8H2O), 

which is a bright pink mineral also known as “cobalt bloom.” In

modern mining of secondary cobalt-rich deposits, the principal

economic phases, in addition to erythrite, are heterogenite 

(CoO(OH)) and asbolane ((Ni,Co)2–xMn(O,OH)4 • nH2O), plus 

local heazlewoodite ((Ni,Co)3S2), together with cobaltiferous 

oxyhydroxides, including goethite (Fe3+O(OH)), limonite 

(FeO(OH)•nH2O), and lithiophorite (AlMnO2(OH)2). With 

respect to the cobaltiferous oxyhydroxides, it is unclear from 

studies to date whether the contained cobalt is present in the 

crystal structure or adsorbed onto surfaces.

Principal Deposit Types

Cobalt deposits of economic or potential economic 

importance are diverse in terms of their geologic setting, age, 

morphology, mineralogy, geochemistry, origin, and gradetonnage relations (for example, Crockett and others, 1987; 

Smith, 2001; British Geological Survey, 2009). The principal 

deposit types that account for most of the world’s cobalt mine 

production are stratiform sediment-hosted Cu-Co deposits, 

Ni-Co laterite deposits, and magmatic Ni-Cu(-Co-PGE) 

sulfide deposits. These deposit types, together with selected 

examples, are described below.

Stratiform Sediment-Hosted Cu-Co Deposits

Most of the world’s cobalt is produced as a byproduct 

of copper mining in sediment-hosted Cu-Co deposits that 

form strata-bound and commonly stratiform zones within 

siliciclastic or carbonate strata (Hitzman and others, 2005). 

Important ore hosts in some deposits are giant breccias that 

originated by the dissolution of former salt beds. Globally, 

the deposits contain chalcopyrite, pyrite, and carrollite, plus 

minor amounts of bornite and chalcocite, in gangue material 

composed mainly of potassium feldspar, muscovite, biotite, 

albite, quartz, and carbonate. The ages of the mineralization 

are principally Neoproterozoic and Permian, but some important deposits are Mesoproterozoic. Although current genetic 

models differ in some respects, the consensus opinion is that 

metalliferous saline hydrothermal fluids were introduced at 

low to moderate temperatures during diagenesis and the early 

stages of deformation and metamorphism (see Zientek and 

others, 2013).

The greatest amount of contained cobalt in this 

deposit type occurs in the Central African Copperbelt in 

Congo (Kinshasa) and Zambia (Selley and others, 2005; Taylor 

and others, 2013). Large deposits in this region from which 

byproduct cobalt is currently or has been produced include 

Kamoto, KOV, and Tenke Fungurume, all in Congo (Kinshasa) 

(El Desouky and others, 2010; Fay and Barton, 2012), and 

Nkana in Zambia (Brems and others, 2009).

Ni-Co Laterite Deposits

Laterites are red regoliths that develop in humid tropical 

climates during the weathering of diverse types of bedrock. 

Those that developed on ultramafic bedrock may contain 

important Ni-Co deposits (Freyssinet and others, 2005; Butt 

and Cluzel, 2013). Ni-Co laterites locally contain abundant 

scandium and, rarely, elevated concentrations of PGEs. 

Ni-Co laterite deposits consist of the following layers, from 

top to bottom: overburden, limonite, saprolite, and weathered 

ultramafic source rocks. Ore zones of laterite deposits range 

in thickness from about 10 meters (m) to as much as 40 m in 

some cases, and generally contain more than 1 percent nickel 

and less than 0.15 percent cobalt. Limonite developed over 
Geology F5

ultramafic rocks tends to have higher cobalt grades, whereas 

saprolite has higher nickel grades. Metal accumulation involves 

supergene processes and several key variables, including 

primary bedrock lithology, climate history, topography, and 

structural preparation (that is, the fracture or joint density of 

bedrock). Ages of the deposits typically are mid-Tertiary to 

Holocene; some fossil Ni-Co laterites are known. Major ore 

constituents include the following: (a) nickeliferous serpentine 

(garnierite), talc, and chlorite; (b) nickel- or cobalt-bearing 

clays (for example, nontronite, and montmorillonite); and (c) 

erythrite, heterogenite, asbolane, heazlewoodite, and millerite; 

and (d) goethite, limonite, and lithiophorite. The gangue 

(non-ore) material may include various amounts of quartz, 

amorphous silica, clays, and serpentine.

The Kalgoorlie and the Murrin Murrin deposits in 

Western Australia (Elias and others, 1981; Gaudin and others, 

2005) and the Goro deposit in New Caledonia (Wells and 

others, 2009) are examples of some of the world’s many large 

Ni-Co laterite deposits. The manganiferous Nkamouna Co-Ni 

deposit in Cameroon is one of the few laterites known for 

which cobalt would be the principal economic metal to be 

mined (Lambiv Dzemua and Gleeson, 2012).

Magmatic Ni-Cu(-Co-PGE) Sulfide Deposits

Large resources (including reserves) of cobalt are 

contained in Ni-Cu(-Co-PGE) sulfide deposits hosted in mafic 

and ultramafic igneous rocks (Naldrett, 2004; Eckstrand and 

Hulbert, 2007). This deposit type comprises semimassive to 

massive sulfides that occur within or near the basal zones of 

layered intrusive complexes, in discordant magmatic conduits, 

and within ultramafic intrusions and lava flows. The ages of 

the deposits, which approximate those of the host intrusions, 

range from Archean to Tertiary. Nickel is the principal metal 

commodity, and it is accompanied by subequal proportions 

of copper in most deposits; cobalt and PGEs are mining 

byproducts. Cobalt resides mainly in cobaltiferous pentlandite 

and, to a lesser extent, in linnaeite. The gangue minerals 

consist mostly of primary magmatic minerals, such as olivine, 

pyroxene, and plagioclase. Mineralizing processes involve 

magmatic segregation of sulfides and, in some deposits, 

hydrothermal mobilization into post-magmatic structures.

Deposits of this type that currently feature significant 

byproduct production of cobalt include Voisey’s Bay in 

Newfoundland and Labrador, Canada (Naldrett and Li, 2007); 

the Sudbury district in Ontario, Canada (Ames and Farrow, 

2007); and the Norilsk-Talnakh district in Siberia, Russia 

(Naldrett and others, 1996).

Other Deposit Types

Types of deposits that on a global scale have produced 

relatively minor amounts of cobalt include the following: 

(a) black-shale-hosted Ni-Cu-Zn-Co, (b) Fe-Cu-Co skarn 

and replacement, (c) iron oxide-Cu-Au(-Ag-U-REE-Co-Ni), 

(d) metasedimentary-rock-hosted Co-Cu-Au, (e) Mississippi 

Valley-type Zn-Pb(-Co-Ni), (f) polymetallic (Ag-Ni-CoAs-Bi) and other cobalt-rich veins, and (g) volcanogenic 

Cu(-Zn-Co-Ag-Au) massive sulfides. Identified cobalt deposits 

on the sea floor, which may be mined in the future, include 

Fe-Mn(-Ni-Cu-Co-Mo) nodules, Fe-Mn(-Co-Mo-REE) crusts, 

and volcanogenic Cu(-Zn-Co-Ag-Au) massive sulfides.

Black-Shale-Hosted Ni-Cu-Zn-Co Deposits

Black shales are well known for containing elevated 

contents of many metals of economic interest, including Cu, 

Mo, Ni, Zn, Co, Cd, Ag, Au, Se, Cr, V, U, and PGEs (for 

example, Desborough and Poole, 1983; Coveney, 2003). 

Metals concentrated in black shales may reside in pyrite; 

organic matter; aluminosilicate minerals, such as illite; and 

locally in sphalerite and chalcopyrite. Some deposits contain 

very high contents of molybdenum, nickel, zinc, and (or) 

vanadium within primary sedimentary beds or laminae. In 

other deposits, high contents of cobalt formed by hydrothermal 

leaching, mobilization, and concentration into sulfide minerals 

during deformation and regional metamorphism.

One of the few black-shale-hosted deposits from which 

byproduct cobalt is recovered is the giant Talvivaara orebody 

in central Finland, which is being mined for nickel, zinc, and 

copper (Loukola-Ruskeeniemi and Lahtinen, 2013).

Fe-Cu-Co Skarn and Replacement Deposits

Pluton-related skarn and replacement deposits form 

by the introduction of hydrothermal fluids into chemically 

reactive rocks, mainly carbonate (limestone and dolostone), 

and by metasomatic processes that introduce metals and other 

components into the precursor strata. These deposits occur 

proximal or distal to intrusive bodies and contain a diverse 

suite of metals (Megaw, 1998; Meinert and others, 2005). 

Sulfide minerals may include pyrite, pyrrhotite, chalcopyrite, 

sphalerite, and galena, together with abundant magnetite 

in some deposits. Cobaltiferous deposits generally contain 

mainly copper residing in chalcopyrite; cobalt occurs in 

cobaltite or cobalt-rich pyrite. In replacement deposits, which 

commonly are located distal from pluton contacts, gangue 

minerals typically include quartz or carbonate. The gangue 

minerals of skarn deposits, which occur near plutons, are 

generally different and include garnet, pyroxene, amphibole, 

epidote, olivine, plagioclase, and (or) scapolite.

The Cornwall and the Morgantown (Grace Mine) Fe-Cu-Co

deposits, which are two of only a few large cobaltiferous skarn 

deposits in the world, are located in southeastern Pennsylvania 

(Lapham, 1968) but are no longer operating. Another large 

deposit of this type, although it lacks significant amounts of 

iron, is Ruby Creek in northwestern Alaska (Bernstein and 

Cox, 1986). The Mount Elliott Cu-Au(-Co-Ni) skarn deposit 

in Queensland, Australia, differs in that it contains a significant 

amount of gold (Wang and Williams, 2001).
F6 Critical Mineral Resources of the United States—Cobalt

Table F2. Cobalt concentrations in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency; cm, centimeter; ppm, part per million; μg/g, microgram per gram; 

μg/L, microgram per liter; μm, micrometer; ng/m3, nanogram per cubic meter]

Environment

and (or) location

Cobalt

concentration Unit Notes Reference(s)

Rocks

Upper continental crust 10 ppm Average Taylor and McLennan (1995)

Bulk continental crust 29 ppm Average Taylor and McLennan (1995)

Lower continental crust 35 ppm Average Taylor and McLennan (1995)

Basalt 47 ppm average Krauskopf and Bird (1995)

Dunite (alpine-type) 109 ppm Average Gülaҫar and Delaloye (1976)

Granite 2.4 ppm Average Krauskopf and Bird (1995)

Pyroxenite (alpine-type) 55.2 ppm Average Gülaҫar and Delaloye (1976)

Serpentinite (alpine-type) 115 ppm Average Gülaҫar and Delaloye (1976)

Shale 19 ppm Average Krauskopf and Bird (1995)

Soils

United States 5 to 30 ppm Regional background soil 

in Idaho

Shacklette and Boerngen (1984)

United States 29 to 940 ppm Near mining in the Idaho 

copper belt

Giles and others (2009)

Conterminous United States 0.5 to 160 ppm 0 to 5 cm depth; median is 

7.2 ppm

Smith and others (2005)

Conterminous United States 1.8 to 14 ppm O horizon, if present; median 

is 3 ppm

Smith and others (2005)

Conterminous United States 0.9 to 143 ppm A horizon; median is 7.1 ppm Smith and others (2005)

Conterminous United States 0.7 to 191 ppm C horizon; median is 7.8 ppm Smith and others (2005)

Western United States 7.1 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Eastern United States 5.9 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Canada 2 to 6 ppm 0 to 5 cm; distant from 

Sudbury mining district

Narendrula and others (2012)

Canada 22 to 37 ppm 0 to 5 cm; near Sudbury 

mining district

Narendrula and others (2012)

Canada 5 to 18 ppm 0 to 5 cm; distant from roast beds Hutchinson and Symington (1997)

Canada 58 to 299 ppm 0 to 5 cm; near roast beds Hutchinson and Symington (1997)

Congo (Kinshasa) 17 to 33 ppm 0 to 5 cm; distant from 

African copper belt mines

Narendrula and others (2012)

Congo (Kinshasa) 204 to 6,150 ppm 0 to 5 cm; near African 

copper belt mines

Narendrula and others (2012)

Egypt 13 to 25 ppm 0 to 20 cm; distant from industry Zohny (2002)

Egypt 26 to 65 ppm 0 to 20 cm; pollution from 

industry

Zohny (2002)

Sweden 0.5 to 2.3 ppm Range of profile developed on 

quartzite and gniess

Tyler (2004)

Proposed DOE benchmark 20 ppm Contaminant screening benchmark for terrestrial plantsEfroymson and others (1997)

Soil-quality guideline 40 ppm Canadian agricultural 

soil guideline

Canadian Council of Ministers

of the Environment (2013)
Geology F7

Table F2. Cobalt concentrations in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency; cm, centimeter; ppm, part per million; μg/g, microgram per gram; 

μg/L, microgram per liter; μm, micrometer; ng/m3, nanogram per cubic meter]

Environment 

and (or) location

Cobalt 

concentration Unit Notes Reference(s)

Waters

Seawater, South Atlantic and 

Southern Oceans

Seawater, Sargasso Sea

African rivers

European rivers

North American rivers

South American rivers

Huanghe River, China

Streamwater, Canada—

Cobalt

Streamwater, Sweden 

(northern)

Streamwater, Sweden 

(southern)

Streamwater, United States—

Cobalt

Streamwater, United States—

Cobalt

Sediment, world rivers

Tailings pore water, 

Khovu-Aksy, Russia

Proposed EPA benchmark

Proposed EPA benchmark

0.0003 to 0.004

0.001 to 0.004

0.04 to 0.43

0.08 to 0.26

0.02 to 0.15

0.02 to 0.18

0.006 to 0.03

3 to 20

0.057

0.24

0.67 to 30

10 to 1,100

22.5

70 to 61,000

1,500

23

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

μg/g

μg/L

μg/L

μg/L

Dissolved (<0.22 μm)

Dissolved and colloidal

(<0.4 μm)

Dissolved load (<0.2 μm)

Dissolved load (<0.2 μm)

Dissolved load (<0.2 μm)

Dissolved load (<0.2 μm)

Dissolved load (<0.2 μm)

Dissolved load (<0.45 μm); 

mine drainage

Median 

Median

Samples distant from mines

Samples near mines

Average suspended sediment

Median is 540 μg/L

Tier II secondary acute value

Tier II secondary chronic value

Bown and others (2012)

Shelley and others (2012)

Gaillardet and others (2003)

Gaillardet and others (2003)

Gaillardet and others (2003)

Gaillardet and others (2003)

Gaillardet and others (2003)

Kwong and others (2007)

Huser and others (2011)

Huser and others (2011)

Eppinger and others (2003)

Eppinger and others (2003)

Viers and others (2009)

Bortnikova and others (2012)

Suter and Tsao (1996)

Suter and Tsao (1996)

Air

Worldwide

South Pole

Greenland

0.13 to 37

0.1 to 1.2

70 to 150

ng/m3

ng/m3

ng/m3

“Polluted air”; median 

is 3 ng/m3

None

None

Reimann and de Caritat (1998)

Kabata-Pendias and Pendias 

(2001)

Kabata-Pendias and Pendias 

(2001)
F8 Critical Mineral Resources of the United States—Cobalt

Iron Oxide-Cu-Au(-Ag-U-REE-Co-Ni) Deposits

The iron oxide-Cu-Au(-Ag-U-REE-Co-Ni) class of 

mineral deposits is globally important as a major source of 

copper, gold, and in some cases, silver, uranium, and rareearth elements (REEs) (Williams and others, 2005). These 

iron oxide-copper-gold (IOCG) deposits are mainly of Late 

Archean, Proterozoic, and Mesozoic ages, and they occur 

either near plutons or in metamorphic terranes without a clear 

link to intrusive bodies. Their ore mineralogy is dominated 

by magnetite, hematite, chalcopyrite, bornite, and gold, with 

concentrations in some deposits of uraninite, bastnäsite (or 

bastnaesite), monazite, cobaltite, glaucodot, bismuthinite, 

native bismuth, arsenopyrite, millerite, molybdenite, and 

galena; common gangue minerals are fluorite, barite, and 

quartz; and tourmaline is present locally. This deposit type 

is thought to have originated from one of the following fluid 

types and sources: magmatic-hydrothermal, metamorphic, 

and evaporitic brine.

Very large IOCG deposits that contain appreciable 

amounts of cobalt include, in Australia, the Olympic Dam 

deposit in South Australia (Reynolds, 2000) and the Ernest 

Henry deposit in Queensland (Mark and others, 2000), and 

in Brazil, the Sossego deposit (Monteiro and others, 2008).

Metasedimentary-Rock-Hosted Co-Cu-Au Deposits

Co-Cu-Au deposits hosted in metasedimentary rocks are 

strata-bound zones of semimassive to locally massive sulfides 

within deformed and metamorphosed siliciclastic strata chiefly 

of Proterozoic age (Slack, 2013). The mineralized zones range 

from strata-bound and discordant to stratiform and include 

lenses, veins, and breccias. The deposits consist of cobaltite 

and (or) other cobalt-rich sulfarsenide or sulfide minerals, or 

cobaltiferous pyrite, together with chalcopyrite and magnetite, 

in a gangue of quartz, muscovite, biotite, chlorite, potassium 

feldspar, albite, and (or) scapolite, accompanied in many 

deposits by minor amounts of tourmaline. The origins of these 

deposits are thought to be varied; a range of mineralizing 

processes, from diagenetic to epigenetic (the latter occurring 

both before and during metamorphism) are thought to be 

involved. In some deposits, geochronological and geochemical 

evidence suggests links to granitic and (or) gabbroic plutons, 

but the origin of most deposits of this type—and sources of the 

hydrothermal fluids and metals—are enigmatic.

The largest tonnages of cobalt in this deposit type are 

contained in the Blackbird district of east-central Idaho (Slack, 

2013, and references therein), the Kuusamo belt of northeastern Finland (Vanhanen, 2001), and the NICO deposit in 

the Northwest Territories in Canada (Goad and others, 2000). 

Cobalt deposits of this type in the Modum district in Norway, 

especially those at the Skuterud Mine (Grorud, 1997), were 

historically important as the major European source of blue 

pigment used during the 17th and 18th centuries.

Mississippi Valley-Type Zn-Pb(-Co-Ni) 

Sulfide Deposits

Strata-bound Zn-Pb(-Co-Ni) sulfide concentrations 

hosted by carbonate strata are termed Mississippi Valley-type 

(MVT) deposits (Leach and others, 2005; Paradis and others, 

2007). These deposits typically consist of various proportions 

of sphalerite and galena in a gangue of fluorite or barite; 

quartz is uncommon. The principal cobalt mineral is carrollite. 

Most deposits of this type are Phanerozoic, although a few 

Proterozoic examples are known. Mineralization characteristically involves the migration of low-temperature, highly saline 

brines and the subsequent precipitation of ore and gangue 

minerals as open-space fillings of paleokarst structures or as 

replacement zones in carbonate rocks.

MVT deposits generally lack high cobalt contents, but a 

few, such as the Mine La Motte-Fredericktown and the Higdon 

deposits in Missouri (Seeger, 2008; Parra and others, 2009) 

have produced minor amounts of cobalt as a byproduct of the 

mining of lead and zinc.

Polymetallic (Ag-Ni-Co-As-Bi) and 

Other Cobalt-Rich Vein Deposits

Cobalt-rich vein deposits fill fractures and faults within 

metasedimentary or metaigneous rocks of Proterozoic or 

younger age (Kissin, 1992). The majority of these deposits 

were mined for silver and cobalt; some also produced 

byproduct copper, lead, zinc, gold, uranium, and (or) barite. In 

these types of deposits, multistage mineralization is common, 

typified by a diverse ore mineralogy consisting of native silver 

and native bismuth, gold, argentite, rammelsbergite, safflorite, 

skutterudite, gersdorffite, niccolite, uraninite, arsenopyrite, 

pyrite, pyrrhotite, sphalerite, galena, bornite, chalcopyrite, 

tetrahedrite, and Sb-As-Ag sulfosalts. Gangue minerals may 

include quartz, carbonate, fluorite, and (or) barite. The origins 

of these deposits vary widely and include mineralization by 

magmatically derived hydrothermal fluids, metamorphic 

fluids, basinal brines, and meteoric waters.

Economically, the most important vein deposits have 

been those in the Erzgebirge region of the Czech Republic and 

Germany (Hermann, 2005), the Kongsberg district of southern 

Norway (Neumann, 1944), the Cobalt district of Ontario, 

Canada (Marshall and Watkinson, 2000), and the Bou Azzer 

district of Morocco (Bouabdellah and others, 2016).

Volcanogenic Cu(-Zn-Co-Ag-Au) Massive 

Sulfide Deposits

Volcanogenic massive sulfide (VMS) deposits occur in 

marine volcanic and volcanosedimentary belts of Archean 

to Holocene age and are mined principally for copper, zinc, 

lead, silver, and gold (Galley and others, 2007; Shanks and 
Geology F9

Thurston, 2012). These deposits precipitate from hydrothermal fluids on the sea floor within irregular chimneys and 

mounds, and in the subsurface within strata-bound or discordant zones. The major sulfide minerals are pyrite, pyrrhotite, 

chalcopyrite, and sphalerite; some deposits have appreciable 

amounts of galena, tetrahedrite, and gold. Gangue minerals 

may include quartz, muscovite, chlorite, siderite, and barite. 

Most VMS deposits lack elevated contents (> 0.1 weight 

percent) of cobalt but several have higher grades, especially 

deposits hosted predominantly by ultramafic or mafic volcanic 

rocks. Extensive research on this deposit type has allowed 

the development of refined genetic models that involve 

subsea-floor emplacement of igneous magma that drives 

hydrothermal convection and seawater entrainment, coupled 

with upflow of metalliferous hydrothermal fluids and subsequent sulfide precipitation (see Shanks and Thurston, 2012, 

and references therein).

Among ancient deposits, the few that have significant 

amounts of contained cobalt include Outukumpu in central 

Finland (Peltonen and others, 2008), Windy Craggy in 

northwestern British Columbia, Canada (Peter and Scott, 

1999), and Deerni in Qinghai Province, China (Hou and 

others, 1999).

Sea-Floor Fe-Mn(-Ni-Cu-Co-Mo) Nodules

Ferromanganese nodules on the modern sea floor 

contain large resources of several metals, including nickel, 

copper, cobalt, and molybdenum (Hein and others, 2013). 

These nodules occur mainly on abyssal plains at depths of 

3,500 to 6,500 m, having formed by the precipitation of iron 

and manganese oxyhydroxide phases from seawater and 

pore fluids within enclosing sediments. The growth rates 

of nodules vary greatly, from a few millimeters per million 

years, where components are precipitated only from seawater, 

to as much as 250 millimeters per million years, where 

precipitation is dominantly from pore fluids. Nodule diameters 

typically are 2 to 8 centimeters (cm). Major constituents are 

delta-manganese dioxide (δ-MnO2), todorokite, and other 

similar manganates, with lesser amounts of birnessite, iron 

oxyhydroxides, and detrital aluminosilicate minerals. Metals 

of economic interest, such as cobalt, are sorbed onto surfaces 

of the manganese oxyhydroxides, where they are retained 

by surface oxidation processes. The formational ages of the 

nodules generally are no older than the Holocene owing to 

dissolution of manganese oxides within anoxic pore fluids 

during diagenesis, but ancient nodules of Miocene and 

Cretaceous age also are known.

The best-documented and largest nodule fields are 

located in the Pacific Ocean between the Clarion and 

Clipperton fracture zones, in the Cook Islands Exclusive 

Economic Zone (EEZ), in the Penrhyn-Samoa Basin, the Peru 

Basin, and in the Pioneer area of the central Indian Ocean 

(Hein and others, 2013).

Sea-Floor Fe-Mn(-Co-Mo-REE) Crusts

Ferromanganese crusts on the modern sea floor also 

represent immense resources of cobalt, molybdenum, 

REEs, and other metals (Glasby and others, 2010; Hein and 

others, 2013). These crusts occur on seamounts, plateaus, 

and spreading ridges at relatively shallow depths of 

800 to 3,000 m. Growth rates of the crusts are extremely slow, 

generally in the range of 1 to 6 millimeters per million years 

owing to precipitation of components only from seawater. The 

thickness of the crust varies from less than 1 cm to as much as 

26 cm, and the thickest parts occur on the oldest seamounts. 

The principal minerals are δ-MnO2, iron oxyhydroxides, and 

carbonate fluorapatite, together with local goethite and minor 

amounts of detrital minerals, such as quartz and feldspar. The 

incorporation of cobalt and other metals occurs by processes 

similar to those that take place in ferromanganese nodules.

The greatest resources in ferromanganese crusts are 

located in the Pacific prime crust zone of the central equatorial 

Pacific Ocean (Hein, 2002). Other regions contain appreciable 

resources, however, such as in the Nameless, Unicorn, and 

Maderia-Tore Rise (MTR) zones of the northeastern Atlantic 

Ocean (Muiños and others, 2013), and the Afanasiy-Nikitin 

seamount in the Indian Ocean (Parthiban and Banakar, 1999).

Sea-Floor Volcanogenic Cu(-Zn-Co-Ag-Au) 

Massive Sulfide Deposits

More than 100 VMS deposits have been discovered 

on the modern sea floor (Hannington and others, 2005). 

Cobaltiferous examples are few and limited to occurrences in 

ultramafic rocks. These relatively cobalt-rich VMS deposits 

are present in serpentinite, which is a rare lithology on the 

modern sea floor, having been exposed by detachment faulting 

and core complex formation that raise deeper rocks of the 

lower oceanic crust and upper mantle to the ocean floor. 

The major sulfide minerals present in ancient VMS deposits 

hosted in basalt are chalcopyrite, isocubanite, pyrite (locally 

cobaltiferous), sphalerite, and wurtzite; the gangue minerals 

consist of abundant serpentine and smectite, in addition 

to chlorite.

The Rainbow vent field on the Mid-Atlantic Ridge 

(Bogdanov and others, 2002) has VMS deposits that contain 

the greatest amount of cobalt known among modern sea-floor 

VMS deposits. Other deposits of this type that are found on 

the Mid-Atlantic Ridge occur in the Logatchev (Mozgova 

and others, 1999) and the Nibelungen fields (Melchert and 

others, 2008).
F10 Critical Mineral Resources of the United States—Cobalt

Resources and Production

Production and Identified Resources

Figure F5 shows cobalt production in 2011, by deposit 

type. Cobalt production from Congo (Kinshasa) and most 

of the output from Zambia was as a byproduct of the copper 

mining of stratiform sediment-hosted Cu-Co deposits. Cobalt 

mine production from most other countries was as a byproduct 

of nickel mining. Production from Cuba and New Caledonia 

was from Ni-Co laterite deposits. Production from Canada 

and Russia and most of the production from China was from 

magmatic Ni-Cu(-Co-PGE) sulfide deposits. Production from 

Australia and Brazil was from Ni-Co laterite and magmatic 

Ni-Cu(-Co-PGE) sulfide deposits. Only in Morocco was 

cobalt produced as the principal commodity from a current 

mining operation; the deposits in Morocco are hydrothermal 

polymetallic veins.

The proportion of cobalt produced from different deposit 

types has varied over time. During the 1990s, production from 

stratiform sediment-hosted Cu-Co deposits was relatively low 

owing to a period of civil unrest in Congo (Kinshasa) and 

a major collapse at the Kamoto Mine in Congo (Kinshasa) 

in 1990. The percentage of production from Ni-Co laterite 

deposits has increased since the mid-1990s owing to development of second-generation hydrometallurgical processes for 

refining those ores and investment by a Canadian company in 

one of the laterite operations in Cuba.

Grade-Tonnage Characteristics

The grade-tonnage plot in figure F6, which combines all 

types of cobalt-bearing mineral deposits and districts, shows 

a large range in the amounts of contained cobalt. In the figure, 

several to many deposits are grouped into districts for some 

areas (such as Sudbury, Ontario, Canada, and Norilsk-Talnakh, 

Russia) because, in these cases, grade and tonnage data for the 

individual deposits were not available. The sizes of the plotted 

deposits and districts, which include past production, reserves, 

and other resources (where known) are given in table F1. The 

summation of reserves and other resources has the potential 

to introduce double-counting when data are from reports in 

which reserves are included as part of resources. In table F1, 

where possible, reserves were not included when the potential 

for such double-counting was identified. Past production is 

included to give an indication of the original size of the deposit

before mining. For many deposits—especially stratiform 

sediment-hosted Cu-Co, Ni-Co laterites, and magmatic 

Ni-Cu(-Co-PGE) sulfides—cumulative past production is

small compared with the amount of cobalt contained in 

reserves and other resources. Figure F7 shows the global 

distribution of the largest cobalt deposits on land and on the sea 

floor; the terrestrial deposits shown in the figure also include 

selected small deposits that represent minor deposit types.

The total terrestrial cobalt resource (plus past production, 

where available) calculated from data in table F1 (back of 

chapter) is 25.5 million metric tons, which is very similar 

to the terrestrial cobalt resource of 26.1 million metric tons 

presented in a recent independent study by Mudd and others 

(2013). Among terrestrial deposits, contained cobalt can vary 

by three orders of magnitude or more (fig. F6). Most types of 

cobalt deposits contain between 5,000 and 500,000 metric tons 

of cobalt. The largest tonnages (greater than 500,000 metric 

tons of contained Co) are present in numerous magmatic 

Ni-Cu(-Co-PGE) sulfide deposits (some grouped into 

districts, such as Sudbury), several stratiform sediment-hosted 

Cu-Co deposits, and numerous lateritic Ni-Co deposits. 

Similarly, cobalt grades differ greatly—mostly low grades 

(0.01 to 0.2 percent cobalt) for magmatic Ni-Cu(-Co-PGE) 

sulfide deposits, intermediate grades (0.03 to 0.2 percent 

cobalt) for the majority of lateritic Ni-Co deposits, and 

relatively high grades (0.03 to 1.0 percent cobalt) for most 

stratiform sediment-hosted Cu-Co deposits. The highest 

average grades (about 1.5 percent cobalt) are in relatively 

small cobalt-rich vein deposits of the Bou Azzer district in 

Morocco. As shown in figures F6 through F8, the largest 

amount of cobalt on land is contained within stratiform 

sediment-hosted Cu-Co deposits (41 percent) and lateritic 

Ni-Co deposits (36 percent), followed at much lower 

proportions by that in magmatic Ni-Cu(-Co-PGE) sulfide 

deposits (15 percent) and other terrestrial deposits (8 percent), 

of which the greatest portion is in VMS deposits (2.7 percent).

Other terrestrial 2%

Cobalt Figure 5

Stratiform

sediment-hosted

Cu-Co

60%

Magmatic

Ni-Cu(-Co-PGE) 

sulfide

23%

Ni-Co

laterite

15%

Figure F5. Pie chart showing percentage of world cobalt 

mine production in 2011, by deposit type. Other terrestrial 

deposits are grouped together and include black-shale-hosted 

Ni-Cu-Zn-Co deposits, polymetallic (Ag-Ni-Co-As-Bi) and 

other cobalt-rich veins, and volcanogenic Cu(-Zn-Co-Ag-Au) 

massive sulfide deposits. No recoverable cobalt was known 

to be produced in 2011 from deposit types not listed. The 

sources of production are cobalt, copper, nickel, platinumgroup-element (PGE), and zinc operations. Ag, silver; 

As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; 

Ni, nickel; Zn, zinc
Resources and Production F11

108

 t Co

107

 t Co

106

 t Co

105

 t Co

104

 t Co

103

 t Co

EXPLANATION

Cobalt deposit type

Metasedimentary-rock-hosted Co-Cu-Au

Fe-Cu-Co skarn and replacement

Polymetallic (Ag-Ni-Co-As-Bi) and 

 other cobalt-rich vein

Ni-Co laterite

Sea-floor Fe-Mn(-Ni-Cu-Co-Mo) nodules

Sea-floor Fe-Mn(-Co-Mo-REE) crusts

BA

BB MU

KF

SC

KM MTR

TF WC

NT SB

TV

TM

DM

MS

KN

NO

OD

NM

JC

PC

CI

CC

Cobalt Figure 6

Magmatic Ni-Cu(-Co-PGE) sulfide

Stratiform sediment-hosted Cu-Co

Black-shale-hosted Ni-Cu-Zn-Co

Iron oxide-Cu-Au(-Ag-U-REE-Co-Ni)

Volcanogenic Cu(-Zn-Co-Ag-Au) 

 massive sulfide

0.001

0.01

0.1

1.0

10

Grade, in weight percent cobalt

1 10 100 1,000 10,000 100,000

Deposit tonnage, in million metric tons

Figure F6. Grade-tonnage plot for 214 cobalt deposits worldwide. Grades and tonnages 

include production plus reserves plus other resources where known; reserve and 

resource data are from publicly available reports and company Web sites (see table F1), 

but in some cases are not defined by a National Instrument 43–101 standard, Joint Ore 

Reserves Committee code, or similar mineral-resource classification scheme. Labeled 

deposits represent most of those containing more than 500,000 metric tons (t) of cobalt, 

many of the U.S. deposits, some examples of the less common deposit types, and some 

that are discussed in the text. Small deposits—those with less than 1,000 metric tons of 

cobalt—are not shown. Diagonal lines are isolines of contained cobalt, in metric tons. 

Abbreviations: BA, Bou Azzer (Morocco); BB, Blackbird (Idaho); CC, Clarion-Clipperton 

zone (Pacific Ocean); CI, Cook Islands Exclusive Economic Zone (Pacific Ocean); 

DM, Dumont (Canada); JC, Jacaré (Brazil); KF, Kisanfu (Congo [Kinshasa]); KM, Kamoto–

KOV–Musonoi–Mupine deposits (Congo [Kinshasa]); KN, Kalgoorlie Nickel (Australia); MS, 

Mesaba (Minnesota); MTR, Maderia-Tore Rise (Atlantic Ocean); MU, Mutanda (Congo 

[Kinshasa]); NM, Nkamouna (Cameroon); NO, Northmet (Minnesota); NT, Noril’sk Talnakh 

(Russia); OD, Olympic Dam (Australia); PC, Pacific prime crust zone (Pacific Ocean); SB, 

Sudbury (Canada); SC, Sheep Creek (Montana); TF, Tenke Fungurume (Congo [Kinshasa]); 

TM, Twin Metals (Minnesota); TV, Talvivaara (Finland); WC, Windy Craggy (Canada)
F12 Critical Mineral Resources of the United States—Cobalt

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°

50°

40°

30°

20°

10°

0°

10°

20°

30°

40°

50°

80°

70°

70°

80°

60°

Blackbird

Sudbury

NICO

Windy Craggy

Moa

Jacaré

Nkamouna

Mukondo

Clarion-Clipperton

zone

Cook

 Islands

 Exclusive

 Economic

 Zone

Pioneer

Olympic Dam

PPCZ

MaderiaTore Rise

Bou Azzer

Noril’sk-Talnakh

Magnitogorsk Talvivaara

AfnasiyNikitinseamount Kisanfu

Tenke Fungurume Mutanda DIMA

Kamoto-KOVMusonoi-Mupine

PPCZ

Cobalt figure 7. 

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Terrestrial (land-based) cobalt deposit types Sea-floor cobalt deposit types

EXPLANATION

≥1.0 Mt cobalt

SSH Cu-CoMagmatic Ni-Cu (-Co-PGE) sulfide

0.5 to 0.99 Mt cobalt

SSH Cu-CoNi-Co lateriteMagmatic Ni-Cu (-Co-PGE) sulfide

Iron oxide-Cu-Au (-Ag-U-REE-Co-Ni)

<0.5 Mt cobalt

Fe-Cu-Co skarn and replacementMSRH Co-Cu-Au

NodulesCrustsClarion-Clipperton zonePacific prime crust zone (PPCZ)

SSH Cu-CoNi-Co lateriteMagmatic Ni-Cu (-Co-PGE) sulfideBSH Ni-Cu-Zn-CoPolymetallic vein

>0.1 Mt cobalt

Volcanogenic  Cu(-Zn-Co-Ag-Au) massive sulfide

etric tons of cobalt) and selected 

Additional data on these and

el; SSH, 

orld map showing global distribution of major cobalt-bearing mineral deposits (those containing equal to or more than 500,000 m

represent minor types. The sizes of the symbols for terrestrial (land-based) deposits reflect the amounts of contained cobalt. 

ore Rise, which is located west of other the deposits are given in table F1. Note that the Nameless and Unicorn sea-floor deposits (table F1) are within the area of the Maderia-T

Au, gold; BSH, black-shale-hosted; Co, cobalt; Cu, copper; MSRH, metasedimentary-rock-hosted; Mt, million metric tons; Ni, nick .)3102 ,srehto dna 

 t

s

a

o

ht

ñi

st

u M

W

iso

e

p

e

Figure F7.

e

s(

d

 l

 r

a

e

g

ll

u

a

tr

m

o

sPstratiform sediment-hosted; Zn, zinc
Resources and Production F13

A. Terrestrial deposits B. Terrestrial and sea-floor deposits 

Stratiform sediment-hosted Cu-Co 7% 

Magmatic Ni-Cu(-Co-PGE) 3% 

Ni-Co laterite 6% 

Other terrestrial 1% 

Sea-floor

Fe-Mn(-Ni-Cu-Co-Mo)

nodules

45% 

Sea-floor

Fe-Mn(-Co-Mo-REE)

crusts

38%

Cobalt Figure 8

Stratiform

sediment-hosted

Cu-Co

41%

Magmatic

Ni-Cu(-Co-PGE)

15%

Ni-Co

laterite

36%

Other terrestrial

8%

Figure F8. Pie charts showing proportions of cobalt contained in mineral deposits worldwide (cumulative past production plus 

reserves plus other resources), by deposit type, for A, terrestrial deposits, and B, terrestrial and sea-floor deposits. Mine production in 

2011 from terrestrial deposit types is shown in figure F5. Globally, the amount of cobalt contained in reserves plus other resources in 

identified mineral deposits is significantly more than the amount of cobalt contained in past production. Co, cobalt; Cu, copper; Fe, iron; 

Mn, manganese; Mo, molybdenum; Ni, nickel; PGE, platinum-group element; REE, rare-earth element

The United States has 0.64 million metric tons of cobalt 

contained in several deposit types (table F1). The vast majority 

occurs within magmatic Ni-Cu(-Co-PGE) sulfide deposits of 

the Duluth Complex in northeastern Minnesota (which include 

the Nokomis deposit [291,000 metric tons], the Northmet 

deposit [92,400 metric tons], and the Mesaba deposit 

[89,000 metric tons]), and in metasedimentary-rock-hosted 

Co-Cu-Au deposits of the Blackbird district in east-central 

Idaho (123,000 metric tons). Much smaller amounts of cobalt 

are present within the Sheep Creek (Black Butte) stratiform 

sediment-hosted Cu-Co-Ag deposit in western Montana 

(17,700 metric tons) and the Eagle magmatic Ni-Cu(-Co-PGE) 

deposit in northern Michigan (4,000 metric tons). Resources 

also may be present in the Ni-Co laterites of southern Oregon 

and northern California (Foose, 1991, 1992). Compared with 

the estimated global terrestrial resources (plus past production, 

where available) of 25.5 million metric tons of cobalt, the 

United States has only 2.5 percent of the total amount of 

cobalt contained in mined ore (past production) plus reserves 

plus other resources.

On a global scale, the largest resources of cobalt occur 

in modern sea-floor ferromanganese nodules and crusts. 

Identified resources for eight different areas, each containing 

0.2 to 50 million metric tons of cobalt, have been reported. 

Crusts tend to have slightly higher cobalt grades but a similar 

range in deposit sizes. Significantly, the amount of cobalt in 

the Pacific prime crust zone in the Pacific Ocean (fig. F6) is 

three to four times the cobalt reserves contained in all known 

terrestrial deposits (Hein and others, 2013).

Production of cobalt during the next several decades will 

likely be dominated by the mining of Ni-Co laterites and stratiform sediment-hosted Cu-Co deposits, with lesser contributions 

from magmatic Ni-Cu(-Co-PGE) sulfide deposits. The giant 

cobalt resources present in modern sea-floor ferromanganese 

nodules and crusts (table F1; fig. F6) will probably be mined in 

the future, but given the technological challenges, environmental 

issues, and uncertain economics involved in such deep-sea 

mining ventures, it is unclear when these sea-floor mineral 

resources might be exploited profitably for cobalt production.

Mining Methods

The mining of cobalt deposits is done by conventional 

underground and open pit methods. Underground mining is 

used for most magmatic Ni-Cu(-Co-PGE) sulfide, iron oxideCu-Au(-Ag-U-REE-Co-Ni), volcanogenic Cu-(Co-Zn-Ag-Au) 

massive sulfide, and metasedimentary-rock-hosted Co-Cu-Au 

deposits. In contrast, open pit mining is the predominant 

method used for stratiform sediment-hosted Cu-Co and Ni-Co 

laterite deposits. Ore from the black-shale-hosted Talvivaara 

Ni-Cu-Zn-Co deposit in Finland—the only one of its type 

presently being mined—is extracted from large open pits.
F14 Critical Mineral Resources of the United States—Cobalt

Typical Mine Life

The mine life for cobalt-bearing deposits depends on 

a variety of parameters, but it is ultimately limited by the 

size of the resource and the development of reserves, which 

is the economically extractable portion of the demonstrated 

resource. Mines that produce cobalt generally are worked for 

decades. For example, several stratiform sediment-hosted 

Cu-Co deposits in the Central African Copperbelt have been 

mined for more than 50 years, such as Chibuluma, which 

began production in 1955 and is still being mined (Metorex 

Pty Ltd., 2013). In some other countries, mining started earlier 

and continues to the present day; for example, the magmatic 

Ni-Cu(-Co-PGE) sulfide deposits at the Creighton Mine in the 

Sudbury district of Ontario were first worked in 1901 (Python 

Mining Consultants, 2013); mining of the cobalt-rich veins in 

the Bou Azzer district of Morocco began in 1928 (Ahmed and 

others, 2009), and mining of the magmatic Ni-Cu(-Co-PGE) 

sulfide deposits of the Norilsk-Talnakh district in Russia 

began in 1936 (Naldrett, 2004). Possibly the longest period 

of production has come from the Thio Ni-Co laterite mine 

in New Caledonia, which was first mined in 1880 and is still 

being mined (Eramet Group, 2013).

Ore-Processing Methods

Cobalt-bearing ores are processed by a wide variety of 

extractive metallurgical techniques, depending on the type 

of ore, the availability of energy, environmental concerns, 

market demand for primary products and byproducts, and 

overall project economics. Ni-Co laterite ores generally 

require minimal beneficiation before being refined. Ores 

from magmatic Ni-Cu(-Co-PGE) sulfide and stratiform 

sediment-hosted Cu-Co deposits are processed by standard 

rock crushing and grinding techniques, typically followed 

by froth flotation to produce mineral concentrates, which are 

then refined to recover the metals (De Cuyper, 1988).

The ores or concentrates can be processed in one of the 

following ways: (a) leaching, (b) roasting and then leaching, 

or (c) smelting and then leaching. The resulting solutions 

are purified and refined to separate out the individual metals. 

The purification and refining are generally carried out by 

one or more hydrometallurgical and (or) electrometallurgical 

methods, including chemical precipitation, electrowinning, 

hydrogen reduction, ion exchange, and solvent extraction. 

Leaching of cobalt-bearing ores can be done at various 

temperatures and pressures using acids or other solutions, 

such as ammonia, ammonia–ammonium carbonate, chloride, 

or chlorine (Kerfoot and Weir, 1988). For more detailed 

information on the processing of cobalt ores, see, for example, 

Crundwell and others (2011).

In a few cases, leaching is accelerated by the introduction 

of bacteria. For example, ore from the black-shale-hosted 

Talvivaara deposit in Finland is processed by bio-heapleaching 

(Saari and Riekkola-Vanhanen, 2012). Bacterial leaching was 

also used at Kasese, Uganda, to recover cobalt from stockpiled 

pyrite concentrates that were generated during earlier 

copper mining.

Undiscovered Resources

Because stratiform sediment-hosted Cu-Co, Ni-Co 

laterite, and magmatic Ni-Cu(-Co-PGE) sulfide deposits are 

the sources of the greatest amount of cobalt production and 

contain the largest reserves of cobalt, most undiscovered 

resources of cobalt are likely to occur in these same deposit 

types. Descriptive and genetic models for such deposits are 

relatively well developed; hence, the exploration for new 

orebodies probably will target geologic terranes that host 

known deposits. In the case of stratiform sediment-hosted 

Cu-Co deposits, prospective geologic settings have been 

thoroughly documented (for example, Hitzman and others, 

2005; Taylor and others, 2013), yet potentially favorable 

terranes exist in many remote parts of the world where modern 

deposit models have not been applied. A similar situation 

exists for Ni-Co laterite deposits in that, despite a robust 

understanding of key ore-forming processes (for example, 

Freyssinet and others, 2005; Butt and Cluzel, 2013), areas 

that may contain deeply weathered ultramafic rocks have not 

been well explored. For magmatic Ni-Cu(-Co-PGE) sulfide 

deposits, the most important undiscovered resources are likely 

to be present in mafic and ultramafic igneous rocks, including 

large intrusive complexes (Naldrett, 2004; Eckstrand and 

Hulbert, 2007). The global distribution of such complexes 

is well known for those exposed at the surface, but in some 

remote parts of the world, buried mafic intrusive complexes 

still await discovery.

Unconventional Resources

Metamorphosed Black Shales

The black-shale-hosted Talvivaara Ni-Cu-Zn-Co deposit 

in Finland contains an average of 200 ppm cobalt (LoukolaRuskeeniemi and Lahtinen, 2013); however, the economic 

part of this deposit is located within remobilized structures in 

which lower metal contents of the surrounding Proterozoic 

black shale were increased during deformation and metamorphism. Epigenetic cobalt enrichment has also been described 

for the Permian Kupferschiefer Cu-Ag-Au-PGE deposits in 

Germany and Poland, where elevated cobalt contents occur 

preferentially within small, paragenetically late veins (Schmidt 

and Friedrich, 1988; Sun and Püttmann, 1997). Other black 

shales may have potential for containing low-grade cobalt 

deposits, such as the Upper Devonian-Lower Mississippian 

Chattanooga Shale in the Southeastern United States that 

typically has 70 ppm cobalt but locally contains as much as 

300 ppm (Leventhal and others, 1983).
Exploration for New Deposits F15

Modern Sea-Floor Volcanogenic Massive 

Sulfide Deposits

Some VMS deposits on the modern sea floor contain 

elevated amounts of cobalt. Documented examples of high 

contents of cobalt in deposits on the Mid-Atlantic Ridge are all in 

settings underlain by ultramafic rocks, including in the Rainbow 

vent field (Bogdanov and others, 2002), in the Logatchev 

field (Mozgova and others, 1999), and in the Nibelungen field 

(Melchert and others, 2008). Samples of massive sulfide rock 

from the Logatchev field have as much as 1,310 ppm cobalt 

(Murphy and Meyer, 1998). None of these modern deposits 

has been explored sufficiently for resource delineation, 

however, because of their occurrence at relatively deep water 

depths of 3,000 m or more, which are considered unfavorable 

for the profitable mining of such deposits (Herzig and others, 

2002). Despite this constraint, ultramafic-hosted modern 

VMS deposits may represent a resource for cobalt.

Ancient Ni-Co Laterites

Because Ni-Co laterite deposits of Cretaceous or 

younger ages are the predominant source of cobalt mined 

today (fig. F7; Berger and others, 2011), ancient laterites—if 

preserved and not eroded—could represent a nonconventional 

resource for nickel and cobalt. Gleeson and Herrington (2005) 

evaluated this potential, using the data of Bárdossy and Aleva 

(1990) on time periods of extensive global weathering, to 

suggest that fossil Ni-Co laterites preferentially formed during 

the Carboniferous, late Permian, Eocene-Oligocene, Miocene, 

and Pliocene, if favorable settings (including tropical to 

subtropical climates) existed at those times (see Thorne 

and others, 2012). Two examples of ancient laterites are the 

Çaldağ Ni-Co deposit in Turkey and the Ni-Co laterites of 

Greece, which formed during the Late Cretaceous to Eocene 

and Late Jurassic to Cretaceous, respectively (Çağatay and 

others, 1983; Valeton and others, 1987). Even older formational ages are possible, although paleoclimatic conditions 

during the Paleozoic and Precambrian are not well known.

Ancient Fe-Mn(-Cu-Co-Ni) Nodules

Ancient Fe-Mn(-Cu-Co-Ni) nodules are also a possible 

nonconventional resource of cobalt. Fossil nodules have 

been described from Miocene strata in Chile (Achurra and 

others, 2009) and Cretaceous strata in West Timor, Indonesia 

(Margolis and others, 1978). In both cases, however, the metal 

contents in the ancient nodules are much lower than those in 

modern nodules on the Pacific Ocean floor; such low metal 

concentrations may reflect shallow formation on continental 

slopes above the carbonate compensation depth (CCD). 

Metal-rich ancient nodules are most likely to be found within 

strata that formed in deepwater, abyssal plain settings below 

the CCD. These types of strata are generally carried on top 

of subduction zones and hence are not preserved, except in 

sedimentary sequences that are obducted onto continents.

Ancient Cobalt-Rich Manganese Deposits

Cobalt-rich manganese deposits occur in Pliocene 

alkaline basalts of the Calatrava region in central Spain 

(Crespo and Lunar, 1997). These unusual manganiferous 

deposits, which contain up to 1.7 weight percent cobalt, are 

interpreted to have formed in a hot-spring environment within 

a predominantly subaerial volcanic sequence. Although the 

known cobalt deposits are small, the potential may exist 

for larger deposits elsewhere in this volcanic field and also 

within other subaerial alkaline basaltic provinces that display 

evidence of hot-spring environments.

Ultramafic Igneous Rocks

The highest cobalt concentrations in common rock types 

are in ultramafic igneous rocks, including dunite, peridotite, 

and pyroxenite, and in altered varieties, such as serpentinite. 

Dunite tends to have the greatest contents because this rock 

type consists mainly of olivine, which is a ferromagnesian 

silicate mineral that preferentially concentrates cobalt. 

The average cobalt contents in dunite and serpentinite are 

109 and 115 ppm, respectively (Gülaҫar and Delaloye, 1976); 

hence, these rock types potentially could become largetonnage and low-grade cobalt resources if the extraction of 

cobalt from olivine and serpentine can be made commercially 

viable. The large amount of energy required to extract such 

cobalt suggests that this type of deposit has only a small 

chance of being developed and mined, however.

Exploration for New Deposits

Because of the diversity of deposit types that contain 

economic cobalt concentrations, exploration methods vary 

greatly. In the exploration for stratiform sediment-hosted 

Cu-Co deposits, prospecting may involve the integration 

of data from soil geochemical surveys, as well as structure, 

stratigraphy, geophysical, and drilling data. For lateritic 

Ni-Co deposits, key exploration criteria are the presence of 

deeply weathered ultramafic rocks, development of a thick 

lateritic profile, and optimum topography and rainfall required 

to effectively concentrate cobalt and nickel within the weathering zone. Geochemical sampling, including of vegetation, 

is typically effective in locating Ni-Co laterite deposits. Many 

magmatic Ni-Cu(-Co-PGE) sulfide deposits occur within large 

mafic intrusive complexes where refined deposit models and 

local geologic knowledge can be used to focus exploration 

to geophysics and drilling. Exploration for this deposit type 

is centered on the basal zones of the complexes using mainly 

electrical and magnetic geophysical methods.
F16 Critical Mineral Resources of the United States—Cobalt

Environmental Considerations

Cobalt is an essential nutrient for most life, but it 

can cause toxic effects when present in solids or waters at 

concentrations that are significantly enriched relative to 

normal background levels. Such elevated concentrations may 

result from cobalt mining, processing, manufacturing, and 

use. Consequently, an understanding of the behavior of cobalt 

in the environment can assist in developing technologies 

and working practices that do not lead to potentially toxic 

concentrations of cobalt during its production and use. The 

control, mitigation, and prevention of potential environmental 

impacts from mining and related processing facilities will vary 

depending on the deposit type, but they are also dependent on 

regulations in place within the host country, State, and local 

area. Therefore, no attempt is made herein to describe every 

type of mine and its potential for cobalt pollution. Instead, 

the focus is on some of the better documented environmental 

impacts associated with cobalt mining.

Sources and Fate in the Environment

In weathering environments at Earth’s surface, cobalt 

typically occurs in the +2 and +3 oxidation states, and 

may dissolve from host minerals and form complexes with 

hydroxide, fluoride, sulfate, phosphate, chloride, and (or) 

organic material. The behavior of cobalt in weathering environments follows that of iron and manganese. Manganese oxide 

minerals, in particular, have a strong capacity to sorb dissolved 

cobalt. Much of the literature about the behavior of dissolved 

cobalt focuses on optimizing cobalt sorption to various natural 

and synthetic solids as a means of removing dissolved cobalt 

from solution (for example, Chen and others, 2011).

The tendency for cobalt to be dissolved and transported 

largely depends upon the pH and temperature of weathering 

solutions. Because cobalt typically is concentrated in sulfide 

and arsenide minerals, it can be expected to be relatively 

mobile under some weathering conditions because of the 

instability of these minerals at Earth’s surface. Dissolution 

of sulfide minerals, in particular, releases metals and forms 

sulfuric acid, and the low pH values thus produced allow 

higher concentrations of metals to be dissolved—potentially 

causing the environmental problem known as acid mine 

drainage (AMD). Metals, including cobalt, which are 

dissolved in AMD can be naturally attenuated through 

precipitation, sorption to (oxyhydr)oxide minerals, or dilution 

by mixing with water at circumneutral pH; if not fixed in this 

manner, these metals are transported downstream and become 

more widely dispersed in the environment.

Natural concentrations of cobalt in rocks, soils, waters, 

and air are given in table F2. The cobalt content in soil varies 

widely depending upon the type of parent rock, but soils 

unaffected by pollution from industry generally contain a 

range of 0.5 to 30 ppm cobalt (Shacklette and Boerngen, 1984; 

Hutchinson and Symington, 1997; Tyler, 2004; Smith and 

others, 2005; Narendrula and others, 2012). Cobalt concentrations in seawater are 0.0003 to 0.004 micrograms per liter 

(μg/L, or parts per billion) (Bown and others, 2012; Shelley 

and others, 2012). In streams and rivers, cobalt concentrations 

vary from 0.006 to 0.43 μg/L (Gaillardet and others, 2003; 

Huser and others, 2011). Elevated background concentrations 

of as high as 30 μg/L cobalt have been reported near Cobalt, 

Idaho (Eppinger and others, 2003). Suspended particulates in 

world rivers contain an average of 22.5 micrograms per gram 

cobalt (Viers and others, 2009). Cobalt occurs naturally in the 

atmosphere within mineral dust particles. In relatively clean 

air over the South Pole, cobalt ranges from 0.1 to 1.2 nanograms per cubic meter (ng/m3

) (Kabata-Pendias and Pendias, 

2001, and references therein).

Concentrations of cobalt in the environment that are 

higher than background concentration can result from 

mining and ore processing (table F2). For example, soils 

near mines and processing facilities may contain as much as 

22 to 6,150 ppm cobalt (Hutchinson and Symington, 1997; 

Zohny, 2002; Giles and others, 2009; Narendrula and others, 

2012). Likewise, streams near the mining towns of Cobalt, 

Ontario, Canada, and Cobalt, Idaho, have 3 to 20 μg/L and 

10 to 1,100 μg/L cobalt, respectively (Kwong and others, 

2007; Gray and Eppinger, 2012). Extremely high cobalt 

concentrations of 70 to 61,000 μg/L (median is 540 μg/L) 

occur in the pore waters of mine waste impoundments at the 

Khovu-Aksy Mine in Russia (Bortnikova and others, 2012). 

Cobalt also can be emitted to the atmosphere during the 

metal refining process; air affected by industry may contain 

0.13 to 37 ng/m3 cobalt (Reimann and de Caritat, 1998, and 

references therein).

Mine Waste Characteristics

Mine waste is generally considered to be the material that 

originates and accumulates at a mine site but has no current 

economic value (Lottermoser, 2010); it includes both solid and 

liquid waste. The character of the waste generated from cobalt 

mining varies according to the geology of the deposit and the 

methods used to extract the ore. Mine wastes produced by 

the extraction of copper and nickel ores from which cobalt is 

produced typically consist of waste rock, tailings, and possibly 

pit lakes. Tailings are the residual silt- to fine sand-sized grains 

generated from ore grinding and processing, and generally are 

stored in dams or ponds.

The mining of cobalt typically produces large volumes 

of solid and liquid waste. Estimates for the amounts of solid 

waste generated from the mining of many deposits are difficult 

to obtain. The mining of laterite deposits requires the stripping 

of overburden before extraction of the exposed ore. As a 

result, reclamation of the mined area is often a significant 

issue, as soil needs to be reestablished and revegetated. 

Further, once the nickel and cobalt have been extracted, a 

volume of material nearly equal in volume to the original 

deposit must be secured in tailings so that it cannot be a source 
Environmental Considerations F17

of contamination. Laterite deposits typically are between 

20 and 200 million metric tons (table F1). In addition, laterite 

ores that are mined and then leached to recover nickel and 

cobalt can generate large amounts of metal-rich sludge that 

must be contained in impoundments, which are known as 

“dry stacks” (Power and others, 2011), or, if deemed environmentally safe, combined with additives to revegetate mined 

areas (Powers and Siemens, 1983).

Other open pit mines, such as those that are used to 

develop many strata-bound copper deposits, can be quite deep 

(for example, Taylor and others, 2013). A constraint on their 

economic viability is the stripping ratio, which determines 

how much waste rock must be moved. For example, the large 

Cu-Co mine at Tenke Fungurume (Congo [Kinshasa]) has an 

approximate stripping ratio of 3.3:1 (International Mining, 

2012), which means that 3.3 metric tons of rock must be 

moved to extract 1 metric ton of ore. The result is that large 

volumes of rock must be moved and managed, and waste rock 

must be placed in holding areas and secured. Underground 

mines, such as those typically used to mine certain sulfide ore 

deposits (magmatic Ni-Cu sulfide deposits and some stratiform 

sediment-hosted Cu-Co deposits), generally produce less 

waste than open pit mines; however, the waste commonly is 

highly reactive because of its high content of sulfide minerals. 

These minerals are capable of producing significant amounts 

of AMD, which, if not contained, must be remediated. One 

example of waste generated from a past-producing sulfidebearing cobalt mine is the Blackbird Mine near Cobalt, Idaho, 

which was mined intermittently from 1949 to 1960, yielding 

4.8 million metric tons of waste rock and 2 million metric 

tons of tailings, and resulting in AMD (U.S. Environmental 

Protection Agency, 2012).

Solid mine waste includes overburden (the soil and rock 

that overlie an orebody that must be removed to access the ore 

in an open pit mine) and gangue (the material that surrounds 

or is mixed with the valuable metallic minerals in the ore 

deposit). The mineralogy of solid mine waste in cobalt-rich 

deposits tends to be similar to the mineralogy of the deposit, 

except that the proportion of ore minerals is lower relative 

to gangue minerals. Because of their association with sulfide 

minerals, trace elements present in many cobalt-bearing 

deposits and wastes typically include As, Au, Bi, Co, Cr, 

Cu, Pb, Se, and Zn, and locally include Ag, Ba, Fe, Hg, Mg, 

Ni, Sb, Sc, Sn, Ti, V, and PGEs (Evans and others, 1995; 

Foose and others, 1995; Lindsey and others, 1995). Common 

elements associated with lateritic Ni-Co deposits are Al, Co, 

Cr, Fe, Mg, Mn, Ni, and Si (Watling and others, 2011; Lambiv 

Dzemua and others, 2013). Under oxidizing and acidic 

(pH<3) weathering conditions, which would be expected in 

mine waste having little to no acid-neutralizing capacity (such 

as those typically found in some stratiform sediment-hosted 

Cu-Co deposits, nearly all magmatic Ni-Cu(-Co-PGE) sulfide 

deposits, and in metasedimentary-rock-hosted Co-Cu-Au 

deposits), many of the aforementioned accessory elements 

are expected to be mobile (Smith and Huyck, 1999). The 

acid-neutralizing capacity of some stratiform sediment-hosted 

Cu-Co deposits is predicted to be greater owing to the presence of such gangue minerals as calcite and dolomite. For 

example, the paste pH of tailings samples from the mines 

in the Central African Copperbelt are as high as 6.9 at the 

Mindolo Mine and 8.5 at the Chambishi Mine (Sracek and 

others, 2010). Zinc in carbonate-hosted sulfide deposits of the 

Central African Copperbelt can make up a large percentage 

of the total dissolved metals that drain from the deposits and 

associated waste piles (Plumlee and others, 1999). Lateritic 

Ni-Co deposits generally lack acid-generating minerals, so 

the likely oxidizing and circumneutral (5<pH<8) weathering 

conditions in these and carbonate-bearing, stratiform 

sediment-hosted Cu-Co deposits allow many of the accessory 

elements listed above to be less mobile or immobile (Smith 

and Huyck, 1999).

Liquid waste includes ore-processing fluids that are 

stored onsite, as well as natural water that interacts with solid 

mine waste. Liquid mine waste can occur as surface water, 

groundwater, and soil pore water within and surrounding the 

mine site. As an example of liquid mine waste at a former 

cobalt mine, surface water draining from the Blackbird 

Mine near Cobalt, Idaho, is contaminated with As, Co, Cr, 

Cu, Fe, Mn, Ni, Pb, and Zn (Agency for Toxic Substances 

and Disease Registry, 1995). Cobalt concentrations in the 

waters in surrounding creeks may be nearly 70 times higher 

than background values (10 to 2,000 μg/L cobalt), as well 

as have high copper (10 to 1,800 μg/L copper), and arsenic 

(0.45 to 6.2 μg/L arsenic) concentrations (Mok and Wai, 

1989). A survey of streams within the broader Idaho cobalt 

belt revealed elevated concentrations (in μg/L) of cobalt 

(0.67 to 1,100), copper (1 to 2,000), arsenic (less than 

0.2 to 44), and iron (28 to 3,800) (Eppinger and others, 2003). 

Metal contamination of groundwater, surface water, soil, and 

stream sediments near the Blackbird Mine, as a result of runoff 

from waste rock, tailings, and mill debris, led to the proposed 

addition of the Blackbird Mine to the U.S. Superfund National 

Priorities List in 1993 (U.S. Environmental Protection Agency, 

2012). Acid mine drainage like that at the Blackbird Mine 

may be typical for historic mining districts, but, with modern 

controls, similar damage can generally be avoided or the 

harmful effects can be mitigated.

Human Health Concerns

The Agency for Toxic Substances and Disease Registry 

(2004) provides a useful summary of the health effects of 

cobalt on humans. The general public is most likely to be 

exposed to cobalt through consumption of food and drinking 

water. Human dietary intake of cobalt typically ranges from 

10 to 30 micrograms per day (Schrauzer, 2004). Occupational 

exposure generally results from inhalation of cobalt-bearing 

dust during cobalt processing and industrial activities that 

use cobalt, although the current use of appropriate protective 

equipment is thought to effectively minimize such exposure to 

levels that do not cause harm. At low levels, cobalt is essential 
F18 Critical Mineral Resources of the United States—Cobalt

to human health, particularly because it is the central atom in 

the critical nutrient vitamin B12. Overexposure to high levels 

of cobalt, however, may cause lung and heart dysfunction, as 

well as dermatitis.

Isotopes of cobalt range from 50Co to 71Co. The 

naturally occurring isotope is 59Co. 60Co, which has a 

half-life of 5.27 years and is the most commercially 

important radioactive isotope of cobalt, is made by placing 

59Co in a nuclear reactor for 1.5 to 2 years (Reimann and 

de Caritat, 1998). 60Co is used in commercial and medical 

applications. The Agency for Toxic Substances and Disease 

Registry (2004) stated that exposure to radiation from 

radioactive cobalt can damage cells and potentially cause 

nausea, vomiting, diarrhea, bleeding, coma, cancer, and, 

in rare cases, death. Additionally, nonradioactive cobalt 

does not appear to cause cancer in humans or animals when 

ingested, but when inhaled, animals have shown cancer 

development. More recently, the National Toxicology 

Program (NTP), made up of three U.S. Government health 

organizations, listed nonradioactive cobalt sulfate and 

cobalt-tungsten carbide in the Report on Carcinogens as 

reasonably anticipated to be human carcinogens (National 

Toxicology Program, 2014a, b). Beyond these listings, 

the NTP’s recent report revealed that there is not enough 

unequivocal evidence connecting cancer development in 

humans with exposure to cobalt and cobalt compounds 

(National Toxicology Program, 2016).

Primary and secondary drinking water regulations 

for cobalt currently do not exist in the United States, but 

the Occupational Safety and Health Administration (2013) 

has set an exposure limit of 0.1 milligram per cubic meter 

for cobalt-bearing dust in workplace air over an 8-hour 

workday. Likewise, the U.S. Nuclear Regulatory Commission (2013a, b) limits radioactive cobalt in workplace air 

to 1x10– 6 microcuries per milliliter (μCi/mL) of 57Co and 

7x10-8 μCi/mL of 60Co.

Mining of cobalt-rich ore deposits can potentially 

mobilize elements that are known human toxins. Perhaps 

the best-known examples of the effects of elements that 

are known human toxins are the neurological impacts 

of lead on children (Holecy and Mousavi, 2012) and the 

carcinogenic effects of arsenic in drinking water (Gupta 

and others, 2012). Other commonly associated elements, 

such as nickel, copper, and zinc, also have the potential 

to affect human health when present above threshold 

concentrations in air, drinking water, and soil resources. 

The current U.S. National Ambient Air Quality Standard for 

lead is 0.15 microgram per cubic meter (U.S. Environmental 

Protection Agency, 2013), and the current U.S. primary and 

secondary drinking water standards for arsenic, lead, and 

copper are 0.01, 0.0, and 1.3 milligrams per liter (mg/L), 

respectively (U.S. Environmental Protection Agency, 2016). 

Canadian agricultural soil quality guidelines for nickel, 

copper, lead, and zinc are 50, 63, 70, and 200 milligrams per 

kilogram (mg/kg, which is equivalent to ppm), respectively 

(Canadian Council of Ministers of the Environment, 2013).

Ecological Health Concerns

Many studies focus on the ecological impacts of cobalt 

bioaccessibility in the environment. Various organisms, 

including fish, amphibians, crustaceans, insects, and plants, 

are used in chronic (lower doses over longer time periods) 

and acute (higher doses over shorter time periods) tests of a 

substance toxicity. The aquatic toxicity of cobalt depends on 

multiple factors, including its chemical form (for example, free 

cobalt ion [Co2+], cobalt(II) chloride [CoCl2], and cobalt (II) 

sulfate [CoSO4(H2O)x]) and ambient physicochemical factors, 

such as pH, temperature, and overall water chemistry. One 

of several useful measurements used in toxicity tests is the 

lethal concentration that leads to 50 percent mortality (LC50 ) 

after exposure to a substance for a certain amount of time. 

Another commonly used measure of toxicity is the effective 

concentration that results in 50 percent of the tested population 

exhibiting decreased functionality (EC50), such as inhibited 

growth.

The freshwater amphipod Hyalella azteca is known to 

be sensitive to dissolved cobalt, exhibiting a 1-week LC50 of 

16 μg/L in soft water and 61 μg/L in hard water (Borgmann 

and others, 2005). The toxic effects of cobalt on this and other 

species decrease with increasing water hardness because less 

toxic ions like Ca2+ and Mg2+ are able to outcompete Co2+

for biological sites. Acute toxicity tests with dissolved cobalt 

chloride and the freshwater green alga Chlorella vulgaris

resulted in an EC50 of 530 μg/L cobalt, based on inhibited 

growth over 96 hours (Rachlin and Grosso, 1993). Fish 

appear to be more resistant to dissolved cobalt. For example, 

acute toxicity tests with dissolved cobalt (ranging from 

0 to 2,000 mg/L) and rainbow trout (Oncorhynchus mykiss) 

showed no lethality and an LC50 value of 1.4 mg/L cobalt 

after 2 and 3 days of exposure, respectively (Marr and others, 

1998). Interestingly, when this latter species was exposed to 

solutions containing dissolved copper only, dissolved cobalt 

only, or a mixture of the two metals, toxicity for time periods 

greater than 2 days decreased in the following order: 250 μg/L 

Co+Cu, 50 μg/L Co+Cu, Cu only, and Co only (Marr and 

others, 1998). These results indicate that exposure to the 

combination of dissolved cobalt and copper, which is closer to 

actual environmental conditions, is more toxic than exposure 

to either metal alone. Some regions of the United States have 

adopted secondary acute and chronic screening benchmarks of 

1,500 μg/L cobalt and 23 μg/L cobalt, respectively, for aquatic 

freshwater life (Suter and Tsao, 1996).

Although cobalt may be toxic to some species, it is an 

essential nutrient for many life forms, primarily because it 

is an integral component in vitamin B12 (Schrauzer, 2004). 

Microorganisms are responsible for the biosynthesis of natural 

vitamin B12. Some microorganisms secrete enzymes known 

as siderophores, which are known to have a high capacity for 

dissolving iron from relatively insoluble minerals, thereby 

increasing the bioavailability of iron in iron-limited environments. Recent work has shown that a commonly studied 

siderophore has a binding capacity for cobalt that is five 
Problems and Future Research F19

orders of magnitude greater than that for iron (Duckworth 

and others, 2009). Microorganisms may use siderophores to 

increase cobalt solubility and bioavailability in cobalt-limited 

environments as a means of contending with the generally 

low cobalt concentrations in the environment (Bi and others, 

2010). Microorganisms are responsible for synthesizing 

vitamin B12 in the gut of ruminants, including those that 

humans depend on for food, such as cows and sheep. Cobaltdeficient sheep and cattle may experience weight loss and 

frequent miscarriages, but such symptoms are often curtailed 

by adding cobalt salts to feed and cobalt additives to soil 

(Schrauzer, 2004).

Fewer studies have focused on the toxic effects of 

cobalt on higher level plants. The cobalt content of food 

plants varies widely (from 1.1 to 380 micrograms per 

kilogram) and depends upon the soil conditions and the 

species (Kabata-Pendias and Pendias, 2001, and references 

therein). This dependence is illustrated by a study that tested 

the phytotoxicity of cobalt to barley (Hordeum vulgare L.), 

oilseed rape (Brassica napus L.), and tomato (Lycopersicon 

esculentum L.) in 10 different soils collected from around the 

world (Li and others, 2009). The EC50, based on decreased 

shoot growth and biomass, was highly variable. The most 

sensitive EC50 was 7 mg/kg of cobalt for oilseed rape growing 

in soil from Athens (Georgia), and the least sensitive EC50

was 1,708 mg/kg of cobalt for barley growing in soil from 

Brécy (France). The bladder campion (Silene vulgaris 

[Moench] Garcke) plant is native to Europe and is widespread 

in North America. This species has the capacity for high 

metal tolerance, as expressed by seedling specimens collected 

in Canada from mine tailings near Cobalt, Ontario, having 

a greater ability to withstand elevated arsenic, cobalt, and 

nickel concentrations in the growth medium than counterparts 

collected from an uncontaminated site near Baymouth, 

Ontario (Paliouris and Hutchinson, 1991). For healthy 

terrestrial plant growth, some regions of the United States 

have adopted a soil screening benchmark of 20 mg/kg cobalt 

(Efroymson and others, 1997). The Canadian agricultural soil 

quality guideline for cobalt is 40 mg/kg (Canadian Council of 

Ministers of the Environment, 2013).

Mine Closure

Most recent and new mining operations include closure 

plans that address issues related to the mine’s footprint. A 

mine’s footprint includes the waste left on site and locally 

affected soil and water, as well as ecological impacts, such as 

habitat destruction and loss of biodiversity. Long-term plans 

for closed mining operations where cobalt is an economic 

byproduct also depend upon country, state, and local 

regulations. Following mining in many cobalt-rich deposits, 

a typical long-term mine closure issue is the potential for 

AMD derived from the site. Acidic drainage may seep from 

waste piles or tailings ponds. Common methods for treating 

AMD include active water treatment facilities, passive 

limestone-lined channels, or constructed wetlands (Plumlee 

and Logsdon, 1999). The end result of both active and passive 

approaches is the eventual precipitation of dissolved metals. 

Precipitated metals in passive wetland systems tend to be 

more stable under the prevailing anoxic conditions, whereas 

the metal-rich precipitates that result from active treatment 

facilities form a sludge that can be similar to Ni-Co laterite 

processing wastes, and both materials can cause environmental 

problems if not disposed of responsibly.

At large mines, mine waste is typically consolidated 

into pits and submerged under water, forming a tailings pond 

or impoundment. Acid-generating minerals are less reactive 

under water, but any seepage usually needs to be treated. A 

tailings impoundment in the Chambishi catchment in Zambia 

covers 1.6 hectares and is submerged during the wet season, 

but becomes completely dry by the middle of the dry season 

(von der Heyden and New, 2004). The acid-neutralizing 

capacity of the Chambishi tailings is sufficient to attenuate 

groundwater cobalt, nickel, and zinc concentrations to within 

local drinking water guideline levels, but evaporation in the 

tailings impoundment causes precipitation of metal-bearing 

evaporite minerals that can be mobilized as wind-blown dust 

during the dry season (von der Heyden and New, 2004). Likewise, dried tailings pond dust has settled in soils near the Tuva 

cobalt plant (Khovu-Aksy Mine site, Russia), contributing to 

soil arsenic concentrations as high as 540 mg/kg (Bortnikova 

and others, 2012).

Another common long-term mine-closure issue related 

to the mining of cobalt-rich deposits is the generation of 

large volumes of waste rock and tailings piles. These waste 

piles have the potential to become unstable and can be a 

source of metal-rich dust. If the use of mine waste as backfill 

into mine workings is not an option, waste pile stability and 

dust-generating issues can often be addressed through grading 

and covering of the piles with vegetation. Securing waste piles 

and prevention and treatment of AMD typically are taken into 

account in the long-term costs of active and proposed metal 

mining projects.

Problems and Future Research

Several types of studies could benefit the delineation 

of new cobalt resources for the future and the mining and 

processing of known cobalt deposits. First would be better 

geologic models for cobalt-rich ores, which could be used in 

the exploration for new deposits. Second would be improved 

methods for increased recovery of cobalt from Ni-Co laterites 

based on the typically lower recovery rates for cobalt relative 

to nickel (Xu and others, 2005); in situ leaching of laterite 

deposits could also be evaluated. Third would be the development of processes for the efficient and economic recovery of 

cobalt from silicate minerals, such as olivine in dunite and 

other types of ultramafic rocks. Fourth, and last, would be 

technological advances to help make the mining and extraction 
F20 Critical Mineral Resources of the United States—Cobalt

of cobalt from deep-sea ferromanganese nodules and crusts 

economically profitable.

Considerable challenges exist for the profitable 

extraction of cobalt from sea-floor resources in ferromanganese nodules and crusts (figs. F6 and F8). These cobalt 

resources lie at water depths of as great as 6,000 m, so 

technological, economic, and legal barriers have prevented 

their exploitation to date. The International Seabed Authority 

has developed guidelines for the environmental impact 

statement (EIS) that must be submitted to governing bodies 

by companies applying for licenses to explore and mine 

sea-floor mineral deposits. According to the EIS guidelines, 

developers must address all environmental and social issues 

that could result from the mining activity, as well as the 

relevant mitigation measures (International Seabed Authority, 

2012). Discussion of all potential issues is beyond the scope 

of this work. Interested readers can find more information 

in the various technical reports published online by the 

International Seabed Authority (2015).

Processing of sea-floor ferromanganese nodules and 

crusts would likely take place on land, so some of the same 

mine-waste issues for terrestrial cobalt resources will also 

apply to sea-floor resources. Additionally, perhaps one of the 

most controversial issues involving sea-floor mining is the 

potential impact on benthic, middepth, and pelagic ecosystems. Benthic ecosystems on abyssal plains are particularly 

threatened by mining in the Clarion-Clipperton zone (CCZ). 

The understanding of abyssal, benthic ecosystems has grown 

rapidly in the past several decades. An international effort 

to evaluate the biodiversity in the CCZ revealed that “high, 

unanticipated, and still poorly sampled levels of species 

diversity for all three sediment-dwelling faunal components 

(foraminifera, nematodes, and polychaetes) [exist] at [the] 

individual study sites” (International Seabed Authority, 

2008, p. 2). These findings led to the recommendation that 

“marine protected areas” be maintained throughout the 

CCZ in order to preserve abyssal biodiversity (International 

Seabed Authority, 2008). Another major environmental 

concern relates to wastewater being discharged from the 

ships carrying out the sea-floor mining. This waste could 

contain crushed nodules and trace metals, and has the 

potential to disrupt photosynthesis by pelagic organisms 

and, being colder and denser than surface water, change the 

natural circulation patterns in the water column (Markussen, 

1994). As a result, current conceptual designs are engineered 

to separate nodules from waste material (for example, clay, 

sediments, and muddy water) and to return waste to the sea 

floor at water depths of nearly 5,000 m, so as not to disturb 

the middepth and pelagic ecosystems (for example, Agarwal 

and others, 2012). As sea-floor mineral exploration continues 

and if mining begins in the future, the scientific community 

likely will be documenting its impacts on ocean life.

Acknowledgments

We thank Michael P. Foose (U.S. Geological Survey) and 

Andrew G. Gunn (British Geological Survey) for thorough and 

constructive reviews of the manuscript, and John H. DeYoung, Jr. 

(U.S. Geological Survey) for editorial suggestions. Comments

by David Weight and Paul Marsh (The Cobalt Development

Institute), and William Stubblefield (Oregon State University)

are also appreciated. Keith Labay (U.S. Geological Survey)

helped create the deposit location maps, and Suzanne Nicholson

(U.S. Geological Survey) provided additional editorial 

comments after the technical review was complete.

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Williams, P.J., Barton, M.D., Johnson, D.A., Fontboté, Lluís, 

de Haller, Antoine, Mark, Geordie, Oliver, N.H.S., and 

Marschik, Robert, 2005, Iron oxide copper-gold deposits—

Geology, space-time distribution, and possible modes of 

origin, in Hedenquist, J.W., Thompson, J.F.H., Goldfarb, R.J.,

and Richards, J.P., eds., Economic Geology—One hundredth

anniversary volume, 1905–2005: Littleton, Colo., Society 

of Economic Geologists, p. 371–405. [Appendixes are on a 

CD–ROM inside the back cover.]

Wilson, A.B., Causey, J.D., Denning, P.D., Hayes, T.S., 

Horton, J.D., Kirschbaum, M.J., Parks, H.L., Taylor, C.D., 

and Zientek, M.L., 2013, Appendix A—Relational database 

of sediment-hosted copper deposits, resources, and production in the Central African Copperbelt, Democratic Republic 

of the Congo and Zambia, in Taylor, C.D., Causey, J.D., 

Denning, P.D., Hammarstrom, J.M., Hayes, T.S., Horton, J.D.,

Kirschbaum, M.J., Parks, H.L., Wilson, A.B., Wintzer, N.E., 

and Zientek, M.L., Descriptive models, grade-tonnage 

relations, and databases for the assessment of sedimenthosted copper deposits—With emphasis on deposits in 

the Central African Copperbelt, Democratic Republic of 

the Congo and Zambia: U.S. Geological Survey Scientific 

Investigations Report 2010–5090–J, p. 103–144. [Also 

available at http://pubs.usgs.gov/sir/2010/5090/j/.]

Xstrata plc, 2011, Mineral resources & ore reserves: Xstrata 

plc, accessed April 22, 2013, at http://www.xstrata.com/

publications/reservesandresources/.

Xu, Y.-B., Xie, Y.-T., Yan, Lan, and Yang, R.-D., 2005, A new 

method for recovering valuable metals from low-grade 

nickeliferous oxide ores: Hydrometallurgy, v. 80, no. 4, 

p. 280–285. [Also available at http://dx.doi.org/10.1016/

j.hydromet.2005.08.007.]

Xu, Yong, and Zhu, Xinyou, 2000, Cobalt deposits in the 

central China orogenic belt: Acta Geologica Sinica, v. 74, 

no. 3, p. 540–543.

Zientek, M.L., Hayes, T.S., and Hammarstrom, J.M., 2013, 

Overview of a new descriptive model for sediment-hosted 

stratabound copper deposits, chap. 1 of Taylor, C.D.,

Causey, J.D., Denning, P.D., Hammarstrom, J.M., Hayes, T.S.,

Horton, J.D., Kirschbaum, M.J., Parks, H.L., Wilson, A.B., 

Wintzer, N.E., and Zientek, M.L., Descriptive models, 

grade-tonnage relations, and databases for the assessment 

of sediment-hosted copper deposits—With emphasis on 

deposits in the Central African Copperbelt, Democratic 

Republic of the Congo and Zambia: U.S. Geological Survey 

Scientific Investigations Report 2010–5090–J, p. 2–16. 

[Also available at http://pubs.usgs.gov/sir/2010/5090/j/.]

Zijin Mining Group Co. Ltd., 2010, Information of Deziwa and

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Zijin Mining Group Co. Ltd., accessed April 22, 2013, at 

www.zjky.cn/portals/1/ltn20100507861.pdf.

Zohny, E.A.M., 2002, Cobalt in alluvial Egyptian soils 

affected by industrial activities: Journal of Environmental 

Sciences (China), v. 14, no. 1, p. 34–38.
Table F1
F34 Critical Mineral Resources of the United States—Cobalt

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.—Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage

(Mt)

Grade

(% Co)

Co

content

(Mt)

Principal data source

and notes (decimal degrees

WGS 84)

Stratiform sediment-hosted Cu-Co deposits

Browns Australia –13.07 131.08 84.0 0.120 0.101 Compass Resources NL (2007)

Emmie Bluff Australia –31.10 137.17 25.0 0.100 0.025 Hitzman and others (2005)

Mount Gunson Australia –31.45 137.25 11.7 0.060 0.007 Hitzman and others (2005)

Mount Isa Australia –20.57 139.47 182.0 0.050 0.091 Mount Isa Mines Ltd. (2012)

Pyrite Hill-Big Hill Australia –32.12 141.20 20.8 0.085 0.018 Broken Hill Prospecting Ltd. (2013)

Windabout Australia –31.30 137.16 18.7 0.050 0.009 Gunson Resources Ltd. (2013)

Tangdan (Dongchuan) China 26.12 103.17 146.8 0.010 0.015 Hitzman and others (2005)

Deziwa Congo (Kinshasa) –10.78 25.77 319.4 0.120 0.383 Zijin Mining Group Co. Ltd. (2010)

Dilala East Congo (Kinshasa) –10.70 25.47 19.1 0.900 0.172 Metorex Pty Ltd. (2011)

DIMA Congo (Kinshasa) –10.74 25.38 280.0 0.189 0.530 Wilson and others (2013)

Disele Congo (Kinshasa) –10.75 26.25 4.1 0.068 0.003 Wilson and others (2013)

Écaille C Congo (Kinshasa) –10.76 26.12 21.8 0.060 0.013 Zijin Mining Group Co. Ltd. (2010)

Etoile (Ruashi) Congo (Kinshasa) –11.63 27.58 52.6 0.300 0.158 Metorex Pty Ltd. (2011)

Kababankola Congo (Kinshasa) –10.73 26.42 1.9 1.120 0.021 Wilson and others (2013)

Kabolela Congo (Kinshasa) –10.69 26.45 11.7 0.501 0.059 Wilson and others (2013)

Kakanda Congo (Kinshasa) –10.74 26.40 32.4 0.180 0.058 Wilson and others (2013)

Kalukundi-Kii Congo (Kinshasa) –10.62 25.91 12.8 0.457 0.058 Wilson and others (2013)

Kambove Congo (Kinshasa) –10.85 26.61 46.9 0.213 0.100 Wilson and others (2013)

Kamfundwa Congo (Kinshasa) –10.81 26.59 26.8 0.200 0.054 Wilson and others (2013)

Kamoto-KOVMusonoi-MupineCongo (Kinshasa) –10.72 25.40 416.0 0.387 1.608 Wilson and others (2013)

Kananga Congo (Kinshasa) –10.68 25.48 8.1 0.884 0.072 Wilson and others (2013)

Kazibizi Congo (Kinshasa) –10.83 26.57 1.0 0.600 0.006 Wilson and others (2013)

Kipapila Congo (Kinshasa) –12.02 27.90 2.6 0.540 0.014 Wilson and others (2013)

Kipoi Central Congo (Kinshasa) –11.26 27.10 4.6 0.143 0.007 Wilson and others (2013)

Kipoi North Congo (Kinshasa) –11.26 27.10 5.3 0.049 0.003 Wilson and others (2013)

Kisanfu Congo (Kinshasa) –10.77 25.98 108.0 1.080 1.166 Freeport-McMoRan Copper & 

Gold Inc. (2013)

Luishia Congo (Kinshasa) –11.17 27.01 72.3 0.152 0.110 Wilson and others (2013)

Luiswishi Congo (Kinshasa) –11.51 27.42 8.0 1.100 0.088 Wilson and others (2013)

Lupoto Congo (Kinshasa) –11.60 27.26 24.1 0.088 0.021 Wilson and others (2013)

M’Sesa Congo (Kinshasa) –10.85 26.58 8.0 0.213 0.017 Wilson and others (2013)

Mukondo (includes 

C19 and C21)

Congo (Kinshasa) –10.73 26.35 70.0 0.714 0.500 CAMEC Plc (2007)

Mutanda Congo (Kinshasa) –10.79 25.81 299.5 0.59 1.767 Glencore International plc (2011)

Mutoshi Congo (Kinshasa) –10.65 25.51 10.8 0.304 0.033 Wilson and others (2013)

Tenke Fungurume1 Congo (Kinshasa) –10.58 26.19 674.0 0.245 1.651 Nilsson and others (2011)

Tilwezembe Congo (Kinshasa) –10.80 25.69 23.4 0.637 0.149 Wilson and others (2013)
Table F1 F35

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.— Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage 

(Mt)

Grade 

(% Co)

Co 

content 

(Mt)

Principal data source 

and notes (decimal degrees 

WGS 84)

Stratiform sediment-hosted Cu-Co deposits—Continued

Boleo

Kilembe

Sheep Creek

(Black Butte)

Baluba-MuliashiLuanshya

Chambishi Southeast

ChibulumaChibuluma West

Mexico

Uganda

United States

Zambia

Zambia

Zambia

27.33

0.22

46.73

–13.50

–12.66

–12.92

–112.30

30.05

–110.90

28.38

28.05

28.08

424.5

16.2

11.6

339.0

178.0

19.9

0.054

0.170

0.100

0.033

0.069

0.186

0.231

0.028

0.012

0.111

0.123

0.037

Baja Mining Corp. (2011)

Vangold Resources Ltd. (2013)

Tintina Resources Inc. (2016)

Wilson and others (2013)

Wilson and others (2013)

Wilson and others (2013)

Chimiwungo

Chingola-Nchanga

Kalumbila

Malundwe

Mindola-Nkana N-S

Nama

Zambia

Zambia

Zambia

Zambia

Zambia

Zambia

–12.18

–12.51

–12.21

–12.16

–12.80

–12.26

25.85

27.87

25.35

25.75

28.18

27.62

761.0

576.0

340.0

162.0

656.0

185.8

0.009

0.026

0.030

0.014

0.065

0.059

0.069

0.149

0.102

0.023

0.429

0.109

Wilson and others (2013)

Wilson and others (2013)

Wilson and others (2013)

Wilson and others (2013)

Wilson and others (2013)

Wilson and others (2013)

Ni-Co laterite deposits

Bell Creek-NeckMinnamoolkaAustralia –18.14 144.79 18.4 0.050 0.009 Metallica Minerals Ltd. (2013)

Cawse

Claude Hills

Greenvale-KokomoLucknow

Australia

Australia

Australia

–30.38

–26.00

–18.97

121.15

129.30

144.94

217.0

33.3

59.5

0.040

0.070

0.070

0.087

0.023

0.042

Berger and others (2011)

Metals X Ltd. (2013)

Metallica Minerals Ltd. (2013)

Kalgoorlie Nickel2

Marlborough

Murrin Murrin

Ora Banda

Ravensthorpe

Summervale/Westlynn

Weld Range

Wingellina

Yerilla

Young

Jacaré

Niquelandia

Santa Fe

Serra do Tapa

Vale dos Sonhos

Australia

Australia

Australia

Australia

Australia

Australia

Australia

Australia

Australia

Australia

Brazil

Brazil

Brazil

Brazil

Brazil

–30.10

–23.01

–28.77

–30.24

–33.65

–31.55

–26.82

–26.06

–29.73

–35.77

–6.21

–14.35

–15.67

–6.85

–6.97

121.05

149.89

121.89

121.00

120.40

147.17

117.75

128.97

122.04

148.82

–52.86

–48.40

–51.12

–49.33

–49.26

727.0

210.0

464.0

30.0

386.0

16.0

330.0

347.7

135.4

167.0

495.0

56.3

61.0

60.3

13.0

0.044

0.060

0.076

0.080

0.027

0.064

0.060

0.080

0.052

0.070

0.130

0.060

0.060

0.050

0.070

0.320

0.126

0.351

0.024

0.104

0.010

0.198

0.278

0.070

0.117

0.644

0.034

0.037

0.030

0.009

Heron Resources Ltd. (2013)

Berger and others (2011)

Minara Resources Pty Ltd. (2011)

Berger and others (2011)

Berger and others (2011)

Jervois Mining Ltd. (2013)

Minara Resources Pty Ltd. (2005)

Metals X Ltd. (2013)

Heron Resources Ltd. (2013)

Jervois Mining Ltd. (2013)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)
F36 Critical Mineral Resources of the United States—Cobalt

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.— Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage 

(Mt)

Grade 

(% Co)

Co 

content 

(Mt)

Principal data source 

and notes (decimal degrees 

WGS 84)

Ni-Co laterite deposits—Continued

Vermelho

Vila Oito

Musongati

Nyabikere

Waga

Nkamouna (includes 

Mada)

Moa

Nicaro 

Pinares de Mayari

Punta Gorda

Sechol

Bhimatangar

Kansa

Saruabil

Gag Island

Halmahera/Weda Bay

La Sampala

Tiuba-Biankouma

Gornostai

Kempirsai

Cikatova

Glavica

Ržanovo

Ambatovy

Analamay

Tavai

Goro

Kouaoua

Nakety

Nepoui

Thio

Tiébaghi

Mambare

Ramu

Wowo Gap

Brazil

Brazil

Burundi

Burundi

Burundi

Cameroon

Cuba

Cuba

Cuba

Cuba

Guatemala

India

India

India

Indonesia

Indonesia

Indonesia

Ivory Coast

Kazakhstan

Kazakhstan

Kosovo (Serbia)

Kosovo (Serbia)

Macedonia

Madagascar

Madagascar

Malaysia

New Caledonia

New Caledonia

New Caledonia

New Caledonia

New Caledonia

New Caledonia

Papua New 

Guinea

Papua New 

Guinea

Papua New 

Guinea

–6.24

–7.79

–3.79

–3.28

–3.56

3.27

20.60

20.63

20.54

20.63

15.43

21.01

21.06

21.06

–0.45

–0.50

–2.70

7.74

50.61

50.34

42.22

42.54

41.17

–18.87

–18.85

5.58

–22.31

–21.42

–21.51

–21.22

–21.62

–20.45

–9.00

–5.53

–9.68

–49.75

–49.35

30.19

30.09

29.91

13.81

–74.96

–75.54

–75.72

–74.89

–89.66

85.74

85.86

85.82

129.88

127.94

121.93

–7.62

78.76

58.28

21.26

21.01

21.99

48.35

48.37

117.33

166.93

165.74

166.05

165.04

166.18

164.22

147.70

145.19

148.96

290.0

12.9

185.0

46.4

35.0

391.5

290.0

200.0

400.0

310.0

36.1

63.9

40.9

18.4

240.0

277.0

162.0

293.0

9.7

179.7

13.0

7.5

43.0

164.4

56.3

270.0

323.0

30.0

229.0

16.0

208.0

39.2

162.5

143.2

190.0

0.026

0.065

0.080

0.030

0.040

0.220

0.180

0.100

0.100

0.110

0.080

0.150

0.080

0.080

0.080

0.090

0.080

0.110

0.059

0.054

0.070

0.050

0.060

0.091

0.104

0.070

0.110

0.080

0.120

0.080

0.100

0.090

0.090

0.100

0.060

0.075

0.008

0.148

0.014

0.014

0.860

0.522

0.200

0.400

0.341

0.029

0.096

0.033

0.015

0.192

0.249

0.130

0.322

0.006

0.097

0.009

0.004

0.026

0.150

0.059

0.189

0.355

0.024

0.275

0.013

0.208

0.035

0.146

0.143

0.114

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Geovic Mining Corp. (2013)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Tiger International 

Resources Inc. (2006)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Direct Nickel Pty Ltd. (2013)

Berger and others (2011)

Berger and others (2011)
Table F1 F37

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.— Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage 

(Mt)

Grade 

(% Co)

Co 

content 

(Mt)

Principal data source 

and notes (decimal degrees 

WGS 84)

Ni-Co laterite deposits—Continued

Acoje/Zambales

Agata/Bolobolo/ 

Karihatag

Berong

Cagdianao

Ipilan

Mindoro

Nonoc

Rio Tuba (Coral Bay)

Taganaan

Taganito/Adlay

Guanajibo

Las Mesas

Akkermanov

Buruktal

Cheremshanskoe 

Lipovskoe

Rogozhinskoe

Serovskoe

Sinarskoe

Rudjinci

Isabela Island (Bugotu)

Dutwa

Çaldağ

Philippines

Philippines

Philippines

Philippines

Philippines

Philippines

Philippines

Philippines

Philippines

Philippines

Puerto Rico

Puerto Rico

Russia

Russia

Russia

Russia

Russia

Russia

Russia

Serbia

Solomon Islands

Tanzania

Turkey

15.72

9.28

9.39

9.92

8.97

12.94

9.84

8.59

9.78

9.53

18.13

18.19

51.21

50.85

56.12

57.43

56.17

59.71

56.07

43.59

–8.39

–3.67

38.62

120.04

125.52

118.14

125.67

117.98

121.07

125.62

117.41

125.72

125.83

–67.16

–67.10

58.23

60.83

60.30

61.11

60.26

60.56

60.40

20.84

159.71

33.42

27.77

72.3

35.0

275.0

8.0

77.0

315.0

144.7

57.7

20.0

123.0

44.4

22.7

40.0

79.0

45.0

46.0

2.5

52.8

18.0

14.0

24.0

107.0

37.9

0.070

0.060

0.080

0.140

0.100

0.060

0.110

0.123

0.092

0.116

0.080

0.120

0.010

0.100

0.024

0.057

0.060

0.043

0.055

0.050

0.063

0.028

0.050

0.051

0.021

0.220

0.011

0.077

0.189

0.159

0.071

0.018

0.143

0.036

0.027

0.004

0.079

0.011

0.026

0.002

0.023

0.010

0.007

0.015

0.030

0.019

Berger and others (2011)

Gifford and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Nickel Asia Corp. (2011)

Nickel Asia Corp. (2011)

Nickel Asia Corp. (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

Berger and others (2011)

African Eagle Resources Plc (2013)

Berger and others (2011)

Magmatic Ni-Cu(-Co-PGE) sulfide deposits

Kambalda

Mt. Keith

Savannah (Sally 

Malay)

Americano do Brasil 

(Salgado)

Fortaleza de Minas

Santa Rita

Dumont

Ferguson Lake

Makwa (Maskwa)

Australia

Australia

Australia

Brazil

Brazil

Brazil

Canada

Canada

Canada

–31.64

–27.71

–17.35

–16.10

–21.00

–14.19

48.65

62.92

50.47

121.65

120.53

128.02

–50.08

–48.50

–39.73

–78.44

–96.97

–95.43

67.0

478.0

8.1

7.9

10.3

159.3

2,134.2

44.2

19.1

0.207

0.014

0.073

0.040

0.200

0.015

0.011

0.080

0.010

0.139

0.067

0.006

0.003

0.021

0.024

0.226

0.035

0.002

Naldrett (2004)

Naldrett (2004)

Panoramic Resources Ltd. (2012)

Votorantim Group (2006)

Votorantim Group (2006)

Mirabela Nickel Ltd. (2012)

Staples and others (2012)

Starfield Resources Inc. (2013)

Mustang Minerals Corp. (2013)
F38 Critical Mineral Resources of the United States—Cobalt

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.— Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage 

(Mt)

Grade 

(% Co)

Co 

content 

(Mt)

Principal data source 

and notes (decimal degrees 

WGS 84)

Magmatic Ni-Cu(Co-PGE) sulfide deposits—Continued

Nunavik (includes 

Mesamax)

Raglan

Shakespeare

Sudbury

Thompson

Turnagain

Voisey’s Bay

Wellgreen

Jinchuan

Yangliuping

Hitura

Kevitsa

Kotalahti

Laukunkangas

Stormi

Bruvann (Råna)

Stormyra (including 

Dalen)

Ertelien

Flåt

NKT

Noril’sk-Talnakh area

Pechenga

Sopchuaivench

Nkomati

Aguablanca

Kabanga

Eagle

Mesaba (Babbitt/

Minnamax)

Northmet

Twin Metals 

(Nokomis)3

Munali

Canada

Canada

Canada

Canada

Canada

Canada

Canada

Canada

China

China

Finland

Finland

Finland

Finland

Finland

Norway

Norway

Norway

Norway

Russia

Russia

Russia

Russia

South Africa

Spain

Tanzania

United States

United States

United States

United States

Zambia

61.57

61.59

46.35

46.63

53.76

58.48

45.50

60.75

38.47

30.67

63.84

67.70

62.52

62.05

61.33

68.33

61.34

60.07

58.60

67.94

69.36

69.33

67.88

–25.67

37.85

–2.87

46.75

47.63

47.60

47.80

–16.00

–73.26

–74.29

–81.83

–81.38

–98.81

–128.83

–73.57

–139.50

102.17

102.00

25.05

29.97

27.68

28.77

24.90

16.94

9.66

10.05

7.87

32.82

88.31

29.74

32.82

30.50

–6.18

30.58

–87.90

–91.88

–91.97

–91.76

28.00

27.1

44.0

11.8

2,648

150.0

763.0

137.0

461.0

515.0

63.8

19.4

435.4

12.4

7.9

9.0

43.0

11.1

3.1

3.2

8.2

1,309.0

339.0

131.2

290.6

6.0

58.2

5.2

1,106.5

923.9

2,425.9

27.7

0.048

0.060

0.020

0.038

0.046

0.013

0.090

0.021

0.019

0.016

0.020

0.015

0.030

0.030

0.040

0.010

0.022

0.074

0.075

0.230

0.061

0.045

0.014

0.020

0.600

0.198

0.080

0.008

0.010

0.012

0.069

0.013

0.026

0.002

1.006

0.069

0.102

0.123

0.097

0.098

0.010

0.004

0.064

0.004

0.002

0.004

0.004

0.002

0.002

0.002

0.019

0.798

0.153

0.018

0.058

0.036

0.115

0.004

0.089

0.092

0.291

0.019

Armstrong, Puritch, and Yassa 

(2010)

Xstrata plc (2011)

Prophecy Platinum Corp. (2013a)

Naldrett (2004)

Naldrett (2004)

Hard Creek Nickel Corp. (2013)

Naldrett (2004)

Prophecy Platinum Corp. (2013b)

Naldrett (2004)

Pirajno (2013)

Eilu (2012)

First Quantum Minerals Ltd. (2011)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Naldrett (2004)

Naldrett (2004)

Eilu (2012)

African Rainbow Minerals Ltd. (2011)

Lundin Mining Corp. (2012)

Xstrata plc (2011)

Owen and Meyer (2013)

Teck Resources Ltd. (2010)

Polymet Mining Corp. (2013)

Duluth Metals Ltd. (2012)

Albidon Ltd. (2007)

Black-shale-hosted Ni-Cu-Zn-Co deposits

Talvivaara Finland 63.98 28.17 2,053 0.020 0.411 Talvivaara Mining Company Plc 

(2013)
Table F1 F39

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.— Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage 

(Mt)

Grade 

(% Co)

Co 

content 

(Mt)

Principal data source 

and notes (decimal degrees 

WGS 84)

Fe-Cu-Co skarn and replacement deposits

Goroblagodat

Magnitogorsk

Cornwall

Russia

Russia

United States

61.42 60.50 160.0 0.022

53.42 59.10 500.0 0.018

40.27 –76.40 100.0 0.025

0.035

0.090

0.025

Herrington and others (2005)

Herrington and others (2005)

Rose and others (1985); grade 

calculated

Iron oxide-Cu-Au(-Ag-U-REE-Co-Ni) deposits

Basil Australia –22.07 135.55 26.5 0.050 0.013 Mithril Resources Ltd. (2013)

Ernest Henry Australia –20.43 140.70 166.0 0.050 0.083 Williams and Pollard (2001)

Greenmount Australia –21.03 140.53 23.8 0.049 0.012 Williams and Pollard (2001)

Mount Oxide Australia –19.48 139.47 15.5 0.050 0.008 Chalice Gold Mines Ltd. (2012)

Olympic Dam Australia –30.43 136.88 605.0 0.020 0.121 Williams and Pollard (2001)

Rover 1 Australia –20.00 133.66 6.8 0.060 0.004 Metals X Ltd. (2013)

Lala (Lalachang) China 26.10 101.90 200.0 0.022 0.044 Chen and Zhou (2012)

Guelb Moghrein Mauritania 19.75 –14.38 23.7 0.014 0.003 Kolb and others (2006)

Metasedimentary-rock-hosted Co-Cu-Au deposits

NICO Canada 63.55 –116.75 31.0 0.120 0.037 Slack (2013)

Werner Lake Canada 50.47 –94.97 1.3 0.270 0.004 Slack (2013)

Haarakumpu Finland 66.42 28.57 4.7 0.170 0.008 Slack (2013)

Juomasuo Finland 66.29 29.20 5.0 0.128 0.006 Slack (2013)

Kouvervaara Finland 66.13 28.82 1.6 0.100 0.002 Slack (2013)

Vähäjoki Finland 66.11 25.28 10.5 0.030 0.003 Slack (2013)

Skuterud Norway 59.91 9.89 1.0 0.260 0.003 Slack (2013)

Blackbird (district) United States 45.12 –114.35 16.8 0.735 0.123 Slack (2013)

Polymetallic (Ag-Ni-Co-As-Bi) and other cobalt-rich veins

Cobalt-Gowganda 

(districts)

Canada 47.40 –79.69 2.0 1.000 0.020 Petruk and others (1971)

Bou Azzer (district) Morocco 30.53 –6.91 4

5.7 41.500 0.085 Bouabdellah and others (2016)

Karakul Russia 49.91 89.55 14.9 0.260 0.039 Puget Ventures Inc. (2013)

Volcanogenic Cu(-Zn-Co-Ag-Au) massive sulfide deposits

Chu Chu Canada 51.38 –120.06 7.0 0.100 0.007 Mosier and others (2009)

Soucy No. 1 Canada 58.32 –69.87 5.4 0.090 0.005 Mosier and others (2009)

Windy Craggy Canada 59.73 –137.73 297.0 0.069 0.205 Peter and Scott (1999)

Deerni China 34.40 100.13 54.0 0.080 0.043 Galley and others (2007)

Dur’ngoi China 34.33 100.24 31.9 0.089 0.028 Xu and Zhu (2000)

Luikonlahti Finland 62.92 28.67 3.4 0.130 0.004 Mosier and others (2009)

Outokumpu Finland 62.73 29.01 29.2 0.250 0.073 Mosier and others (2009)
F40 Critical Mineral Resources of the United States—Cobalt

Table F1. Location, grade, tonnage, and other data for selected cobalt deposits of the world.— Continued

[Tonnage and grade include reserves, other resources, and past production (where available). WGS 84, World Geodetic System of 1984; negative values 

for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; 

Congo (Kinshasa), Democratic Republic of the Congo; EEZ, Exclusive Economic Zone; Mt, million metric tons; %, percent. Element: Ag, silver; Au, gold; 

As, arsenic; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Mn, manganese; Mo, molybdenum; Ni, nickel; REE, rare-earth-elements; Zn, zinc]

Deposit Country

Latitude Longitude Tonnage 

(Mt)

Grade 

(% Co)

Co 

content 

(Mt)

Principal data source 

and notes (decimal degrees 

WGS 84)

Volcanogenic Cu(-Zn-Co-Ag-Au) massive sulfide deposits—Continued

Outokumpu 

(Kylylahti)

Outokumpu 

(Luikonlahti)

Outokumpu 

(Saramäki)

Outokumpu (Vuonos)

Gaiskoye (Gai and 

Gai East)

Ishkinino

Ivanovskoe

Komsomolskoye

Letneye

Sibaiskoye

Kure (Asikoy)

Turner Albright

Finland

Finland

Finland

Finland

Russia

Russia

Russia

Russia

Russia

Russia

Turkey

United States

62.72

63.60

62.94

62.76

51.46

51.45

52.45

51.23

51.06

52.70

41.81

42.00

29.00

27.40

29.13

29.09

58.49

58.28

57.92

56.97

59.37

58.63

33.69

–123.76

7.8

7.7

3.4

5.9

300.0

1.0

10.0

25.0

10.0

100.0

13.6

4.0

0.240

0.120

0.090

0.150

0.020

0.200

0.400

0.020

0.090

0.130

0.300

0.050

0.019

0.009

0.003

0.009

0.060

0.002

0.040

0.005

0.009

0.130

0.041

0.002

Altona Mining Ltd. (2013)

Eilu (2012)

Eilu (2012)

Eilu (2012)

Prokin and Buslaev (1999)

Herrington and others (2005)

Mosier and others (2009)

Prokin and Buslaev (1999)

Prokin and Buslaev (1999)

Prokin and Buslaev (1999)

Mosier and others (2009)

Mosier and others (2009)

Sea-floor Fe-Mn(-Ni-Cu-Co-Mo) nodules

Clarion-Clipperton 

zone

Cook Islands EEZ

Pioneer

Pacific Ocean

Pacific Ocean

Indian Ocean

10.00

–10.00

–14.00

–140.00

–170.00

74.00

21,100.0

5,122.0

700.0

0.209

0.410

0.100

44.00

21.00

0.700

Hein and others (2013)

Hein and others (2013)

Banakar (2010)

Sea-floor Fe-Mn(-Co-Mo-REE) crusts

Pacific prime 

crust zone

Nameless

Unicorn

Maderia-Tore Rise 

(MTR)

Afanasiy-Nikitin 

seamount

Pacific Ocean

Atlantic Ocean

Atlantic Ocean

Atlantic Ocean

Indian Ocean

15.00

35.20

34.60

36.00

–3.00

160.00

–15.00

–14.50

–15.00

83.00

7,533.0

71.0

130.0

1,100.0

180.0

0.664

0.340

0.350

0.350

0.580

50.00

0.240

0.450

3.800

1.044

Hein and others (2013)

Muiños and others (2013)

Muiños and others (2013)

Muiños and others (2013)

Parthiban and Banakar (1999)

1

Data for Tenke Fungurume include results for 7 deposits: Fwaulu, Fungurume, Kansalawile, Kwatebala, Mambilima, Mwadinkomba, and Tenke.

2

Data for Kalgoorlie Nickel include results for 10 deposits: Big Four, Black Range, Bulong, Ghost Rocks, Goongarrie Hill, Goongarrie South, Highway, 

Kalpini, Siberia, and Siberia North.

3

Cobalt data from D.M. Peterson, Duluth Metals Ltd., written commun. to J.F. Slack, March 25, 2013.

4

Size and grade estimated.
Fluorine

By Timothy S. Hayes, M. Michael Miller, Greta J. Orris, and Nadine M. Piatak

Chapter G of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–G

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter G

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................G1

Introduction..................................................................................................................................................G2

Uses, Applications, and Consumption ............................................................................................G3

Geology.........................................................................................................................................................G4

Geochemistry......................................................................................................................................G4

Fluorine in Rocks .......................................................................................................................G4

Fluorine in Water .......................................................................................................................G4

Mineralogy........................................................................................................................................G10

Deposit Types ...................................................................................................................................G12

Carbonatite-Related Fluorspar Deposits ............................................................................G13

Alkaline-Intrusion-Related Fluorspar Deposits .................................................................G13

Alkaline-Volcanic-Related Epithermal Fluorspar Deposits ..............................................G15

Mississippi Valley-Type Fluorspar Deposits .......................................................................G16

Salt-Related Carbonate-Hosted Fluorspar Deposits ........................................................G18

Fluorspar Deposits Related to Strongly Differentiated Granites ....................................G20

Subalkaline-Volcanic-Related Epithermal Fluorspar Deposits .......................................G21

Fluorspar Deposits Apparently Conformable Within Tuffaceous 

Limy Lacustrine Sediments ......................................................................................G22

Residual Fluorspar Deposits .................................................................................................G23

Fluorine Recovery From Brines ............................................................................................G23

Fluorspar Deposit Grades and Tonnages by Deposit Class .............................................G23

Resources and Production ......................................................................................................................G25

Identified Resources in the United States ...................................................................................G25

Identified Resources in Other Countries ......................................................................................G26

Undiscovered Resources in the United States ...........................................................................G27

Undiscovered Resources in Other Countries ..............................................................................G28

Production and Consumption ........................................................................................................G28

Exploration for New Deposits .................................................................................................................G30

Environmental Considerations ................................................................................................................G31

Sources and Fate in the Environment ..........................................................................................G31

Mine Waste Characteristics ..........................................................................................................G32

Human Health Concerns .................................................................................................................G33

Ecological Health Concerns ...........................................................................................................G35

Carbon Footprint...............................................................................................................................G36

Mine Closure.....................................................................................................................................G36

Problems and Future Research ..............................................................................................................G37

Acknowledgments....................................................................................................................................G37

References Cited.......................................................................................................................................G38

Appendix G1. Selected Fluorspar Districts, Deposits, and Prospects of the World .....................G53

Suggested citation:

Hayes, T.S., Miller, M.M., Orris, G.J., and Piatak, N.M., 2017, Fluorine, chap. G of Schulz, K.J., DeYoung, J.H., Jr., 

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. G1– G80, http://dx.doi.org/

10.3133/pp1802G.
Fluorine

By Timothy S. Hayes, M. Michael Miller, Greta J. Orris, and Nadine M. Piatak

Abstract

Fluorine compounds are essential in numerous chemical 

and manufacturing processes. Fluorspar is the commercial 

name for fluorite (isometric CaF2 ), which is the only fluorine 

mineral that is mined on a large scale. Fluorspar is used 

directly as a fluxing material and as an additive in different 

manufacturing processes. It is the source of fluorine in the 

production of hydrogen fluoride or hydrofluoric acid, which 

is used as the feedstock for numerous organic and inorganic 

chemical compounds.

The United States was the world’s leading producer 

of fluorspar until the mid-1950s. In the mid-1970s, the 

U.S. fluorspar mining industry began to decline because of 

foreign competition. By 1982, there was essentially only a 

single U.S. producer left, and that company ceased mining in 

1996. Consumption of fluorspar in the United States peaked in 

the early 1970s, which was also the peak period of U.S. steel 

production. Since then, U.S. fluorspar consumption has 

decreased substantially; the United States has nonetheless 

increased its imports of downstream fluorine compounds, such 

as, in order of tonnage imported, hydrofluoric acid, aluminum 

fluoride, and cryolite. This combination of no U.S. production 

(until recently) and high levels of consumption has made the 

United States the world’s leading fluorspar-importing country, 

in all its various forms.

The number of fluorspar-exporting countries has 

decreased substantially in recent decades, and, as a result, the 

United States has become dependent on just a few countries 

to supply its needs. In 2013, the United States imported the 

majority of its fluorspar from three countries, which were, in 

descending order of the amount imported, Mexico, China, and 

South Africa.

Geologically, in igneous systems, fluorine is one of a 

number of elements that are “incompatible.” These incompatible elements become concentrated in the residual magma 

while the common silicates crystallize upon magma ascent 

and cooling, leading to relatively high fluorine concentrations 

in the more evolved or differentiated igneous rocks and in 

hydrothermal deposits associated with those evolved igneous 

rocks. In sedimentary rocks, fluorine’s highest concentrations 

are found in phosphorites because fluorine substitutes for 

hydroxyl ions in apatite, which leads to fluorine concentrations 

of, typically, from 2 to 4 weight percent in phosphorites. 

Because of the presence of fluorine, phosphate fertilizer 

manufacturers can produce a fluorosilicic acid byproduct. 

Most deposits mined for fluorine are hydrothermal, however, 

and consist of fluorine minerals that precipitated from hot 

water. Magmatic brines and brines from deep within sedimentary basins that have high concentrations of dissolved fluoride 

are the mineralizing fluids for various types of hydrothermal 

fluorspar deposits. Relatively dilute hydrothermal fluids that 

formed in some volcanic rocks can also transport sufficient 

fluoride to form a high-grade fluorspar deposit. Fluorite has 

low solubility in a common range of hydrothermal temperatures, particularly from about 160 degrees Celsius (°C) down 

to 60 °C. The increasing fluorite solubility below 60 °C partly 

explains why some water with exceptionally high levels of 

dissolved fluorine are found even at ambient temperatures in 

evaporitic lake basins in some East African Rift valleys in 

Kenya and Tanzania. The geologic conditions that led to the 

high concentrations there are known to exist in a number of 

other places in the world as well, including, perhaps, places 

in the Basin and Range province of the United States.

Eight minerals or mineral groups have sufficient fluorine 

in their structures to be considered as possible ores of the 

element; they are bastnaesite (also spelled bastnäsite) and 

other fluorocarbonates, cryolite, sellaite, villiaumite, fluorite, 

fluorapatite (in phosphorites), various phyllosilicates, and 

topaz. Fluorite is currently the only mineral that is mined for 

fluorine, and no mineral except fluorite is likely to become 

a source of commercially produced fluorine as a primary 

product as long as supplies from relatively thick and highgrade fluorite deposits continue to be available.

At least seven classes (which include one subclass) 

of hydrothermal fluorite deposits are recognized; they are 

classified according to their tectonic and (or) magmatic 

settings, as follows: (1) carbonatite-related fluorspar 

deposits; (2) alkaline-intrusion-related fluorspar deposits; 

(3) alkaline-volcanic-related epithermal fluorspar deposits; 

(4) Mississippi Valley-type fluorspar deposits (and a 

subclass of salt-related carbonate-hosted fluorspar deposits); 

(5) fluorspar deposits related to strongly differentiated granites; 
G2 Critical Mineral Resources of the United States—Fluorine

(6) subalkaline-volcanic-related epithermal fluorspar deposits; 

and (7) fluorspar deposits that appear to be conformable within 

tuffaceous limy lacustrine sediments. An eighth class (not 

hydrothermal) is that of fluorspar deposits concentrated in 

soils and weathered zones; that is, residual fluorspar deposits. 

Generally, fluorspar deposits related to strongly differentiated granites have larger tonnages and lower grades than 

carbonatite-related fluorspar deposits, which, in turn, have 

larger tonnages and lower grades than fluorspar vein deposits 

from various other classes.

The United States has a few identified resources of 

fluorspar, most notably the Klondike II property in the IllinoisKentucky fluorspar district located about 8 kilometers southwest of Salem, Kentucky, which has a large vein that contains 

at least 1.6 million metric tons at a grade of 60 percent CaF2

(Feytis, 2009). Additional fluorspar resources of lower grade 

but larger tonnage have been identified at Hicks Dome in the 

Illinois-Kentucky fluorspar district and at Lost River near the 

western tip of the Seward Peninsula in Alaska, along with a 

couple of dozen smaller, higher grade resources.

Internationally, new mines that either opened before 

the beginning of 2013 or were scheduled to open soon after 

that time include the Nui Phao tungsten-fluorspar-bismuthcopper-gold deposit in northern Vietnam; the St. Lawrence 

project in Newfoundland, Canada, which is located in a 

well-known fluorspar district; the Bamianshan deposit, which 

is related to a strongly differentiated granite in northwestern 

Zhejiang Province, China, near some of that Province’s 

large, subalkaline-volcanic-related epithermal veins; and the 

Nokeng project in South Africa, which is also related to a 

strongly differentiated granite. Other deposits in northwestern 

Australia, Nevada (United States), Norway, South Africa, and 

Sweden have been identified and could be put into production 

within just a few years.

Among undiscovered resources, an interesting possibility 

might be to produce a fluorine product from evaporitic, highfluorine, high-pH sodium-carbonate brines like Lake Magadi 

(Kenya) and Lake Natron (Tanzania) in Africa’s Eastern Rift 

Valley. In addition, apparently conformable fluorspar deposits 

in tuffaceous limy lacustrine sediments, such as those in Italy, 

are likely to occur in similar young alkalic volcanic settings 

elsewhere in the world.

Modern geophysical and geochemical exploration techniques have typically not been brought to bear in exploration 

for new fluorspar deposits, although such techniques are likely 

to be used in future exploration. The tendency for fluorine to 

dissolve in significant concentrations in water at low temperature allows both surface water and groundwater to be used as 

sampling media in geochemical exploration. Evolved-graniterelated fluorspar deposits may be particularly susceptible to 

geophysical exploration methods because crystalline rocks that 

form a basement to sedimentary sections can be approximately 

defined with gravity and magnetic methods, and magnetitebearing skarns can be directly detected with magnetic surveys.

Environmental considerations of fluorine mining focus 

especially on drinking water, where high fluorine concentrations can lead to tooth decay; dental and skeletal fluorosis; 

and bone and cartilage conditions, including genu valgum, 

which is the crippling bone deformity more commonly 

known as knock knee. Trace amounts of other elements in 

fluorspar ores are a concern at some deposits; for example, 

high beryllium concentrations in alkaline-volcanic-related 

epithermal deposits or high cadmium concentrations 

associated with Mississippi Valley-type and salt-related 

carbonate-hosted fluorspar deposits.

Future research might include testing whether fluorine 

can be extracted economically from high-pH, sodiumcarbonate brines and exploring for new occurrences of 

apparently conformable fluorspar deposits in tuffaceous limy 

lacustrine sediments outside of the Latium Region of Italy. 

Other promising new areas of research could be studies of 

fluorspar deposit fluid inclusion compositions by quadrupole 

mass spectrometry, by noble gas mass spectrometry on 

irradiated fluid inclusions, or by chlorine isotopes, while also 

measuring the chemistry of the same fluid inclusions either by 

bulk crush-and-leach methods or by laser ablation-inductively 

coupled plasma mass spectrometry. Advanced studies of 

fluid inclusion chemistry could be applied beneficially to 

some of the enigmatic large epithermal fluorspar veins at 

various places in the world, where they might determine 

those deposits’ possible relationships to igneous intrusions, 

or to dissolved salt, or to heated meteoric water in volcanic 

sections, or perhaps to all three. This knowledge could help 

focus new exploration.

Introduction

Fluorine is an essential element for a number of industries, including aluminum production, steelmaking, gasoline 

refining, glass manufacturing, and the production of enamels, 

insulating foams, refrigerants, and uranium for nuclear 

power. The United States, which is the world’s second-ranked 

consumer of fluorine products (Miller, 2012), obtains most of 

its supply through imports from Mexico. China is the world’s 

leading fluorspar producer. It produced more than 60 percent 

of the world’s output in 2012 (Miller, 2013). Fluorspar, which 

is the commercial name for fluorite (isometric CaF2) (fig. G1), 

is the only mineral that is mined at large scale for fluorine. The 

importance of fluorine in numerous civilian and military uses 

and the dependence of the United States on a limited number 

of foreign suppliers are reasons why fluorine is considered to 

be a critical element for the United States. Beginning in World 

War II, or just after, and ending in 2009, fluorspar was judged 

to be of such importance that it was one of the strategic and 

critical minerals that were stockpiled by the U.S. Government. 

Starting in 2001, substantial decreases in exports of fluorspar 

from China have led to decreases in the world supply.
Introduction G3

Uses, Applications, and Consumption

Fluorspar is used directly as a fluxing material in 

ceramics, glassmaking, and steelmaking; as an additive in 

welding rod coatings; and for other, minor, uses. Fluorspar 

is the feedstock for fluorine chemicals, most of which are 

manufactured after fluorspar is converted to hydrogen fluoride; 

that is, hydrofluoric acid (HF). The major uses for HF are 

the manufacture of aluminum fluoride, synthetic cryolite, 

and fluorocarbons, but it is also used in the production of 

high-octane gasoline and in concentrating uranium; in glass 

etching, metal pickling, and oil and gas well treatment; and 

in the electronics industry. HF is used in the manufacture of a 

host of fluorine chemicals used in agrichemicals, dielectrics, 

metallurgy, mouth washes and toothpastes, pharmaceuticals, 

and wood preservatives.

There are three principal market grades of fluorspar—

acid, ceramic, and metallurgical. Specifications for the acid 

grade and metallurgical grade are generally well defined, 

but the requirements set by individual consumers may vary 

in detail.

Acid-grade fluorspar (sometimes called “acidspar”) is 

defined as containing more than 97 percent CaF2, although 

material with a lower CaF2 content can be used by manufacturers of HF (for example, those in Russia) if the remaining 

impurities are acceptable. Major impurities include calcium 

carbonate and silica content, which commonly have 

maximum allowable contents of 1.0 percent each. Limits are 

normally placed on other minor but relevant impurities, such 

as arsenic, phosphates, sodium chloride (NaCl), and sulfide 

sulfur or total sulfur. Arsenic and phosphate impurities are 

the most problematic because they are difficult to remove 

during ore beneficiation and because they have detrimental 

effects on HF production. If purchased as a wet flotation 

concentrate, the maximum allowable moisture content of the 

fluorspar is 10 percent water (H2O); if purchased as a dried 

concentrate, the specification for acid and ceramic grades 

is usually a maximum of 0.10 percent H2O. Particle size 

requirements for fluorspar may vary by customer, but generally the specification calls for the bulk (85 to 95 percent) 

to be able to pass through a 150-micrometer (μm; or 

0.15-millimeter [mm]) screen.

In the United States, the use of ceramic-grade fluorspar 

has declined, and some former consumers of ceramic-grade 

fluorspar now use acid-grade fluorspar. The two ceramic 

grades for fluorspar are no. 1 (90 to 95 percent CaF2 ) and 

no. 2 (85 to 90 percent CaF2). Some domestic and foreign 

fluorspar suppliers still offer ceramic-grade fluorspar, but 

because practically every ceramic-grade user has its own 

specifications, suppliers tend to tailor their products to 

specific consumers or to offer several product grades that 

fall between metallurgical grade and acid grade. Customers 

may specify limits on the amounts of calcium carbonate, 

ferric oxide, lead sulfide, zinc sulfide, and silica present in 

the fluorspar.

In the United States, metallurgical-grade fluorspar 

(“metspar”) generally contains a minimum of 80 percent 

CaF2 and can range as high as 93 percent CaF2. Steel mills 

are the major users of metspar, and their purchase specifications limit the content of silica to between 4 and 10 percent, 

and that of calcium carbonate, to between 4 and 9 percent. 

The standard size specifications for metallurgical-grade 

fluorspar gravel are that 100 percent of the gravel must be 

able to pass through a 63-mm screen and that a maximum of 

15 percent must be able to pass through a 9.5-mm screen. In 

markets outside of the United States, the acceptable level of 

CaF2 content may be lower (depending on the country), and 

the acceptable level of silica content may be higher than in 

the U.S. specification.

0 1 2 CENTIMETERS

0 0.5 1 INCH

Fluorspar Figure 1 Figure G1. Photograph of a fluorite specimen from the 

Number 1 (Minerva) Mine, Cave-in-Rock subdistrict, IllinoisKentucky fluorspar district, showing the paragenesis (sequence 

of mineral deposition), including (1) yellow fluorite number 2, 

(2) a redissolution surface visible particularly at the upper right 

beneath the blue fluorite, (3) pyrobitumens (black coating), 

and (4) blue fluorite (which is paragenetically equivalent to 

purple fluorite number 2 over most of the district). Worldwide, 

fluorite is the only mineral mined on a large scale solely for its 

fluorine content. Photograph courtesy of the Mineralogical and 

Geological Museum at Harvard University, Bannister Collection, 

MGMH 124614, © Copyright 2012, President and Fellows of 

Harvard College. All rights reserved.
G4 Critical Mineral Resources of the United States—Fluorine

Geology

Geochemistry

Fluorine is one of the five halogen elements, along with 

chlorine (the most common), bromine, iodine, and astatine 

(which is the rarest). Fluorine has only one known chemical 

valence: –1. It has an ionic radius of 1.36 angstroms (Å), 

which allows it to substitute readily for the hydroxyl ion 

(which has an ionic radius of 1.40 Å) in mineral structures 

(Fleischer and Robinson, 1963). Fluorine substitutes less 

commonly for chloride and oxygen ions.

Fluorine in Rocks

In igneous systems, fluorine is one of a number of 

elements that are largely incompatible, in that they do not 

readily go into the common silicate minerals that crystallize 

from the melt. As a result, they become concentrated in the 

residual magma during differentiation. Early in differentiation, 

as mafic silicates and calcium-feldspars crystallize from the 

melt, the fluorine and silica (and typically As, B, Be, Cs, 

Dy, K, Li, Lu, Mn, Mo, Nb, Sb, Sc, Sn, Ta, Tb, Th, Tl, U, 

W, Y, Yb, and Zr) increase in concentration in the residual 

magma (Hildreth, 1981). This happens in both subalkaline and 

alkaline magmas, although the alkaline magmas may achieve 

higher fluorine concentrations (table G1; Shawe, 1976, p. 5). 

Fluorine’s incompatibility broadly leads fluorite to be found 

in economic concentrations close to and within differentiation 

end-product rock types throughout the world, such as granites, 

syenites, and carbonatites, and their shallowly intrusive and 

extrusive equivalents. Shawe (1976) described and discussed 

several additional trends of fluorine in igneous rocks, some of 

which are not wholly predictable from the element’s general 

incompatibility. Typical igneous-rock fluorine concentrations 

are given in table G1.

Because of fluorine’s incompatibility, it may also occur in 

high concentrations in volcanic gases as well as in sublimates 

formed from the gases (White and Waring, 1963, table 1). 

Rocks altered by such gases similarly contain substantial 

amounts of fluorine; the fluorine is substituted mainly into 

hydroxyl sites in alteration micas and clay minerals, but in 

some places, the abundant fluorine occurs as fluorite.

In sedimentary rocks, fluorine’s highest syngenetic 

concentrations (concentrations formed at the time of sedimentation) are found in phosphorites because of the element’s 

substitution for hydroxyl ions in apatite, which is the principal 

phosphorous-bearing mineral in phosphorites; this process 

yields fluorapatite. Phosphorites typically have greater than 

2 weight percent fluorine and may (rarely) have more than 

4 weight percent (Gulbrandsen, 1966, table 1; Cheney and 

others, 1979, tables 3 and 6). Concentrations of fluorine 

in typical sedimentary rocks are given in table G1.

Fluorine geochemistry in metamorphic rocks is not well 

known. Fluorine is present primarily in biotite and muscovite 

in metamorphic rocks, and it tends to be concentrated in micas 

as water is lost during prograde metamorphism. Pyroxeneplagioclase granulites may contain fluorine-rich micas (up to 

3.3 weight percent fluorine in biotite) (Bose and others, 2005). 

A few concentrations of fluorine in typical metamorphic rocks 

are given in table G1.

Fluorine in Water

Most deposits mined for fluorine are hydrothermal (that 

is, the fluorine minerals have precipitated from hot water), 

so understanding the geochemistry of fluorine in water is 

important for understanding how fluorine deposits are formed. 

Table G2 is a listing of some typical concentrations of fluorine 

in various types of water. Surprisingly, fluorine can dissolve in 

high concentrations (concentrations of more than 1,000 parts 

per million [ppm]) in low-temperature water with high pH 

and a low calcium concentration (Kilham and Hecky, 1973; 

Jones and others, 1977). This is especially true for waters in 

contact with rocks that have high fluorine contents, especially 

young volcanic rocks. In such rocks, the glassy components 

commonly have the highest fluorine concentrations and 

are inherently unstable at surface conditions, releasing 

fluorine as they are altered. As with the other halogens at 

low temperature, fluorine occurs commonly as a free ion in 

water, whereas there is a tendency for greater complexing at 

higher temperatures.

Fluoride, like many other dissolved components, is found 

at high concentrations in brines of various origins. Magmatic 

brines and brines from deep within sedimentary basins are 

potential mineralizing fluids for various types of hydrothermal 

fluorspar deposits. Another important feature of fluorine 

geochemistry in water is that fluorite has minimum solubility 

in a common range of hydrothermal temperatures, particularly 

from about 160 degrees Celsius (°C) down to about 60 °C 

(fig. G2). Its solubility increases with increasing temperature, 

as is the case for many dissolved components. Experiments by 

Richardson and Holland (1979), however, show that, unlike 

most minerals, the solubility of fluorite increases as temperature drops below 60 °C down to 0 °C (fig. G2). Over this 

temperature range, the tendency of the mineral to precipitate 

from solution decreases as the temperature decreases. 

Fluorine’s high solubility in cool, calcium-poor, high-pH 

water from young volcanic rocks is related to this retrograde 

solubility. The high fluorine concentrations are found specifically in waters poor in calcium and high in carbonate, where 

fluorite does not control fluorine’s solubility.
Geology G5

Table G1. Fluorine concentrations in various types of rocks.

[Without extremes** means that statistical outliers have not been included in calculating the mean. They are shown as part of the range. no., number; ppm, 

part per million; b.d.l., below method detection limit; —, no data]

Rock type (location) No. of 

samples

Fluorine concentration (ppm) Source(s)1

Range Mean

Igneous rocks—Intrusive

Alkalic ultramafic 7 200 to 2,700 1,100 3, 10

Gabbro, diabase 26 50 to 1,100 420 6

Granodiorite and granite, without extremes** 93 20 to 2,700 810 6

Tertiary intrusives, diorite to granite (central Colorado), without extremes** 592 90 to 33,900 1,170 15

Alkalic intrusives 65 200 to 2,250 1,000 6

Alkalic intrusives 100 — 2,640 13

Granite (mostly Silver Plume, central Colorado), without extremes** 459 70 to 260,000 1,100 15

Granite, (mostly Pikes Peak, central Colorado), without extremes** 284 60 to 170,000 1,116 15

Igneous rocks—Extrusive

Picrite basalts, (Eastern Rift, Kenya and Tanzania) 5 480 to 1,400 904 11

Olivine basalts (Eastern Rift, Kenya and Tanzania) 18 290 to 1,300 900 11

Basalt 130 20 to 1,060 360 6

Andesite 77 b.d.l. to 780 210 6

Andesite 85 — 630 13

Phonolite 14 860 to 1,490 930 6

Phonolite (Eastern Rift, Kenya and Tanzania) 2 2,100 to 2,200 2,150 11

Trachyandesite (Eastern Rift, Tanzania) 1 — 1,400 11

Nephelinites (Eastern Rift, Kenya and Tanzania) 20 1,000 to 4,900 2,480 11

Peralkaline trachyte (Eastern Rift, Kenya and Tanzania) 4 1,700 to 3,600 3,070 12

Melilite, leucitite, and carbonatite (Western Rift, Uganda) 4 1,400 to 3,000 2,500 1, 4

Alkali rhyolite (Eastern Rift, Kenya) 4 1,700 to 6,800 3,870 2

Metamorphic rocks

General metamorphic 69 60 to 1,500 380 6

Schists and gneisses (central Colorado, mostly

Idaho Springs Formation), without extremes**

816 50 to 81,000 1,180 15

Sedimentary rocks

Sandstones, without extremes** 49 10 to 880 180 6

Graywacke 17 — 40 8

Carbonaceous or dark sandstones 11 — 400 8

Varicolored sandstones 23 — 320 8

Pale, nearly white, sandstones 12 — 280 8

Shales, without extremes 79 10 to 7,600 800 6

Pierre Shale 22 560 to 880 682 5

Limestones, without extremes** 98 b.d.l. to 1,210 220 6

Dolostones 14 110 to 400 260 6

Anhydrite rock 3 b.d.l. to 350 121 7

Halite rock 7 b.d.l. to 130 31 7

Phosphorite (Phosphoria Formation, United States) 60 — 31,000 9

Phosphorite (8 mining areas worldwide) 8 30,500 to 41,000 36,560 14

1

Sources of analyses: 1, Holmes and Harwood (1932); 2, Bowen (1937); 3, Higazy (1954); 4, von Knorring and DuBois (1961); 5, Tourtelot (1962); 

6, Fleischer and Robinson (1963); 7, Stewart (1963); 8, Pettijohn (1963, p. S16); 9, Gulbrandsen (1966); 10, Bell and Powell (1969); 11, Gerasimovskiy 

and Savinova (1969); 12, MacDonald and others (1970); 13, Shawe (1976); 14, Cheney and others (1979); and 15, Wallace (2010).
G6 Critical Mineral Resources of the United States—Fluorine

Table G2. Concentrations of fluorine and other elements in waters.—Continued

[Deposit types: 1. carbonatite-related, 2. alkaline-intrusion-related, 3. alkaline-volcanic-related epithermal, 4a. Mississippi Valley-type (MVT), and 4b. saltrelated carbonate-hosted. Chemical abbreviations: F, fluorine; CO3

, carbonate; HCO3, bicarbonate; HF, hydrogen fluoride; SO4, sulfate; Cl, chlorine; SiO2, 

silica; Ca, calcium; Fe, iron; Mg, magnesium; Na, sodium; K, potassium; Cu, copper; Pb, lead; Zn, zinc; CaCO3, calcium carbonate (calcite). calc’d, value 

was calculated using SOLVEQ software (Reed, 1982) and assuming charge balance and saturation with quartz, muscovite, dolomite, fluorite, pyrite, and galena or 

sphalerite. Units of measure: °C, degree Celsius; ppm, part per million; equiv. wt. %, equivalent weight percent; g/kg, gram per kilogram; μS/cm, microsiemens 

per centimeter. n.a., not applicable; —, not determined or not reported; i.d., identification; no. or n, number]

Sample type and area

Number

of 

samples

Associated 

deposit type or 

rock type(s)

pH

Conductivity

(μS/cm at 

25 °C)

F

(ppm)

CO3, aqueous, 

dissolved

(ppm)

CO3

(ppm)

HCO3

(ppm)

SO4

(ppm)

Natural waters 

River water — n.a. — — 0.05 to 2.7 — — — — 

Seawater — n.a. — — 0.03 to 1.4 — — — — 

Mountain region runoff—Rhoads 

Fork of Rapid Creek, Black Hills,

South Dakota, United States

1 Limestone

and

alluvium

7.9 466 <0.10 — — — 2.2 

Near-recharge groundwater in limestone—Site i.d. 440612103152001,

Well No. 10 in Madison aquifer, 

Rapid City, South Dakota,

United States

1 Limestone 7.53 343 0.2 — — — 11 

Basinal groundwater in limestone—

Site i.d. 440415101093002, well

in Madison aquifer, Midland, 

South Dakota, United States

1 Limestone 6.69 — 2.8 — — 150 800 

Basinal groundwater in sandstone—

Well P17 in Lance/Fox Hills 

sandstone, Gillette, Wyoming, 

United States

1 Sandstone 8.18 1,810 3.2 — 2 982 1.6 

Groundwater evaporite brine—Well 

in Paradox sandstone (“posthalite” water), Paradox Basin, 

Grand County, Utah, United States

1 Sandstone

and

anhydrite

4.6 95,000 25 — — 1,010 4 

Mineralizing water for MVT

fluorspar deposits—From fluid 

inclusion measured data, and 

estimates calculated at 120 °C, 

Illinois-Kentucky fluorspar

district, United States

n.a. Granitic 

basement, 

sandstone,

and

dolostone

4.55

calc’d

— 26

calc’d

(CO2 )

17,589

— 2,558 

Measured fluid inclusions in fluorite—

“Complex” sedimentary brine at 

about 140 °C, Berbes subdistrict, 

Asturias Province, Spain. equiv. 

wt. % NaCl: 83; cations: 34

1 4b; metamorphic 

rocks and 

limestone

— — — — — — — 

Surface water—Lake Tulusia of 

the Momela Lakes, Eastern Rift 

Valley, Tanzania

1 Alkalic

volcanic

rocks

— 14,390 437 — (HCO3+CO3

)

189

422 

Water interstitial to newly precipitated, evaporitic trona and 

villiaumite—Borehole H4, 

Lake Magadi, Kenya

1 Trachyte and 

alluvial 

sediments

10.4 — 1,610 — 85,400 2,560 973 

Hypothetical hydrothermal mineralizing water 

Calculated mineralizing water for 

MVT fluorspar deposits after adding hypothetical magmatic gas 

with 0.5 g/kg CO2

 and 0.5 g/kg 

HF and heating to 300 °C, Illinois-Kentucky fluorspar district, 

United States

n.a. 4a 3.7

calc’d

— 2,030

calc’d

(CO2 

) 20,100

calc’d

— 105

calc’d
Geology G7

Table G2. Concentrations of fluorine and other elements in waters.—Continued

[Deposit types: 1. carbonatite-related, 2. alkaline-intrusion-related, 3. alkaline-volcanic-related epithermal, 4a. Mississippi Valley-type (MVT), and 4b. saltrelated carbonate-hosted. Chemical abbreviations: F, fluorine; CO3

, carbonate; HCO3, bicarbonate; HF, hydrogen fluoride; SO4, sulfate; Cl, chlorine; SiO2, 

silica; Ca, calcium; Fe, iron; Mg, magnesium; Na, sodium; K, potassium; Cu, copper; Pb, lead; Zn, zinc; CaCO3, calcium carbonate (calcite). calc’d, value 

was calculated using SOLVEQ software (Reed, 1982) and assuming charge balance and saturation with quartz, muscovite, dolomite, fluorite, pyrite, and galena or 

sphalerite. Units of measure: °C, degree Celsius; ppm, part per million; equiv. wt. %, equivalent weight percent; g/kg, gram per kilogram; μS/cm, microsiemens 

per centimeter. n.a., not applicable; —, not determined or not reported; i.d., identification; no. or n, number]

Cl

(ppm)

SiO2

(ppm)

Ca

(ppm)

Fe

(ppm)

Mg

(ppm)

Na

(ppm)

K

(ppm)

Cu

(ppm)

Pb

(ppm)

Zn

(ppm)

Alkalinity

(ppm as 

CaCO3)

Source(s)1

Natural waters—Continued

— — — — — — — — — — — 15

— — — — — — — — — — — 15

3.1 9 66 0.005 23 1 0.6 <0.001 <0.001 <0.003 219 20

1 — 38 — 19 2.6 1.6 0.002 — 0.01 171 20

28 35 270 — 66 25 9.7 — — — — 8

30 17 2.7 17 0.4 370 4 <0.001 0.001 <0.003 — 6

241,000 — 52,700 750 39,200 5,990 18,800 6 6 60 — 1

92,200

calc’d

59 4,994 .00094

calc’d

6,099 42,016 3,310

calc’d

— 0.14

calc’d

0.68

calc’d

— 11

~130,000 — 12,541 — 2,131 48, 987 6,424 — 92 404 — 18

646 5 3.2 — 6.1 4,600 726 — — — — 4

68,400 513 — — — 115,000 1,430 — — — — 5

Hypothetical hydrothermal mineralizing water—Continued

92,200

calc’d

58.5

calc’d

3,930

calc’d

.00094

calc’d

6,020

calc’d

42,016 3,310

calc’d

— 0.14

calc’d

0.68

calc’d

— 11
G8 Critical Mineral Resources of the United States—Fluorine

Table G2. Concentrations of fluorine and other elements in waters.—Continued

[Deposit types: 1, carbonatite-related; 2, alkaline-intrusion-related; 3, alkaline-volcanic-related epithermal; 4a, Mississippi Valley-type (MVT); and 4b, saltrelated carbonate-hosted. Chemical abbreviations: F, fluorine; CO3

, carbonate; HCO3, bicarbonate; HF, hydrogen fluoride; SO4, sulfate; Cl, chlorine; SiO2, 

silica; Ca, calcium; Fe, iron; Mg, magnesium; Na, sodium; K, potassium; Cu, copper; Pb, lead; Zn, zinc; CaCO3, calcium carbonate (calcite). calc’d, value 

was calculated using SOLVEQ software (Reed, 1982) and assuming charge balance and saturation with quartz, muscovite, dolomite, fluorite, pyrite, and galena or 

sphalerite. Units of measure: °C, degree Celsius; ppm, part per million; equiv. wt. %, equivalent weight percent; g/kg, gram per kilogram; μS/cm, microsiemens 

per centimeter. n.a., not applicable; —, not determined or not reported; i.d., identification; no. or n, number]

Sample type and area

Number

of 

samples

Associated 

deposit type or 

rock type(s)

pH

Conductivity

(μS/cm at 

25 °C)

F

(ppm)

CO3, aqueous, 

dissolved

(ppm)

CO3

(ppm)

HCO3

(ppm)

SO4

(ppm)

Fluorspar mining area waters 

Surface runoff, Northern Pennine 

Orefield, United Kingdom

5 4 — — 0.05 to 0.17 — — — — 

Surface runoff draining mine waste 

area (with fluorite), Northern Pennine Orefield, United Kingdom

25 4 — — 0.05 to 2.3 — — — — 

Domestic water supplies, Northern Pennine Orefield, United 

Kingdom

4 4 — — 0.12 to 0.52 — — — — 

Mine waters from the Frazer’s 

Grove fluorite mine and surrounding area, Northern Pennine 

Orefield, United Kingdom

31

(3 sites)

4 3.8 to 

7.6

335 to 

3,188

— — — up to 

186

34 to 

2,501

Surface water upstream of the 

Frazer’s Grove Mine area, 

United Kingdom

2

(1 site)

4 6.7 to 

7.9

89 to 91 — — — 16 to 

18

9.9 to 

17.9

Mine waters from the Moscona

Mine, Spain

8 4b 7.6 to 

8.2

413 to 

1,779

0.5 to 3.5 — — — — 

Surface spring from the Moscona 

Mine area, Spain

1 4b 7.8 451 0.5 — — — — 

Surface water upstream of the 

Moscona Mine, Spain

1 4b 7.9 405 0.5 — — — — 

Waste waters from the Moscona 

Mine, Spain

29 4b 7.3 to 

8.9

— 0.27 to 4.8 — — — — 

Groundwaters in Coimbatore 

district, Tamil Nadu, India

5 Granite and 

granodiorite

— — 0.8 to 5.9 — — — — 

Mine pit water, Chandidongri fluorite mine, Madhya Pradesh, India1 2 9.0 — 0.3 — 4.0 33.6 15.0 

Groundwater near Chandidongri 

fluorite mine, Madhya Pradesh, 

India

4 2 7.0 to 

9.0

140 to 

242

0.04 to 0.4 — 6.0 13 to 

88

up to 

15

Groundwater in Andhra Pradesh,

India

60 Granites 7.5 to 

8.7

423 to 

1,640

0.78 to 6.1 — 104 to 199 533 to 

1,415

22 to 

96

Surface water, groundwater, and 

mine waters, Kerio Valley, Kenya

28

(14 

sites)

1, 2, or 3 6.0 to 

9.5

42 to 812 0.8 to 306 — — — 23 to 

66

Mine waters from the Lovozero 

niobium mine, Russia

10 2 9.6 to 

12.0

— 750 to 

15,000

— (HCO3+CO3)

1,425 to 2,666

(n=8)

up to 

200

1

Sources: 1, White (1965); 2, Hall and Friedman (1963); 3, Kraynov and others (1969); 4, Kilham and Hecky (1973); 5, Jones and others (1977); 

6, Henderson (1985); 7, Fuge and Andrews (1988); 8, Busby and others (1991); 9, Leach and others (1991); 10, Davies (1994); 11, Plumlee and others (1995); 

12, Chatterjee and Mohabey (1998); 13, Narayana and others (2000); 14, Johnson and Younger (2002); 15, Kabata-Pendias and Mukherjee (2007); and references therein; 16, Jacks and others (2005); 17, Roqueñí and others (2005); 18, Sánchez and others (2009); and 19, Arveti and others (2011); 20, U.S. Geological Survey (2016).
Geology G9

Table G2. Concentrations of fluorine and other elements in waters.—Continued

[Deposit types: 1. carbonatite-related, 2. alkaline-intrusion-related, 3. alkaline-volcanic-related epithermal, 4a. Mississippi Valley-type (MVT), and 4b. saltrelated carbonate-hosted. Chemical abbreviations: F, fluorine; CO3

, carbonate; HCO3, bicarbonate; HF, hydrogen fluoride; SO4, sulfate; Cl, chlorine; SiO2, 

silica; Ca, calcium; Fe, iron; Mg, magnesium; Na, sodium; K, potassium; Cu, copper; Pb, lead; Zn, zinc; CaCO3, calcium carbonate (calcite). calc’d, value 

was calculated using SOLVEQ software (Reed, 1982) and assuming charge balance and saturation with quartz, muscovite, dolomite, fluorite, pyrite, and galena or 

sphalerite. Units of measure: °C, degree Celsius; ppm, part per million; equiv. wt. %, equivalent weight percent; g/kg, gram per kilogram; μS/cm, microsiemens 

per centimeter. n.a., not applicable; —, not determined or not reported; i.d., identification; no. or n, number]

Cl

(ppm)

SiO2

(ppm)

Ca

(ppm)

Fe

(ppm)

Mg

(ppm)

Na

(ppm)

K

(ppm)

Cu

(ppm)

Pb

(ppm)

Zn

(ppm)

Alkalinity

(ppm as 

CaCO3)

Source(s)1

Fluorspar mining area waters—Continued

— — — — — — — — — — — 7

— — — — — — — — — — — 7

— — — — — — — — — — — 7

up to 258 — 52 to 673 <0.1 to 

68

9.4 to 

215

9.1 to 38 4.9 to 

33.7

— — <0.1 to 

111

— 14

13 to 24 — 6.1 to 8.1 0.5 to 0.6 1.8 to 3.0 5.0 to 6.3 1.1 to 1.6 — — 0.2 — 14

— — — <0.1 — — — — — <0.05 to 

2.7

— 17

— — — <0.1 — — — — — <0.05 — 17

— — — <0.1 — — — — — <0.05 — 17

— — — — — — — — — — — 17

— — — — — — — — — — — 16

30.0 — 8.8 — — — — — — — — 12, 13

10 to 30 36 8.0 to 18 3.6 4 to 5 25.5 1.5 — — — — 12

10 to 96 22 to 90 26 to 99 25 to 234 5 to 16 — — — 832 to 

1,920

19

11 to 48 16 to 67 — 0.1 to 0.8 — — — 0.01 to 

0.08

— — 26 to 414 10

8 to 85 140 to 

13,000

up to 1 — 0.6 to 

140

3,680 to 

26,000

35 to 250 0.3 to 0.9

(n=3)

0.02 to 

0.03 

(n=3)

0.4

(n=1)

— 3
G10 Critical Mineral Resources of the United States—Fluorine

Mineralogy

Figure G3A shows the eight possible ore minerals or 

mineral groups that are fluorine-rich and either are, were, or 

might someday become fluorine ore minerals. Although all eight 

are possible ore minerals of fluorine, at present only deposits of 

fluorite are mined specifically for their fluorine content.

Bastnaesite (hexagonal (Ca,La,Nd)(CO3)F) is representative of several fluoro-carbonate minerals that occur in or are 

associated with carbonatites and alkaline ultramafic igneous 

complexes. Bastnaesite is the major mineral from which rareearth elements (REEs) are recovered at a number of carbonatites around the world, including at the Mountain Pass Mine

in California. The Mountain Pass Mine was the world’s 

leading producer of REEs from the 1960s to the 1980s and 

the sole mine that commercially produced REEs in the 

United States in the latter half of the 20th century. The mine 

was closed in 2002 but opened again in 2010; it was then 

placed on care-and-maintenance status in 2015 (Molycorp, 

Inc., 2015a, b). Mountain Pass does not recover fluorine from 

the bastnaesite.

Cryolite (monoclinic Na3(AlF6 )) is a mineral that is 

seldom found outside of pegmatites. It was produced as a 

fluorine source from the Ivigtut pegmatite in Sermersooq 

municipality, Greenland, until 1987, when mining there 

ceased (Singleton, 1989, p. 273).

Sellaite (tetragonal MgF2) is mined at the Suran 1 deposit 

in the southern Ural Mountains in Russia, and fluorine is 

recovered from it. The same deposit contains much larger 

tonnages of fluorite, and the sellaite occurs only where 

the fluorine-bearing hydrothermal solutions encountered 

magnesium-rich diabase (that is, shallowly intrusive basaltic 

rock) in the fault or vein walls (Ellmies and others, 1999). 

Apparently, the magnesium in the basaltic wall rock was used 

in precipitating sellaite, rather than fluorite. The situation is 

not geologically unique; sellaite is found in several different 

fluorite-bearing deposits in dolomite host rocks—for example, 

at Mount Bischoff, Tasmania, Australia (Kwak and Askins, 

1981); at the Shepherd and Murphy Mine in the Moina 

district, Tasmania, Australia (Bottrill and Baker, 2008); at 

the Xianghuapu Mine in Linwu County, Hunan Province, 

China (Yuzhou, 1990); and at the Yaroslavskoye tin deposit 

in Primorskiy Kray, Russia (Bailey, 1980). In all these 

cases, magnesium from the dolomite was probably used in 

precipitating sellaite in preference to fluorite. In a few other 

occurrences, sellaite is found in rocks where calcium had 

earlier been removed by hydrothermal alteration, thus favoring 

the sellaite over fluorite (Imai and others, 1985; Pfaff and 

others, 2012).

Villiaumite (isometric NaF) occurs within trona (monoclinic Na3

(HCO3 )(CO3 ) • 2H2O,) in eastern Africa. Trona is 

the raw material for sodium carbonate (that is, washing soda 

or soda ash). In some valleys of the East African Rift system, 

trona is precipitating because of evaporation of the closed 

basin lakes. With high enough concentrations of villiaumite 

in the trona ore, it could become a feasible byproduct and 

might be recovered as a source of fluorine. The trona deposits 

of eastern Africa are located close to peralkaline to alkaline 

volcanic rocks. An analysis of water pumped from within 

newly precipitated Lake Magadi trona is given in table G2.

110

100

90

80

70

60

50

40

30

20

10

0

Fluorite deposited, in milligrams of CaF2 per kilogram of water

Temperature, in degrees Celsius

Fluorspar Figure G2

260 200 150 100 50 0

0.44 M CaCl2+MgCl2

2.68 M NaCl

0.33 M CaCl2+MgCl2

3.01 M NaCl

0.22 M CaCl2+MgCl2

3.34 M NaCl

0.11 M CaCl2+MgCl2

3.67 M NaCl

0.022 M CaCl2+MgCl2

3.93 M NaCl

4.00 M NaCl

Figure G2. Graph showing the solubility of fluorite (CaF2) as a 

function of temperature for complex Na-Ca-Mg-Cl (sodium-calciummagnesium-chloride) brines from ambient temperatures (25 degrees 

Celsius [°C]) to 260 °C. Total ionic strength for each combination of 

salts was 4 moles per liter. At each molarity (M) of CaCl2 + MgCl2, the 

ratio of calcium to magnesium was 9 to 1. Curvatures in the dashed 

parts of the curves are inferred. Except for pure sodium chloride 

(NaCl) brine compositions, the solubility curves at different complex 

chloride salt molarities all have a maximum temperature of fluorite 

deposition below which the solubility (the opposite of deposition) of 

fluorite increases with decreasing temperature. At those temperatures 

(those below the maximum temperature of fluorite deposition), fluorite 

is said to have retrograde solubility; that is, fluorite solubility increases 

as temperature decreases. First, temperatures near the laboratory 

maxima of fluorite deposition have commonly been measured in fluid 

inclusion microthermometry studies of actual fluorspar deposits (for 

example, Richardson and Pinckney, 1984). Second, the retrograde 

solubility at low temperatures is explanation for relatively high 

concentrations of fluorine in water measured in several surface water 

or shallow groundwater environments (for example, Lake Magadi, 

Kenya; Jones and others, 1977). Graph is from Richardson and Holland 

(1979, fig. 5). CaCl2

, calcium chloride; MgCl2, magnesium chloride
Geology G11

Bastnaesite

Hexagonal

(Ca,La,Nd)(CO3)F

2 to 9 wt. % F.

Carbonatites and

alkaline intrusions

or hydrothermal.

? byproduct F from

mining rare-earth

elements

3. Alkaline-volcanicrelated epithermal

veins, breccia-fills,

and carbonate

replacements

Examples:

Aguachile,

Mexico

Bor Ondor,

Mongolia

?Kimwarer,

Kenya

2. Alkaline-intrusionrelated veins, skarns,

and carbonate

replacements

Examples:

St. Lawrence district,

Canada

Jamestown district,

Colorado, U.S.A.

Yermakovskoye,

Russia

Dulaan Ovoo,

Mongolia

Huahe,

China

1. Carbonatiterelated

Examples:

Amba Dongar,

India

Okorusu,

Namibia

Mato Preto,

Brazil

Kizilcaören,

Turkey

Karasug,

Russia

?Kimwarer,

Kenya

? Hicks Dome,

Illinois, U.S.A.

7. Apparently

conformable

in tuffaceous, 

limy lacustrine

sediments

Examples:

Castel GiulianoPianciano,

Italy

Santo Maria

di Sala,

Italy

Rome, Oregon

occurrence

?Thomas Range,

Utah, U.S.A.

? Sumochagan

Obo, China

6. Subalkalinevolcanic-related

epithermal

veins and

replacements

Examples:

Yangjia-Wuyi,

China

Kalangui,

Russia

? Las Cuevas,

Mexico

?Thomas Range,

Utah, U.S.A.

? Sumochagan

Obo, China

5. Skarns, greisens,

veins, and 

replacements

related to

strongly

differentiated

granites

Examples:

Voznesenka,

Russia

Lost River,

Alaska, U.S.A.

Shizhuyuan,

China

El Hammam,

Morocco

Vergenoeg,

South Africa

Buffalo

South Africa

Nui Phao,

Vietnam

4a. Mississippi

Valley-type

Examples:

Illinois-Kentucky

fluorspar district,

U.S.A.

Northern Pennine

Orefield and

Southern Pennine

Orefield, U.K.

Marico district,

South Africa

4b. Salt-related

carbonate-hosted

Examples:

Asturias,

Spain

Rossignol,

France

? EncantadaBuenavista,

Mexico

Sellaite

Tetragonal

MgF2

61 wt. % F.

Mined together

with fluorite in

hydrothermaldeposits with

Mg-rich wallrocks

or wallrocks

without Ca

Fluorapatite

Hexagonal

Ca5(PO4)3F

3.1 to 4.9 wt. % F.

Fluorosilicic acid byproduct

of mining phosphorites

and nelsonites and

phoscorites from 

ultra-alkalic igneous

complexes; P used

in fertilizers

Topaz

Orthorhombic

Al2(SiO4)(F,OH)2

13 to 20 wt. % F.

Pegmatites, hydrothermally altered

rocks, and a topaz

schist, Front Range,

Colorado, U.S.A.

Phyllosilicates

For example, biotite

Monoclinic

K(Mg,Fe)(AlSi O3 10)(OH,F)2

0 to 7.5 wt. % F.

Hydrothermally altered

rocks; for example,

Henderson porphyry

molybdenum deposit,

Colorado, U.S.A. (Gunow

and others, 1980)

Fluorite

Isometric

CaF2

41.6 to 

48.7 wt.% F

Villiaumite

Isometric

NaF

45 wt. % F.

? byproduct of

trona mining

Cryolite

Monoclinic

Na3(AlF6)

54 wt. % F.

Pegmatites;

natural cryolite

probably not

economic

Classes of primary hydrothermal fluorspar deposits

Possible fluorine ore minerals

Fluorspar Figure G3.

Alkaline-related

AB

Figure G3. Chart showing A, eight minerals or mineral groups from which fluorine has been produced or may be produced in the future, and B, a preliminary 

classification of hydrothermal fluorspar deposits by tectonic and magmatic association. Ca, calcium; F, fluorine; Mg, magnesium; P, phosphorus; 

U.S.A., United States of America; wt. %, weight percent; ?, possible
G12 Critical Mineral Resources of the United States—Fluorine

Fluorite is the most common mineral that contains high 

concentrations of fluorine (fig. G3). Fluorite is typically 

almost transparent, and it has perfect octahedral cleavage. 

It is softer than quartz, harder than calcite, and unreactive 

in weak acids, so it is typically readily distinguishable from 

calcite and quartz, with which it is commonly associated. 

Fluorite occurs as a minor gangue mineral in many types of 

hydrothermal ore deposits; for example, epithermal gold, 

silver, and uranium deposits; porphyry molybdenum (±tungsten) deposits; pegmatites; iron oxide-copper-gold deposits; 

and typically zinc-dominated Mississippi Valley-type (MVT) 

deposits. In the deposit types shown in figure G3B, fluorite is 

the major economic mineral, although coproduct or byproducts Au, barite, Be, Bi, Cu, Li, Mo, Nb, Pb, REEs, Sb, Sn, W, 

and Zn, can be found in the various types of fluorspar deposits.

In many fluorspar deposits, it is clear that fluorine in a 

hydrothermal solution reacted with calcium from the wall 

rocks to form fluorite. In limestone wall rocks, the reaction 

for fluorite precipitation can be modeled as follows:

2HF+CaCO3→CaF2+H2O+O2 (1)

Where the above reaction occurs and calcium is 

conserved (because the density of calcite is 2.71 grams 

per cubic centimeter [g/cm3] and the density of fluorite is 

3.18 g/cm3), new open space (or porosity) results. In fact, 

at standard temperature and pressure, about 15 percent of 

the original limestone (calcium carbonate [CaCO3]) volume 

becomes new porosity. New crystals of fluorite grow into the 

open space as terminated crystals, and the result is that, in 

most carbonate replacement deposits, fluorite occurs as wellformed cubic or cube-octahedral crystals (fig. G1) of many 

different colors; blue, green, purple, white, yellow, and clear 

are typical. Reactions similar to (1) can be written for other 

calcium sources, such as dolomite or plagioclase, most with 

the effect that new porosity is created if calcium is conserved.

Fluorapatite (hexagonal Ca5(PO4 )3F) currently accounts 

for only minor amounts of fluorine production, but it constitutes a very large potential supply if needed in the future. 

Fluorapatite is the major mineral in sedimentary phosphorite 

deposits, which typically have a fluorine grade of 2 to 4 weight 

percent fluorine. Phosphorite is a sedimentary rock composed 

principally of phosphate minerals. Phosphorite deposits can 

exceed 1 billion metric tons of ore. Phosphorites are mined in 

many places in the world as a source of fertilizer. Byproduct 

fluorine is already captured at some of the operations and is 

either converted to synthetic cryolite (a form in which it can 

be used readily in aluminum refining) or to fluorosilicic acid 

for use in drinking water fluoridation. Currently, the supply 

of fluorine obtained from phosphorites is determined by the 

demand for fertilizer—once the fertilizer demand is met, no 

additional byproduct fluorine is produced.

Fluorine is found in a number of silicate minerals, but, 

apparently, no feasible process to recover fluorine from any 

silicate structure has been developed. Topaz (orthorhombic 

Al2

(SiO4)(F,OH)2) is a fluorine-bearing mineral found in 

pegmatites, some hydrothermally altered rocks, and an 

apparently unique topaz-rich schist in Colorado’s Front Range 

(Sheridan and others, 1968).

Fluorine occurs in at least trace concentrations in most 

mica and clay (phyllosilicate) minerals; biotite (monoclinic 

K(Mg,Fe)3

(AlSi3O10 )(OH,F)2) (fig. G3A) is one example 

of fluorine in phyllosilicates. In the phyllosilicates, fluorine 

substitutes for the OH− ion in interlayer sites between the 

sheets of linked alumina octahedrons and silica tetrahedra. 

Theoretically, it should be easier to recover fluorine from such 

mineralogic sites than from topaz, in which the fluorine and 

hydroxyl ions link together adjacent alumina octahedrons; 

however, the relatively small amount of fluorine available 

in the phyllosilicates (fig. G3A) discourages any attempts to 

recover fluorine from them. The maximum amount of fluorine 

measured in phyllosilicates to date is 7.5 weight percent from 

biotite from a hydrothermal vein from the Henderson porphyry 

molybdenum deposit in Clear Creek County, Colorado 

(Gunow and others, 1980).

Deposit Types

Fluorspar deposits occur in a wide variety of geologic 

environments. The lower part of figure G3 (fig. G3B) is a 

preliminary classification of hydrothermal fluorspar deposits 

by their tectonic and (or) magmatic environments. All seven 

of the classes are hydrothermal, but there is a great variation 

of fluid temperatures and compositions among them, and 

they clearly represent various tectonic and magmatic settings, 

although there is some ambiguity in classification of certain 

deposits. The seven classes of primary hydrothermal fluorspar 

deposits shown in figure G3B are discussed below.

Worldwide, many fluorspar deposits are found closely 

associated in space with igneous rocks of several different 

types. Understanding distinctions between several types of 

igneous rocks is needed in order to recognize the various 

magmatic and tectonic environments that are represented. 

Following are explanations for the terms subalkaline, 

peralkaline, alkaline, and carbonatitic, with generalities 

about the tectonic environments represented by each of 

these compositional types of igneous rocks.

Peralkaline igneous rocks are igneous rocks in which the 

molecular proportions of sodium and potassium exceed the 

proportions of aluminum (molar Na2O+K2O>Al2O3). In such 

rocks, the amounts of sodium and potassium are greater than 

can be accommodated in sodium and potassium feldspars and 

must be taken up in other crystallizing minerals; for example, 

aegerine (monoclinic NaFeSi2O6, a pyroxene mineral) or 

riebeckite (monoclinic Na2(Fe2+, Mg)3Fe2

3+Si8

O22(OH)2, 

an amphibole mineral). Subalkaline rocks are those that 

have more than enough aluminum to make the sodium and 

potassium feldspars, with the excess combining with calcium 

and silica to crystallize calcium-bearing plagioclase. In most 

alkaline rocks, however, the molecular proportions of sodium 

and potassium are so great that not even the minerals with 

anomalously high levels of sodium and potassium—that is, 
Geology G13

sodic and potassic pyroxenes and amphiboles, such as aegerine 

and riebeckite, respectively—can take up the excess of alkali 

molecules; consequently, minerals called feldspathoids, such 

as nepheline (hexagonal (Na,K)AlSiO4 ), sodalite (isometric 

Na4(Si3Al3 )O12Cl), leucite (tetragonal K(AlSi2 )O6 ), and 

melilite (tetragonal (Ca,Na)2(Al,Mg)(Si,Al)2O7 ), which have 

more alkalies and aluminum than do feldspars, crystallize 

instead of the alkali feldspars. Alkaline melts also commonly 

have concentrations of silica (SiO2) that are too small to 

make any quartz after the crystallization of feldspars or 

feldspathoids; these melts are thus termed “undersaturated.” 

The end member of the alkaline rocks has very little silica 

at all. These are carbonatites—rare igneous rocks that are 

composed of more than 50 volume percent carbonate minerals 

that crystallized from a magma. Carbonatites occur together 

with syenites and other nepheline-rich intrusive rocks in 

complexes within which the carbonatite is typically the last 

rock type intruded.

As a gross generality, subalkaline igneous rocks are 

the dominant types intruded or erupted in both continental 

and island arcs along convergent plate boundaries, most 

peralkaline rocks are instead unrelated to active compressional 

orogenies at convergent plate margins (they are “anorogenic”), 

and alkaline igneous rocks and carbonatites are found intruded 

through mostly thick continental crust remote from convergent 

margins. Both the anorogenic peralkaline intrusives and the 

alkaline complexes may relate broadly or directly to rifts 

through continental crust.

For each different type of magmatic association—

subalkaline, peralkaline, alkaline, or carbonatitic—the associated fluorspar deposits are, almost without exception, related 

to the most differentiated (latest intruded and crystallized, least 

iron and magnesium-rich, and most silica rich for other than 

the carbonatites) igneous rocks of the igneous province. So, 

for subalkaline magmatic provinces, the fluorspar deposits are 

found almost exclusively with granites that have high contents 

of incompatible elements. The granites associated with the 

fluorspar deposits typically have only biotite as a ferromagnesian mineral. For the peralkaline anorogenic igneous 

provinces, the fluorspar deposits are also associated with 

granites, particularly aegerine granites and riebeckite granites. 

For the alkaline complexes, the fluorspar deposits are most 

closely related to syenites, and only rarely to trachytes, or 

pyroxene-nepheline rocks. And for the carbonatite complexes, 

the fluorspar deposits are found most closely associated with 

the carbonatites themselves.

For association with volcanic rocks rather than intrusive 

rocks, the situation is parallel. The fluorspar deposits are 

nearly ubiquitously associated with rhyolites. They are nearly 

nowhere associated with the basalts.

Carbonatite-Related Fluorspar Deposits

Fluorite deposits associated with carbonatites and other 

ultra-alkaline rocks occur on every continent (fig. G4). In 

such deposits, it is common that the carbonate (commonly 

calcite) that makes up the carbonatite is partially replaced or 

veined by fluorite or is cut by a stockwork of fluorite veinlets. 

The country rock in contact with carbonatite may be altered 

(fenitized) by hydrothermal fluids and also mineralized with 

fluorite. At the Fission (Richardson) deposit in Ontario, 

Canada, and the Speewah deposit in Kimberley, Western 

Australia, Australia, the deposits have the general form of 

a vein or dike, although the carbonate minerals appear to 

be igneous and to contain minerals that are distinctive to 

igneous carbonatites, such as bastnaesite, parisite (rhombohedral Ca(Ce,Nd,La)2

(CO3)2 ), or pyrochlore (isometric 

(Na,Ca)2Nb2O6 (OH,F)). Two examples of carbonatiteassociated fluorspar deposits that have been well studied 

are the Okorusu deposit located in the Otjozondjupa Region 

of Namibia (Kogut and others, 1997; Shivdasan-Gebhardt 

and Hagni, 2008) and the Amba Dongar deposit located in 

Gujarat State, India (Deans and others, 1972; Gwalani and 

others, 1993; Simonetti and Bell, 1995; Simonetti and others, 

1995; Palmer and Williams–Jones, 1996; Doroshkevich and 

others, 2009). At the Amba Dongar deposit, studies of the 

fluid compositions (Palmer and Williams-Jones, 1996) and 

the radiogenic isotopes (Simonetti and Bell, 1995) suggest 

that, despite temperatures of fluorite mineralization of 160 °C 

and less, there was a small component of igneous-derived 

magmatic water mixed with dominant meteoric water (that is, 

groundwater derived ultimately from rain and recharged from 

the surface) in the mineralizing fluids.

Alkaline-Intrusion-Related Fluorspar Deposits

Alkaline-intrusion-related fluorspar deposits are 

fluorite-bearing deposits that are in close spatial association 

with granites that contain aegerine and (or) riebeckite, or 

are in association with syenites, any of those igneous rocks 

commonly being porphyritic (Teng and Strong, 1976; Lykhin 

and others, 2010). A distinction is arbitrarily drawn between 

these types of deposits and a large number of other fluorspar 

deposits that are associated with extrusive alkali rhyolites 

or trachytes (that is, volcanic equivalents of syenites) where 

intrusive rocks are not found nearby. It is somewhat common 

for deposits of both classes to contain either bertrandite 

or phenakite and, at some places, lithium minerals, such 

as lithium-rich white mica (Lykhin and others, 2010; 

Obolenskiy and others, 2010). Some also contain economic 

concentrations of niobium, REEs, tantalum, and zirconium 

(Kovalenko and Yarmolyuk, 1995). The Yermakovskoye 

deposit in Buryatiya Republic, Russia, which was the first 

of the Russian alkaline- intrusion-related deposits to be 

mined, has small skarns—that is, bodies of calcium-silicate 

minerals with additional valuable minerals that have replaced 

limestone adjacent to intrusive rocks—and bodies with veinlet 

stockworks of fluorite-phenakite or of fluorite-bertrandite, 

and in which the fluorite commonly totals 20 volume percent 

(Kislov and others, 2010; Lykhin and others, 2010). Phenakite 

is rhombohedral Be2

SiO4, and bertrandite is orthorhombic 

Be4Si2O7(OH)2, so these deposits are commonly also deposits 
G14 Critical Mineral Resources of the United States—Fluorine 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10°

20°30°40°50°

80°

70°

70°

80°

60°

Lost River

Crystal Mountain

Meyers Cove

Rome, Oregon, occurrence

Northgate district Jamestown district

Thomas Range Browns 

Canyon district

Indian Peak district

Daisy Mine

Huckleberry Mine

Great Eagle MineWhite Eagle Mine

Taxco

district

Aguachile

Encantada-Buenavista

Las CuevasEl Refugio

Parral

district

Hicks Dome and Illinois-Kentucky fluorspar district

Sweetwater district

St. Lawrence

 district

Fission

(Richardson)

Mine

Mato Preto

Santa Catarina district

Cerros Negros

Northern Pennine Orefield

Southern Pennine Orefield

Argentolle Le Barlet

Le Burc

Asturias

Escaro

Osor

Sierra Gador district

El Hammam Jebal Tirremi

Silius Mine Bruncu Molentinu

Santo Maria di Sala

Castel GiulianoPianciano

Kizilcaören

Suran I

and II

Kimwarer

Jebel

Semeih

Okorusu

Marico

district

Zwartkloof

Buffalo

Vergenoeg Hlabisa

ChiocoDjanguire Mount Muambe Canxixe

Macossa

Kalat district

Amba

Dongar

Karadzhal

Takob

Suppatash

Kalangui

Chaakhol

Karasug

Auninsky Yermakovskoye

Naransky

Mushugai- Khudag

Berkh Adag

Zuum-Tsagaan-Del

Bor Ondor

Urgun

Sumochagan Obo Chenlou

Jianshan

Yinzishan

Zhuancum

Taolin Yanjia-Wuyi

Houshu

Hushan

Badu

Donggangshan

Huayuan

district Mahuaping

Ban Hong

Doi Tao

district

Moina

Yang Hak

 district

Voznesenka I and II

 Pogranichny

NuiPhao

Rossignol

Speewah

Taskaynar South Taskaynar

Huahe

Pakozd

Maoniuping

Shuangjiangkou

Shizhuyuan

Hammam Zriba

Dulaan Ovoo

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Fluorspar Figure G4.

Carbonatite-relatedAlkaline-intrusion-relatedAlkaline-volcanic-related epithermal Mississippi Valley-typeSalt-related carbonate-hostedRelated to strongly differentiated granitesSubalkaline-volcanic-related epithermalConformable in tuffaceous limy  lacustrine sedimentsClass unknown

Selected fluorspar deposits—Smaller text indicates reserves of less than 1.5 million metric tons of fluorite

EXPLANATION

Figure G4. World map showing locations of selected fluorspar deposits according to their tectonic and magmatic class as listed in figure G3B.
Geology G15

of beryllium. Other alkaline-intrusion-related deposits 

are located in the St. Lawrence district in Newfoundland, 

Canada, where large, discrete fluorspar veins cut across 

peralkaline granite (Van Alstine, 1944). Other districts, such 

as the Jamestown district in Boulder County, Colo., have ore 

contained mainly in breccia bodies that may grade laterally 

to veins (Goddard, 1946). Studies suggest that the fluids that 

formed these deposits had temperatures that ranged from 

500 °C down to about 100 °C (Nash and Cunningham, 1973; 

Strong and others, 1984). The fluids also had a great range of 

salinities. Those results and oxygen isotope studies suggest 

that the mineralizing fluids were mixtures of magmatic and 

meteoric waters, and the relatively high contents of REEs 

in the fluorite supports that conclusion (Strong and others, 

1984). At Yermakovskoye, magmatic fluids have been well 

documented (Reyf, 2004, 2008) and include (a) a magmatic 

brine highly enriched in beryllium, manganese, and molybdenum, with additional cerium, lanthanum, lead, and zinc 

concentrations, and (b) a carbon dioxide (CO2)-rich fluid 

enriched only in beryllium with a homogenization temperature 

of between 335 and 355 °C and relatively low salinity of 

2.4 to 8.2 equivalent weight percent NaCl. These two fluids 

seem likely to be the products of the boiling of a common 

ancestor hydrothermal fluid.

Alkaline-Volcanic-Related Epithermal 

Fluorspar Deposits

Northern Mexico and Kenya have fluorspar deposits 

and occurrences that are related to alkaline volcanic rocks. 

Examples of such deposits are located at Aguachile, Coahuila 

de Zaragoza, Mexico (McAnulty and others, 1963), and at 

Kimwarer, Elgeyo/Marakwet County, Kenya (Nyambok 

and Gaciri, 1975). These deposits are also in contact with or 

are hosted by carbonates (limestone or calcite marble). The 

igneous rock type that is most closely associated in time with 

ore at Aguachile is rhyolite porphyry, which occurs in a ring 

dike around a central plug of quartz-microsyenite porphyry. 

The microsyenite contains 7 volume percent riebeckite 

and another 8 volume percent chlorite that is an alteration 

product of former riebeckite (Levinson, 1962). Several other 

small districts are related to similar rocks in Coahuila de 

Zaragoza, Mexico, as, perhaps, are the deposits of the very 

large Encantada-Buenavista district of Mexico (Temple and 

Grogan, 1963; Kesler, 1977; González-Partida and others, 

2003). Fluorspar deposits in the Eagle Mountains of Hudspeth 

County, Texas, are related to igneous rocks of the same 

igneous province (Gillerman, 1948). The Hidalgo del Parral 

district in Chihuahua, Mexico, is also related to alkali rhyolites 

(Pickard, 1974).

The Kimwarer deposits in Kenya, though hosted in 

Precambrian calcite marbles and other gneissic rocks, are 

most closely related in time to Tertiary phonolites of the East 

African Rift volcanic rocks (Nyambok and Gaciri, 1975). 

Phonolite is a volcanic rock that consists essentially of 

alkali feldspar and a feldspathoid mineral (the feldspathoid 

mineral is nepheline for the rocks near Kimwarer). Because 

the phonolites are not in contact with fluorspar ore anywhere at 

Kimwarer, it is possible that these deposits are instead related 

to unexposed (still buried) carbonatitic rocks. Carbonatites 

are found in several places in or near the East African Rift 

valleys. In these volcanic-related deposits, low temperatures 

and generally low salinities suggest that the mineralizing 

fluids were dominated by meteoric waters. Fluid inclusion 

homogenization temperatures are generally no greater than 

175 °C, and salinities are commonly less than 10 equivalent 

weight percent NaCl (Kesler, 1977; González-Partida and 

others, 2003). Although intrusive rocks may be present, they 

may have served only as heat sources to cause convection 

of the groundwaters through fluorine-rich alkali rhyolite 

source rocks. Alternatively, and more likely, the magmas that 

contributed to the volcanic rocks also evolved a late-stage 

water-rich phase, which contributed to the mineralizing waters 

and thus at least small amounts of magmatic components to 

the deposits. Although the conditions and chemistry of the 

hydrothermal fluids were generally similar throughout this 

class, the geometry of mineralized bodies varies greatly. Some 

are discrete, large veins (Kimwarer). Others are carbonatehosted breccia pipes or breccias along contacts with intrusions 

(Aguachile). Yet others are peneconformable replacement 

bodies in carbonates, which are known in the United States 

as bedding-replacement deposits (Weller and others, 1952; 

Gillerman, 1948) and in Spanish-speaking countries as mantos 

(for example, the Encantada-Buenavista district in Mexico; 

Temple and Grogan, 1963). A peneconformable configuration 

means that the replacement bodies lie within and parallel to 

the bedding of the sedimentary host rocks, yet have lateral 

boundaries that cut across those host rocks.

Some have suggested that other fluorspar deposits of the 

Western United States and Mongolia are related to alkaline 

volcanic rocks (Lindsey, 1982; Lkhamsuren and Hamasaki, 

1998), but there is some doubt that the associated volcanic 

rocks should truly be called alkaline in the sense that the 

word is used above. Lindsey (1982) describes the deposits of 

the Thomas Range, Utah, as being related to rhyolites of the 

latest magmatic stage in the area—those rocks typically have 

a peralkaline composition. The rhyolites of the Thomas Range 

(the Topaz Mountain rhyolites) have phenocrysts of sanidine, 

biotite, and plagioclase, and lack any aegerine or riebeckite. 

They are alkali feldspar rhyolites, but the rocks are probably 

not truly of the alkaline series as that term is used above in 

defining the distinction between subalkaline, peralkaline, 

and alkaline.

The deposits of the Thomas Range are probably better 

assigned to the class of epithermal deposits related to 

subalkaline volcanic rocks (see below in the section titled 

Subalkaline-Volcanic-Related Epithermal Fluorspar Deposits). 

The rhyolite associated with fluorspar deposits in the Thomas 

Range contains topaz phenocrysts; it is very fluorine-rich. Some 

well-known fluorspar deposits in Colorado—in the Browns 

Canyon district of Chaffee County (Van Alstine, 1969) and in 
G16 Critical Mineral Resources of the United States—Fluorine

the Northgate district of Jackson County (Steven, 1960)—are 

also most closely associated in time with topaz rhyolites, 

which are not rocks in the alkaline series (Burt and others, 

1982); hence, those deposits are also probably best assigned 

to the class of epithermal deposits related to subalkaline 

volcanic rocks. The topaz rhyolites of western North America 

do not, however, date from times of the major andesitic 

volcanic fields and granodiorite batholiths like those of the 

Sierra Nevada Range. Rather, they post date such rocks in 

each area where they occur, and they belong to groups of rocks 

believed to have formed by partial melting of continental 

crust. Numerous deposits in Mongolia are described as being 

related to “trachytic rhyolite” and “rhyolite differentiated from 

trachy-basalts” (Lkhamsuren and Hamasaki, 1998). Whereas 

true trachytes (volcanic rocks containing alkali feldspar, minor 

iron-magnesium-silicate minerals, minor plagioclase, and no 

quartz) typically are alkaline, the term “trachytic” describes a 

texture of lath-shaped phenocrysts of alkali feldspar that are 

subparallel to one another in a volcanic rock; the texture does 

not require the rocks to be truly alkaline. The Late Mesozoic 

Mongolian volcanic rocks are not described as riebeckite- or 

aegerine-bearing, and those volcanic rocks again include 

some topaz rhyolites (Kovalenko and others, 1971). The very 

fluorine-rich topaz rhyolites are seemingly the rocks that are 

the most closely associated with fluorspar deposits. Bulnayev 

and Kaperskaya (1995) demonstrated that fluorspar deposits in 

Mongolia are most frequently associated with rhyolites among 

the volcanic rocks there.

Mississippi Valley-Type Fluorspar Deposits

Mississippi Valley-type (MVT) deposits of leadzinc(-fluorspar-barite) are among the best understood of all 

hydrothermal ore deposits. These deposits formed in carbonate 

aquifer rocks where there was interaquifer groundwater 

mixing of the following two brines: (a) a metalliferous and 

(or) fluorine-rich brine that escaped upward from a confined 

lower aquifer under artesian pressure, and (b) a second brine, 

probably more sulfide-rich (Wenz and others, 2012), that 

migrated laterally, also under artesian pressure, and was 

confined above, typically by a shale bed (Plumlee and others, 

1994, 1995; Bouch and others, 2006; Stoffel and others, 2008; 

Ingebritsen and Appold, 2012; Wenz and others, 2012). Both 

brines likely had their origins as deep basinal groundwaters. 

A variety of geohydrologic situations allow the brine of the 

lower aquifer to escape confinement and rise to mix with the 

second brine in the upper aquifer. The simplest situation that 

leads to mixing may be where fracturing connects the two 

aquifers, as in a vein deposit or a breccia body, regardless 

how the breccia originated. A second and very important 

mixing situation is where a confining bed pinches out laterally 

between the two aquifer units (a “shale edge”). Upward escape 

of lower-aquifer fluids will be focused adjacent to the pinchedout confining bed. This can result in very large deposits 

arrayed along and above the pinchout, stratigraphically at the 

level of the upper aquifer; this is the situation with most of 

the ores of the Southeast Missouri lead district, the Tri-State 

(Missouri-Kansas-Oklahoma) district, and the Pine Point 

district in the Northwest Territories, Canada (Brockie and 

others, 1968; Rhodes and others, 1984; Palmer and Hayes, 

1989). A number of other geohydrologic situations allow the 

fluids of the lower aquifer to escape and mix with the groundwater of upper aquifers, but these situations are less common. 

The studied brine mixtures (represented by fluid inclusions) 

from MVT deposits originated by evaporation of seawater, 

but the mixed fluids also show effects from dolomitization 

and sulfate reduction along the travel paths and generally 

show only small amounts of salt redissolution (Viets and 

others, 1997).

The type of deposits that occur in the Mississippi Valley 

in the United States resulted from water in gravity-driven 

regional groundwater flow systems that affected many tens 

of thousands of cubic kilometers of rock, with recharge in 

or near collisional orogenic mountain belts like the Ouachita 

Mountains (Leach and Rowan, 1986). Fluids traveled down 

through deep basins, then up, driven by artesian pressure from 

the elevated recharge area, across the cratonic margins towards 

the craton centers. The regional nature of the mineralizing 

systems is demonstrated in places where the sequence of 

mineral precipitation (that is, the paragenesis) of hydrothermal 

minerals that have partially filled many different kinds of rock 

porosity can be correlated from MVT district to district, and 

through all of the former aquifer rocks in between, across 

hundreds of kilometers (Hayes and others, 1990; Hayes and 

Anderson, 1992).

Districts of the MVT class that contain significant 

amounts of fluorite account for 3 of the 10 largest sources of 

fluorspar in the world (appendix G1). These are the IllinoisKentucky fluorspar district in the United States, the Northern 

Pennine Orefield and the Southern Pennine Orefield in the 

United Kingdom, and the Marico district of North West 

Province, South Africa.

Richardson and Pinckney (1984) established that 

oscillatory growth layering in fluorite crystals within the 

Cave-in-Rock subdistrict (Hardin County, Illinois) of the 

Illinois-Kentucky fluorspar district could be correlated from 

mine to mine throughout an area of about 8.5 by 2.5 kilometers (km). By correlating generations of hydrothermal 

carbonates, Hayes and Anderson (1992) extended the 

correlation of hydrothermal minerals throughout the IllinoisKentucky fluorspar district, then correlated that paragenesis 

with that of the Central Tennessee (Elmwood-Gordonsville) 

zinc district, and with the paragenesis of the Central Kentucky 

district. The total area with correlatable paragenesis of 

hydrothermal minerals partially filling the former porosity 

in carbonates from Early Ordovician through Mississippian 

age is 600 km east-west by 300 km north-south and averages 

slightly more than 2 km thick, thus yielding about 350,000 

cubic kilometers of rock affected by this single MVT system 

(Hayes and Anderson, 1992). The correlations may extend 

even farther—perhaps to the Upper Mississippi Valley 

district in southern Wisconsin (Heyl and West, 1982) and to 
Geology G17

the Findlay arch in northwestern Ohio (Carlson, 1983). The 

Illinois-Kentucky-Central Tennessee MVT system differs 

markedly from that of the Ozark Region of southern Missouri, 

northern Arkansas, southeastern Kansas, and northeastern 

Oklahoma, however, where no fluorite is found. Furthermore, 

in the Illinois-Kentucky fluorspar district, thin dikes and sills 

of mafic igneous rock are fairly numerous and a few small 

plugs of intrusive breccia occur. Hicks Dome in the northwestern portion of the district is thought to be underlain by a 

larger related igneous body, but that remains unproven.

In the Illinois-Kentucky fluorspar district, the dikes 

and sills are likely near-surface indicators of larger igneous 

masses at depth, and the igneous rocks related to these dikes 

are likely the cause of the district’s unique richness in fluorine 

among the MVT deposits. This district hosts 62 known dikes 

and sills that are 1 to 12 meters (m) thick and intruded along 

faults or bedding planes. It also hosts eight known “diatremes” 

(brecciated volcanic pipes) that are approximately upright and 

cylindrical with diameters from a few tens of meters to nearly 

200 m (Denny and others, 2008; Anderson and Sparks, 2012).

Almost all the district’s igneous rocks studied to 

date were originally ultramafic and alkalic. Where fresh, 

the dike and sill rocks are dark green and porphyritic to 

inequigranular. Their major primary minerals are olivine, 

phlogopite, melilite, garnet, apatite, perovskite, and magnetite 

with secondary serpentine, chlorite, and iron-rich calcite. 

Differentiates from the magmas that produced these dikes 

would be expected to be syenites and perhaps could include 

carbonatite. The xenoliths in the diatremes include aegerine 

syenite (Bradbury and Baxter, 1992, p. 15), and the rocks in 

places are nepheline-bearing (Lewis and Mitchell, 1987). The 

neodymium-samarium isochron age of MVT fluorite mineralization from five mines in the Cave-in-Rock subdistrict is 

277.0±15.6 mega-annum (Ma), or million years before present 

(Early Permian) (Chesley and others, 1994), which overlaps 

entirely the four best-dated igneous rocks in the district that 

would together give an age range of 266.5 to 273.4 Ma (Early 

Permian), inclusive of all uncertainties (Reynolds and others, 

1997; Fifarek and others, 2001). Where the igneous rocks are 

in contact with MVT mineralization, the phlogopite of the 

igneous rocks is hydrothermally altered to illite.

Hicks Dome appears to be the thermal center for 

hydrothermal mineralization in the Illinois-Kentucky fluorspar 

district. Deep drilling at Hicks Dome has encountered mineralization with elemental and mineralogical signatures that 

suggest alkalic igneous rock and possibly carbonatite, which 

agrees with the widespread but small occurrences of alkaline 

ultramafic rock in the greater district. The dome is underlain 

at depths of from 450 m to perhaps 1,000 m by fluorite-baritecemented breccias that contain anomalous concentrations 

of Be, Nb, Th, Ti, Y, and REEs (Brown and others, 1954; 

Kirkemo, 1978). Across the district, fluid inclusion homogenization temperatures in fluorite from growth zones late 

within the sequence of banding described by Richardson and 

Pinckney (1984) decline regularly from the center of Hicks 

Dome (175 °C), to the flanks of the dome (150 °C), to the 

Cave-in-Rock subdistrict (150 to 132 °C), to mines central 

to the district in Kentucky (130 to 120 °C), to the southern 

margin of the district (128 to 117 °C), to the most distant 

outliers of the district (<50 °C?) (Taylor and others, 1992). 

This concentric zonation of homogenization temperatures 

around Hicks Dome has the same generally concentric form 

as contours of lead isotopes from galena (Heyl and others, 

1966) and contours of silver and antimony content of galena 

(Hall and Heyl, 1968). Thus, the Illinois-Kentucky fluorspar 

district records a single large mineralizing system that was 

likely influenced in some way by magmatism at or near 

Hicks Dome. It is not inferred that magmatic fluids migrated 

outward from Hicks Dome; only that it was the highest 

temperature area of the district and probably has intrusive 

rocks at depth.

Plumlee and others (1995) used reaction path modeling 

to try to better understand the Illinois-Kentucky fluorspar 

district. They modeled a mineralizing fluid that had the 

measured compositions for major elements from fluid inclusions in fluorite (Hall and Friedman, 1963; Richardson and 

others, 1988) and assumed saturation with respect to dolomite, 

quartz, muscovite, pyrite, sphalerite, galena, and fluorite 

at 120 °C and a pH of 4.55. They could not reproduce the 

district’s MVT ores in terms of phase abundance or paragenesis until they modeled the addition of 0.5 gram of HF and 

0.5 gram of CO2 to each kilogram of the initial fluid. In the 

HF- and CO2-modified models, they found that any or all of 

the following resulted in copious amounts of fluorite precipitation: (a) cooling the fluid from 150 to 142 °C, (b) replacement of limestone, and (c) mixing of the fluid with another 

fluid with different concentrations of calcium and fluorine. 

Plumlee and others (1995) concluded that HF from alkaline 

igneous sources, particularly at Hicks Dome, was added 

to regional MVT metals-transporting fluids to produce the 

district’s fluorite-rich MVT ores. The fluorite-rich ores formed 

where alkaline igneous rocks and their evolved magmatic 

fluids intruded into an ongoing regional MVT mineralizing 

system that extended to at least the Central Tennessee and 

Central Kentucky districts.

Exposed or geophysically indicated alkalic magmatic 

source rocks have not been identified at either of the 

Pennine orefields nor in the Marico district, which are 

the other identified fluorite-rich MVT districts. Dunham 

(1948) speculated that the Weardale Granite, which forms 

the basement to the sedimentary host rocks of the Northern 

Pennine Orefield, was a heat source simply from decay of 

elevated concentrations of radioactive elements within it, 

and that convection of mineralizing meteoric groundwaters 

through the fracturing systems was set up around the Weardale 

Granite because of the anomalous heat. Bau and others (2003) 

suggested that sources of REEs and yttrium in the fluorite of 

the Southern Pennine Orefield were relatively pure marine 

sedimentary carbonates, and that, in contrast, the sources for 

the Northern Pennine Orefield were aluminosilicate rocks, 

which is consistent with the Weardale Granite being a source 

for REEs in the Northern Pennine Orefield fluorite.
G18 Critical Mineral Resources of the United States—Fluorine

The Bushveld Complex is located relatively near the 

Marico MVT district in South Africa. The large Buffalo 

fluorspar deposit (Limpopo Province) and the very large 

Vergenoeg fluorspar deposit (Gauteng Province) are located 

near the center of the Bushveld Complex and at its highest 

structural level (appendix G1). Buffalo and Vergenoeg appear 

to be related to strongly differentiated Bushveld Complex 

granites (deposit type 5 of fig. G3B). In a contact zone that 

surrounds the Bushveld intrusives, the Marico fluorspar ores 

have been metamorphosed by the heat from the intrusions 

(Martini, 1976), so it was not possible for Bushveld-age 

hydrothermal fluids from those granites to provide fluorine 

for the preexisting MVT fluorspar deposits.

Salt-Related Carbonate-Hosted Fluorspar Deposits

Leach and others (2005, appendix A, table A1) consider 

salt-related carbonate-hosted deposits to be part of their 

classification of MVT deposits and have given them an 

alternate subclass name (“salt dome related”). According 

to the authors, only 5 of their 247 example MVT deposits 

worldwide are said to be salt dome related. Indeed, these 

deposits have the same simple mineralogy, the same kinds of 

orebody geometries, similar parageneses, and probably the 

same types of local geohydrologic controls (various scenarios 

that produce interaquifer groundwater mixing) as the MVT 

deposits within the Mississippi Valley itself. In figure G3, saltrelated carbonate-hosted deposits are shown to be a subclass 

of MVT deposits. What MVT and salt-related carbonatehosted deposits do not share, however, is their regional and 

temporal geologic settings, along with other details that result 

from the differences in their settings.

Salt deposits themselves are commonly found on passive 

continental margins that formed in the arid tropics or in failed 

arms of the rifts extending inland from passive continental 

margins (aulacogens). Thus, in contrast with the MVT deposits 

of the Mississippi Valley, which are associated in time and 

distantly in space with collisional orogenic belts, the saltrelated carbonate-hosted deposits are in many places clearly 

associated with continental rifting, with passive margins, or 

with aulacogens. The salt-related classification originated 

with French scientists who studied deposits in Algeria and 

Tunisia where the salt is Triassic (Charef and Sheppard, 1987; 

Orgeval, 1994; Perthuisot and Rouvier, 1995; Sheppard and 

others, 1997); however, the best example of these types of 

deposits are from the Miocene to recent Red Sea margins 

(El Aref and Amstutz, 1983; Hayes and others, 2002). The 

U.S. Gulf of Mexico margin, where the Middle to Upper 

Jurassic Louann Salt is located, also has deposits of this type 

(Price and others, 1983; Kyle and Saunders, 1997; Saunders 

and Swann, 1994), but none has been proven economic 

to mine.

Differences between MVT deposits and salt-related 

carbonate-hosted zinc-lead-fluorite-barite deposits are important considerations for exploration. The sedimentary sections 

that host salt-related carbonate-hosted deposits contain 

evaporites, which typically form major parts of the local 

sedimentary section and, at places, exceed one-half of the total 

sedimentary thickness. Although evaporites are present in 

traces in the sedimentary sections of the MVT deposits of the 

Mississippi Valley, none appear to be as much as 5 percent of 

the thickness of the involved section anywhere within 100 km 

of a deposit (although adjacent basins have considerably 

more; for example, the Michigan Basin). Salt-related 

carbonate-hosted zinc-lead-fluorite-barite deposits differ 

from the MVT deposits by having paragenetic sequences of 

hydrothermal minerals that cannot be correlated for even as 

much as tens of kilometers. Connected secondary porosity 

needed for long-distance transport of hydrothermal brines does 

not generally exist in strongly fault-segmented continental rifts 

and early passive margins.

Although many of the salt-related zinc-lead-fluoritebarite deposits are located close to salt deposits, with some 

actually within the cap rocks of salt domes, the critical role 

of the salt is probably not in creating sites for mineralization, 

such as caprock hosts, nor in creating structural paths for the 

required interaquifer groundwater mixing. Rather, the critical 

role of salt in these deposits is probably that it dissolves in 

mineralizing brines to supply chloride and thus provides 

high solubilities of zinc, lead, iron, and barium as chloride 

complexes in those mineralizing brines (Hayes and others, 

2001, 2002). Spatial association with salt and the presence of 

fluid inclusions in ore minerals that demonstrate the process 

of dissolving preexisting salt to create an ore solution are true 

indicators of this class. Spatial association is not necessary if 

the fluid inclusions contain redissolved salt—as demonstrated 

from the Cl−

:Br− and Na+:Br− ratios (Carpenter, 1978; Hanor, 

1987; Kesler and others, 1995) —as a major component in the 

ore fluids.

Fluorspar deposits in the Asturias region of Spain are the 

best known examples of fluorine-rich deposits of this class 

as a result of some recent work on the composition of the 

deposits’ fluids (Sánchez and others, 2009). The geographic 

association of the deposits with Late Triassic salt in diapirs 

is shown in figure G5, which shows as well the locations of 

zinc-lead-(fluorspar-barite-copper-iron) deposits and siderite 

iron deposits of the enclosing and larger Cantabrian region 

(Velasco and others, 1994). The zinc-lead-(fluorspar-baritecopper-iron) deposits and the siderite iron deposits probably 

all originated from the same general mineralizing events; 

all these deposits are probably salt-related carbonate-hosted 

deposits (Grandia and others, 2003). In studies of the 

microthermometry and composition of fluid inclusions in 

the Asturias deposits, Sánchez and others (2009) identified 

two discrete fluids that mixed to precipitate fluorspar ore: a 

simple sodium-chloride-dominated groundwater with salinities of 0 to 8.2 equivalent weight percent, and a complex 

sodium chloride–calcium chloride brine with salinity of 

9 to 25 equivalent weight percent sodium chloride. They 

found that the complex higher-salinity brine had Cl−

:Br− and 

Na+:Br− ratios that demonstrated a large proportion of halite 

dissolution. A comparable degree of halite dissolution was 
Geology G19

nowhere measured among sphalerite fluid inclusions from 

12 different MVT districts, including all of the districts of the 

Mississippi Valley itself (Viets and others, 1997). Sánchez and 

others (2009) found that the Asturias complex halite dissolution brine also had very high metal contents; that is, up to 

500 ppm zinc, 170 ppm lead, and 480 ppm barium—or about 

one order of magnitude greater than those found in most other 

metalliferous basinal brines (Land, 1995; Hanor, 1997). The 

Asturias fluorspar deposits were also samarium-neodymium 

isotopically dated, yielding an age of 185±29 Ma (Early 

Jurassic) (Sánchez and others, 2006). This date demonstrates 

the association with rifting instead of collisional tectonics, 

because in the Early Jurassic Period, the Asturias region of 

Spain and the rest of Western Europe were in the process of 

renewed rifting along the middle Atlantic rift following upon 

the initial rifting that took place during the Permian Period 

(Yilmaz and others, 1996).

Many other fluorspar deposits of the western part of 

Europe appear to share many of the characteristics of the 

deposits of Asturias. Similar hypersaline brines at temperatures of from 81 to 165 °C and salinities of from 17 to as 

much as 26 equivalent weight percent NaCl were found in 

northeastern (Catalonia) Spain’s Berta, Osor, and Rigros 

veins (Cardellach and others, 2003; Piqué and others, 2008); 

in the Muscadroxiu-Genna-Tres Montis-S’Acqua Frida 

veins (Silius Mine) (Boni and others, 2009) and other veins 

in southern Sardinia, Italy (Belkin and others, 1984); in 

the Le Burc vein in the Massif Central (Midi-Pyrénées 

Region, France; Deloule, 1982); and in the Rossignol vein 

(Centre Region, France) and exhalative sediment (Sizaret and 

others, 2004). Four of these locations are shown in fig. G4. 

Each of these fluorite deposits has been isotopically or geologically constrained to have formed between the Early Jurassic 

Period and Early Cretaceous Period and is related to the 

renewed rifting of the middle Atlantic (Sizaret and others, 

2004; Munoz and others, 2005; Piqué and others, 2008).

The above deposits are all probably examples of the 

general salt-related carbonate-hosted class and related to 

rifting, even where the wall rocks are Hercynian basement 

rocks. A characteristic magmatic rock of some continental rifts 

or rifted margins is peralkaline, however, and, consequently, 

each of the deposits could instead be of the alkaline-intrusionrelated type, with the intrusive rocks buried and providing 

mainly just the heat to produce hydrothermal circulations. 

The deposits at Hansonburg in Socorro County, New Mexico, 

have been called MVT (Putnam and others, 1983; Norman and 

others, 1985), despite their setting adjacent to the Rio Grande 

Rift, which hosts a number of epithermal, alkaline-volcanicrelated fluorspar deposits (Van Alstine, 1976). Recent results 

demonstrate that the Hansonburg hydrothermal fluids had from 

13 to 49 percent asthenospheric magmatic chlorine (Partey 

and others, 2009), so the deposits are also alkaline-intrusionrelated. Deposits of the Encantada-Buenavista district in 

Coahuila de Zaragoza, Mexico (fig. G4) have been shown to 

have fluid inclusions with homogenization temperatures of 

from 75 to 170 °C and salinities of from 7.8 to 15.4 equivalent 

Vitoria

Bilbao

Santander

Berbes

Gijon

Aviles

Oviedo

Bayonne

San Sebastian

Pamplona

FRANCE

SPAIN

ATLANTIC OCEAN

Bay of Biscay

REOCIN MINE

TROYA

MINE

PAMPLONA

 FAULT

VENTANIELLA FAULT

Outcrop of diapir of Late Triassic salt

Fault trace (earlier thrust faults not shown)

Fluorspar deposit

Zinc-lead-(fluorspar-barite-copper-iron) deposit

Siderite iron deposit

EXPLANATION

Fluorspar Figure G5. 

0 10 20 30 40 MILES

0 10 20 30 40 KILOMETERS

N

ASTURIAS

AUTONOMOUS

COMMUNITY

SPAIN

FRANCE

Map area

Square symbols are the correct symbol for city locations

The location map does not need a source. 

I used the author’s png file as a template for the main map.

Location map does include a bit of France.

Figure G5. Map showing the locations of deposits in the Cantabrian salt-related carbonate-hosted mineral district in north-central 

Spain. The locations of the fluorspar mines and occurrences and the base map of the Asturias Autonomous Community on the west 

are modified from Garcia-Iglesías and Loredo (1994), and the locations of siderite iron deposits, zinc-lead-(fluorspar-barite-copper-iron) 

deposits, and outcrops of Late Triassic salt diapirs, as well as the base map of the eastern two-thirds of the area, are modified from 

Velasco and others (1994).
G20 Critical Mineral Resources of the United States—Fluorine

weight percent NaCl; that is, complex brines of probably 

sedimentary origin (González-Partida and others, 2003). On 

that basis, the Encantada-Buenavista deposits are labeled 

salt-related carbonate-hosted in figure G4. They are probably 

related to the salt-bearing Jurassic basin adjacent to the west. 

Kesler (1977), however, had earlier concluded that these and 

other deposits in the same Cretaceous limestone host rocks in 

Coahuila de Zaragoza, Mexico, were deposited from boiling, 

probably partly magmatic, fluids. Many of the deposits are at 

contacts with rhyolitic or trachytic high-level intrusive rocks 

(Temple and Grogan, 1963; Pickard, 1974; Kesler, 1977). As 

with Hansonburg, the Encantada-Buenavista district is likely 

to prove to have some magmatic component, thus validating 

its classification as both salt-related carbonate-hosted and 

related to peralkaline igneous rocks.

Fluorspar Deposits Related to Strongly 

Differentiated Granites

Fluorspar deposits are known to be present in association 

with strongly differentiated granites with a considerable 

range of compositions of the granite, from rather ordinary 

monzogranites, to two-mica granites, to tin granites. 

Generally, the greater the differentiation, the more common 

the occurrence of associated fluorspar deposits. Although 

the general behavior during differentiation is similar to that 

in alkaline granites, the associated metals and deposit types 

are distinct from the metals and deposit types associated 

with alkaline intrusive rocks. Deposits with the subalkaline 

granites include fluorite-rich tin and tungsten skarns, fluoriterich veins with tin and (or) tungsten, some with antimony, 

and fluorite-rich greisen deposits, and, fewer in number, 

disseminated or stockwork fluorspar deposits. Molybdenum 

and bismuth are additional elements that are commonly 

anomalous in these deposits. Greisens are rocks composed 

principally of muscovite (or lithium-rich white mica) and 

quartz with lesser abundances of one or more of the following: 

bertrandite, beryl, cassiterite, columbite-tantalite, fluorite, 

phenakite, topaz, tourmaline, and wolframite. Minor sulfides 

are found in greisens, principally pyrite, but locally others 

as well, including molybdenite. Although greisens appear to 

form by hydrothermal alteration of intrusive igneous rocks, 

a distinguishing feature is that feldspars are absent and have 

presumably been altered to the white mica.

Among the end-member granites that are included in 

the broad class of strongly differentiated granites, two-mica 

granites are those with both muscovite and biotite, which, 

given those two aluminous micas and the presence of both 

plagioclase and alkali feldspars, are invariably strongly 

peraluminous. (In peraluminous igneous rocks, molar 

Al2O3> (Na2O + K2O).) Tin granite is also a very specific 

type of granite (Tischendorf, 1978; Hudson and Arth, 1983). 

Tin granite can be porphyritic, seriate, or equigranular in 

texture, but all are light-colored peraluminous biotite granites, 

and all are small offshoots of large granodiorite-to-granite 

batholiths from which the tin granite evolved by fractional 

crystallization. The tin granites are found in stable cratonal 

regions and not in continental or oceanic arcs (Hudson and 

Arth, 1983). Fluorspar deposits associated with tin granites 

have a variety of geometries (veins; replacements; skarns at 

intrusive contacts; greisens in small, upward projections on the 

granite body’s roof [that is, endogreisens in granitic cupolas]; 

and greisens in wall rocks of the granite [that is, exogreisens]). 

Lithium-rich annite is an important characteristic mineral in 

tin granites. It was earlier termed “zinnwaldite.” Annite is the 

iron-rich end member of the biotite solid solution—that is, 

monoclinic K2

Fe6 (Si6Al2O20 )(OH)4. In lithium-rich annites, 

the lithium substitutes for aluminum and iron in this mineral 

and produces micas of the zinnwaldite series.

To date, fluorspar deposits related to strongly differentiated granites have been identified on every continent except 

South America and Antarctica (fig. G4 and appendix G1). 

They probably also occur on those two continents as well 

but simply have not yet been discovered. Tin-granite end 

members have not been found in the Andean magmatic 

belts, however, because those belts appear to be composed 

entirely of intrusive rocks of oceanic and continental arcs 

(Steve Ludington, U.S. Geological Survey, retired, written 

commun., December 6, 2013).

At places, assignment of a deposit to the class associated 

with strongly differentiated granites is very obvious, but at 

others, it is not as clear. For example, the very large Vergenoeg 

deposit in South Africa is near to but not in contact with 

granites that host tin deposits. The granites are part of the 

Bushveld Complex and are related to the platinoid-metal-, 

chromium-, vanadium-, and copper-nickel-bearing rocks of 

the layered mafic to ultramafic rocks of the lower part of the 

Bushveld Complex (Hatton and Sharpe, 1989). The Vergenoeg 

deposit is an approximately carrot-shaped, root-downward 

mass of very iron-rich igneous (or metasomatic?—that is, 

formed by replacement from hydrothermal fluids) rocks that 

cuts intermediate to felsic Rooiberg volcanic rocks. Those 

Rooiberg volcanic rocks are thought to have been derived by 

partial melting of upper crustal rocks that lay above the intrusion of the enormous, lobate, sill-like body of the Bushveld 

Complex and are considered to be the eruptive products of 

the magmas that became Bushveld granites and granophyres 

(Kleemann and Twist, 1989; Schweitzer and others, 1997). 

Vergenoeg ore consists of very coarse-grained (pegmatitictextured) fayalite-magnetite-siderite-fluorite rock that contains 

an average of 28 percent CaF2 and less than 30 percent SiO2. 

Flat-lying deposits of fluorite-hematite are located nearby in 

three directions from the pipe; these deposits are interpreted 

by Crocker (1985) as tuffs of unusual composition that erupted 

from the pipe. The average REE pattern of Vergenoeg fluorite 

precisely parallels that of average Bushveld granite but the 

fluorite has slightly less than one-tenth the concentrations 

found in the granites (Goff and others (2004). The origin of 

the Vergenoeg deposit remains uncertain, but its classification 

as related to strongly differentiated granite is indicated by the 

parallel REE patterns.
Geology G21

Definitely within the class of deposits associated with 

strongly differentiated granites are the Voznesenka and the 

Pogranichny deposits in Russia’s Far East Region (Ryazantseva, 1998; Sato and others, 2003) and the Lost River prospect 

in Nome Borough, Alaska (Dobson, 1982). These are fluorite 

deposits in which tin granites and related greisens are exposed 

literally within the map limits of the fluorspar deposits, 

although the tin granites themselves are not fluorspar ore. 

These deposits contain related cassiterite-bearing tin skarn that 

is very distinctive because the skarn (known as “wrigglite”) is 

rich in magnetite and vesuvianite and has a tortuously sinuous 

mineralized structure formed by layers that are a millimeter or 

less in thickness (Kwak and Askins, 1981). Cassiterite, which is 

the major ore mineral of tin, is tetragonal SnO2 and vesuvianite, 

which is the calc-silicate mineral of the skarn, is tetragonal Ca19

Fe(Mg,Al)8Al4(SiO4)10(OH)10. Magnetite is isometric Fe3O4. 

The fluorspar ores, which are replacements of former carbonate 

sedimentary rocks by minerals deposited from hydrothermal 

solutions, contain micas that are very lithium- and fluorinerich, so the ores may also be termed fluorspar greisens. Most 

of the same features are found at Moina in Tasmania, Australia 

(Kwak and Askins, 1981) and, apparently, at Bamianshan in 

Zhejiang Province, China (Han and others, 2012).

The El Hammam deposit in the Meknès-Tafilalet 

Region of Morocco; the Yinkuangchong deposit in Hengyang 

Prefecture, Hunan Province, China; and the Taskaynar 

deposit in Kazakhstan have spatial association to granitic 

rocks, but the description of the granites is not detailed 

enough to determine if they are strongly differentiated, much 

less whether they are tin granites or two-mica granites. The 

fluorspar ores at these places contain tin, tungsten, bismuth, 

or molybdenum minerals, and that association of elements 

leads to the deposits’ tentative classification as related to 

strongly differentiated granites. The veins at El Hammam 

contain minor amounts of stannite (tetragonal Cu2FeSnS4 ) and 

bismuthinite (orthorhombic Bi2S3; Cheilletz and others, 2010). 

At Yinkuangchong, two types of mineralized bodies are recognized. The lesser type consists of quartz veins that contain 

some amount of cassiterite. The larger polymetallic-fluorspar 

bodies are silicified breccias that contain minor amounts of 

scheelite (tetragonal CaWO4) (Chen and others, 1996). The 

carbonate replacement body at Taskaynar contains cassiterite 

in a pre-fluorite stage of the paragenetic sequence, and there is 

a halo of rocks that are high in molybdenum surrounding the 

deposit (Koplus and others, 1977).

Subalkaline-Volcanic-Related Epithermal 

Fluorspar Deposits

Many fluorspar deposits worldwide occur in association 

with subalkaline volcanic rocks. Subalkaline volcanic rocks, 

in contrast to alkaline and peralkaline volcanic rocks, do not 

have an excess of potassium and sodium over what is needed 

for feldspars. They commonly contain both plagioclase 

and an alkali feldspar. They are typically saturated with 

silica, and thus quartz crystallized as a late mineral from the 

differentiating melt. Subalkaline volcanic rocks are the most 

common rocks of continental arcs, and andesites are the most 

common rocks among the subalkaline volcanic rocks.

Fluorspar deposits associated with subalkaline volcanic 

rocks are the most plentiful in China, and many are found 

in Zhejiang Province along the county’s southeastern coast 

(fig. G4). Chen and others (1996) and Zhang and others 

(1997) describe the geology of fluorspar deposits in Zhejiang 

Province. Most of the deposits are described as simple large 

veins that have no close spatial association and no obvious 

genetic association with plutonic or shallowly intrusive 

igneous rocks. Some veins of the Yangjia-Wuyi district are 

continuous for more than 15 km along strike with thicknesses 

of 2.3 to 5.8 m. The ore is not continuous along strike, but ore 

shoots are typically separated by just 15- to 46-m intervals of 

quartz-rich and fluorite-poor parts of the veins. The YangjiaWuyi veins cut a section of Jurassic-aged intermediate to 

dominantly felsic volcanic rocks that are intercalated with 

sedimentary rocks. The section is dominated by rhyolitic 

tuff but is interlayered with limestone, quartzite, tuffaceous 

siltstone, and, locally, with shale and mudstone (Chen and 

others, 1996; Zhang and others, 1997). The veins include 

gangue minerals of quartz, chalcedony, opal, calcite, and 

barite, and small amounts of pyrite (isometric FeS2), apatite 

(hexagonal Ca5(PO4)3(F,OH,Cl)), and kaolinite (triclinic 

Al2Si2O4(OH)2). Wall rocks are altered with the addition of 

quartz, a lesser amount of kaolinite, and local pyrophyllite 

(monoclinic AlSi2O5(OH)2), carbonates, chlorite (monoclinic 

(Mg,Al,Fe,Li,Mn,Ni)4–6(Si,Al,B,Fe)4O10(OH,O)8), and pyrite. 

The ore and gangue mineralogy as well as the fluid temperatures and salinities are consistent with epithermal conditions 

(Chen and others, 1996). Isotopic studies are insufficient to 

address the origins of the fluids, but meteoric water was probably dominant. The fluorspar deposits in Zhejiang Province 

have been dated at 70 to 90 Ma by a samarium-neodymium 

isochron on fluorite and calcite (Li and others, 1998). Studies 

of neodymium and strontium isotopes are interpreted to indicate that most of the neodymium and strontium was derived 

from the Precambrian metamorphic rocks lying below the host 

rock, and that the remainder was derived from the Mesozoic 

volcanic and sedimentary wall rocks themselves (Li and 

others, 1998). Although the deposits contain a record of 

multiple igneous events dating from 135 through 97 Ma with 

various chemical signatures, the intrusive rocks from 97 Ma 

are those that, in time, could have contributed in small proportions to the mineralizing fluids. Those rocks, which belong to 

a high-potassium subalkaline suite of volcanic rocks (Lapierre 

and others, 1997), are peraluminous granites of an apparently 

ordinary continental arc. In these epithermal systems, the role 

of the high-potassium subalkaline volcanic rocks as sources 

for fluorine may have been more important than that of the 

granites, which may have merely been heat sources needed 

to cause groundwater convection. In other countries, also, the 

deposits associated with subalkaline volcanic rocks tend to be 

simple epithermal veins with crystals that grow symmetrically 

inward from both vein walls.
G22 Critical Mineral Resources of the United States—Fluorine

Other deposits that may tentatively be included in 

the class of epithermal fluorite deposits associated with 

subalkaline volcanic rocks are the Las Cuevas deposit in 

the State of San Luis Potosi, Mexico, and the nearby mines 

at El Realito and El Refugio in Guanajuato State, Mexico. 

Each of these three deposits is at the faulted and brecciated 

contact of limestone and rhyolite breccia (Ruiz and others, 

1980). The rhyolites are part of a province of high-potassium 

subalkaline rocks (Ruiz, 1983) that lies east of the main 

subalkaline Sierra Madre Occidental volcanic complex. 

Although the rhyolite lying along one wall of the Las Cuevas 

orebody is relatively fluorine-rich, and although Ruiz (1983) 

showed that the composition of the Las Cuevas deposit fluorite 

reflects the composition of its local wall rocks (including that 

rhyolite), the Las Cuevas rhyolite is not one of the “topaz 

rhyolites” that were erupted at approximately the same time 

as the Las Cuevas rhyolite, that have the highest fluorine of 

any volcanic rocks in the region, and that have associated tin 

mineralization in the Mexican tin belt (Huspeni and others, 

1984). Clark and others (1982) included these deposits 

together with those in a belt to the north, including Aguachile 

in Pico Etereo district, and El Tule district in Coahuila de 

Zaragoza, Mexico, which are associated with peralkaline rhyolites, as well as deposits of the Eagle Mountains in Hudspeth 

County, Tex., which are associated with a quartz syenite 

intrusion. That inclusion appears questionable based on the 

contrast in chemistry of the volcanic rocks near Las Cuevas, 

El Realito, and El Refugio compared with the chemistry of 

volcanic rocks associated with the deposits in Coahuila. The 

rhyolites at those three deposits are part of a high-potassium 

subalkaline province; they are not peralkaline like the rhyolites 

in Coahuila.

The Sumochagan Obo deposit in China’s Inner Mongolia 

Autonomous Region may also belong in the class of epithermal fluorspar deposits associated with subalkaline volcanic 

rocks. It is a large, rich deposit approximately conformable 

with the layering of calcareous tuffs and tuffaceous limestones 

(Nie and others, 2008) within a high-potassium subalkaline 

volcanic and sedimentary pile (Xu, 2009) of Early Permian 

age (appendix G1). At its closest point, the deposit is about 

1,200 m distant from a peraluminous granite intrusion (Xu 

and others, 2008) into that volcanic pile. It thus resembles 

the other deposits listed in this section in its tectonic and 

magmatic setting; however, its tuffaceous limy host rock and 

its approximately conformable geometry would perhaps allow 

it to be alternatively placed in the class of apparently conformable deposits within tuffaceous limy lacustrine sediments.

Fluorspar Deposits Apparently Conformable 

Within Tuffaceous Limy Lacustrine Sediments

A number of deposits in the Latium Region of westcentral Italy make up the class of apparently conformable 

deposits within tuffaceous limy lacustrine sediments, together 

with a single fluorite occurrence in southeastern Oregon, 

and, as noted above, possibly the Sumochagan Obo deposit 

in China. The deposits in Latium, Italy, are fine grained with 

more than 85 weight percent of the fluorite less than 40 μm 

(that is, silt and clay size), and, in fact, the proportion of 

fluorite in the Latium Region resource increases as the grain 

size decreases (Mastrangelo, 1976, p. 31). In an apparently 

conformable occurrence near Rome, Oregon, the fluorite grain 

size is predominantly less than 1 μm (Sheppard and Gude, 

1969). The occurrence at Rome, Oreg., is in Pliocene rock 

(1.8 to 5.3 Ma). If Sumochagan Obo belongs within this class, 

then it must have been considerably modified from its original 

form because fluorite in the deposit has grain sizes of from 

110 to 6,370 μm; that is, it has the grain sizes of very fine sand 

to fine conglomerate using standard grain sizes for sand. The 

texture of the “primary industrial ore” of Sumochagan Obo is 

“sugar-granular” (Yong and others, 2011, p. 14); thus, it ranges 

from coarse to very coarse sand. That the grain size of the 

Sumochagan Obo ore would have coarsened since the time of 

its deposition is reasonable. The deposit is in Early Permian 

age rocks (about 270 Ma), whereas the Italian deposits are in 

Pleistocene to Pliocene age rocks (0.9 to 3.7 Ma).

The deposits of the Latium Region appear to be conformable with layering in a mixed volcanic and sedimentary 

section, and they appear to constitute a hydrothermal sediment 

interlayered within lake-deposited tuffaceous limy mudstone 

and siltstone. The volcanic rocks of this region are alkaline, 

varying from trachytic tuffs to leucitites to phonolites to 

tephrites (the series of names of alkalic volcanic rocks just 

listed is from richest to poorest in alkali feldspar; all are 

undersaturated with silica). Most of the volcanic rocks contain 

leucite, analcime (isometric or monoclinic Na(AlSi2 )O6 • H2O), 

or nepheline. Flat-lying volcanic rocks dominate the section, 

but lake-deposited and lesser amounts of stream-deposited 

sedimentary rocks occur in multiple places in the section. The 

Latium Region is pockmarked by maar lakes (that is, lakes 

formed in shallow volcanic craters), which occur in the same 

40-by-110-km zone as the fluorspar deposits. Maar lakes are 

nearly circular, have no volcanic cone, and have very steep 

walls. They are produced by single volcanic explosions caused 

by magma intruding into groundwater-rich rocks, which 

changes the groundwater explosively to steam, which then 

throws the overlying rocks upward and outward. Although a 

number of Latium Region fluorspar deposits were mined in 

the past, most notably the Castel Giuliano-Pianciano deposit, 

little of the fluorite was recovered, probably because of the 

very fine grain sizes of the minerals, and no deposit there 

is currently being mined despite the presence of identified 

resources.

The Latium Region deposits consist of various proportions of fluorite, barite, apatite, calcite, and quartz (or 

chalcedony), and the overall grades are from 37 to 83 weight 

percent CaF2

. The content of barite varies inversely with that 

of fluorite. The deposits are from 1 m thick to as much as 15 m 

thick over areas of up to tens of square kilometers. The host 
Geology G23

lacustrine rocks include beds of diatomite, but diatoms (fossils 

consisting of silica that are each the size of a sand grain or 

less) are not typically found in the fluorspar resource. From a 

1976 symposium about the deposits, the majority opinion 

was that the fluorspar deposits had formed from fumarolic or 

hydrothermal solutions that discharged onto the floors of maar 

lakes (Matteucci, 1976). A strong support of that hypothesis 

was that, at places, a tawny, plastic earthy facies of fluoriterich rock is slump-folded but overlain by an undeformed 

white-to-yellowish gray gritty facies of fluorite-rich rock.

The occurrence near Rome, Oreg., underlies an area of 

about 18 square kilometers (km2 ) with an average thickness of 

about 2.3 m and an average grade of about 2.65 weight percent 

CaF2. Using semiquantitative X-ray diffraction, Sheppard 

and Gude (1969) found a regular set of changes in fluorite 

content and in other authigenic minerals through the section of 

basaltic tuffs and lacustrine tuffaceous mudstone. An overlying 

yellow tuff section appears to have lost fluorine, perhaps 

during a weathering period that produced alkaline, sodium 

bicarbonate (NaHCO3)-type soil waters. The climate and local 

rock compositions would probably have led to developing the 

same kind of NaHCO3-rich water of high pH as was found in 

the Eastern Rift Valley of eastern Africa that had anomalously 

high concentrations of fluorine (table G2) (Kilham and Hecky, 

1973), although probably at considerably lesser concentrations in Oregon. Probably the once glassy components of 

that yellow tuff altered to the zeolite minerals clinoptilolite 

(monoclinic (Na,K)6(Al6Si30))072 • 20H2O), phillipsite (monoclinic K(Ca0.5,Na)2

(Al3Si5)O16 • 6H2O), and erionite (hexagonal 

K2NaCa1.5Mg(Al8Si28)O72•28H2O). Beneath it, the fluoritebearing tuffaceous mudstone became enriched in calcite, 

fluorite, quartz (or chalcedony), and mordenite, a potassiumrich zeolite (orthorhombic K2.8Na1.5Ca2

(Al9Si39)O96 • 29H2O). 

Although some parts of their geologic settings are shared, there 

is no observed or predicted evidence that the same mechanism 

might have created the conformable fluorite deposits of the 

Latium Region, Italy, and the occurrence at Rome, Oreg. 

At just 2.65 percent CaF2, the Rome, Oreg., occurrence is 

unlikely ever to be considered economic for mining.

Residual Fluorspar Deposits

Many fluorspar deposits of the world were first discovered in residuum, and many constituted ore at the surface, 

particularly where costs allowed the ore to be hand sorted. 

Residual fluorspar deposits were well known in the IllinoisKentucky fluorspar district in the United States, where the 

carbonate gangue of the veins, predominantly calcite, and the 

limestone wall rocks were dissolved away during weathering 

(Weller and others, 1952, p. 29). Residual ore extended to 

depths locally as great as 30 m. A large volume of the ore in 

the Marico district in North West Province, South Africa, is 

residual, where it is termed “kokkoman.” Large bodies of 

kokkoman can exceed 50 percent CaF2 (Martini, 1986; Ryan, 

1986). The ore is residually enriched by the dissolving away 

of dolomite from the host rocks during weathering. Residual 

fluorite deposits were also widespread in the northern area of 

fluorspar production in Thailand near the Ban Hong deposit 

(fig. G4; Fulton and Montgomery, 1983).

Fluorine Recovery From Brines

Fluorine products are not currently recovered from 

brines anywhere in the world. As pointed out by Shawe 

(1976, p. 22), however, production of a fluorine product from 

a brine is a distinct possibility, perhaps in the near future. 

A number of other mineral commodities are produced from 

brines, for example, lithium (Kunasz, 1974; Ericksen and 

others, 1978). If a fluorine product is produced from a brine, 

it would not be a brine similar to those that produce lithium, 

however. Fluorine-rich brines are very alkaline (pH commonly 

about 10), are dominated by sodium (Na) and bicarbonate 

(HCO3

−), and are brines from which trona is precipitating. The 

brine from Lake Magadi, Kenya (table G2), is an example of 

such a fluorine-rich brine from within the trona sediment that 

makes up the floor of the lake. This is the area from which 

villiaumite has been identified in the trona (Nielsen, 1999). 

The surface water of Lake Magadi averages about 1,360 ppm 

fluorine, and intrastratal brines from boreholes in bottom 

sediments average about 1,200 ppm (Jones and others, 1977). 

At present, there is no process by which fluorine can be 

recovered from the water even with such high concentrations; 

however, because fluorite has been observed in calcareous 

rocks where they are in contact with this high-fluorine water, 

and where the fluorite has replaced calcite (Surdam and 

Eugster, 1976), it is likely that addition of calcium by almost 

any means is a possible way to extract the fluorine from the 

water (for example, see Aldaco and others, 2007). The water 

could be run across a bed of ground limestone (the area 

has caliche developed in several levels of the Tertiary and 

Quaternary stratigraphy—Eugster, 1980); however, production 

of an aluminum-fluoride (AlF3) product might be preferable. 

Possibly, bauxite could be reacted with fluorine-rich brine to 

make an AlF3

 product directly.

Fluorspar Deposit Grades and Tonnages 

by Deposit Class

From the previous discussions, it could be expected that 

the deposit classes have different typical grades and tonnages. 

The only published previous attempt to examine fluorspar 

deposit grades and tonnages systematically was Orris (1992), 

who restricted her study to fluorspar veins. The advent of 

measures such as Canadian Government National Instrument 

(NI) 43–101 reports and the South African Code for Reporting 

Exploration Results (SAMREC)-compliant resource and 

reserve estimates has made information on the grades and 

tonnages of many more fluorspar deposits publicly available 

since the 1992 study.
G24 Critical Mineral Resources of the United States—Fluorine

Figure G6 is a plot of fluorspar deposit grades and 

tonnages made from the data in appendix G1. The plotted 

grades and tonnages have not been exhaustively vetted to 

assure their validity. A number of pitfalls are possible in the 

reporting of deposit grades and tonnages by private interests. 

The data in appendix G1 have come from many different 

sources, so the resulting plot presents data collected using a 

wide variety of reporting criteria. If the majority of the data 

used are valid, then the overall trends shown on the plot are 

probably real.

In figure G6, deposits with vein geometry have been 

treated as a single category regardless of their class in 

figure G3B, but deposits without vein- or dike-like geometry 

(that is deposits associated with carbonatites, and deposits 

associated with strongly differentiated granites) are each 

treated separately. The plot in figure G6 probably includes a 

few small occurrences in which only the resources considered to be sufficient to assure the next planned advance of 

mining were reported, although in cases where this could be 

confirmed, the deposit was not plotted. A number of other 

occurrences that were mined, but then closed, or that were 

mined intermittently only when prices were high have been 

included. Examples are the Huckleberry (Carlton County), the 

White Eagle (Sierra and Grant Counties), and the Great Eagle 

(Grant County) occurrences in New Mexico; the Canxixe 

(Sofala Province), the Mount Muambe (Tete Province), and 

the Macossa (Manica Province) occurrences in Mozambique; 

and the Jebel Sumayh occurrence in North Kordofan State, 

Sudan. Consequently, not all the plotted deposit grades and 

tonnages are for reserves (resources that were profitable to 

mine). The reasons why unprofitable resources were included 

is explained below.

Although individual classes on the plot appear to show 

very little tendency for tonnage and grade to systematically 

covary or to vary inversely, the three deposit classes together 

appear to show that, of reported deposits, high-grade fluorspar 

deposits tend to be smaller in tonnage, and lower grade 

deposits tend to be larger in tonnage. That overall inverse 

relation is common to deposits of many different kinds of 

mineral commodities. The lack of reporting for low-grade, 

low-tonnage deposits also is a contributing factor to the 

inverse relation observed with the reported data. The plot 

shows that economic fluorspar veins are generally smaller 

and higher grade than economic carbonatite-related fluorspar 

deposits which are, in turn, generally smaller and of higher 

grade than economic fluorspar deposits associated with 

strongly differentiated granite. These results may be overly 

simplistic, however, because some veins have barite, lead, 

and zinc byproducts or coproducts; some carbonatite-related 

deposits have niobium and barite byproducts; and some 

granite-related deposits have tin, tungsten, and possibly 

antimony, bismuth, and (or) molybdenum byproducts 

or coproducts.

On the plot in figure G6, the Zwartkloof deposit in 

the Warmbaths district of Limpopo Province, South Africa, 

which is a deposit related to strongly differentiated granite, 

is labeled Z, and the Argentolle deposit, Bourgogne Region, 

France, which is a vein deposit probably of the salt-related 

carbonate-hosted type, is labeled A. The Zwartkloof Mine 

operated only from 1971 to 1973 and was closed when the 

operating company concluded that the grade of the deposit 

(13.7 percent CaF2

) was insufficient (Brian Hodge, consultant 

on world fluorspar deposits, written commun., April 29, 2013). 

The deposit at Argentolle was mined from 1971 to 1982, and 

was also abandoned as uneconomic. Zwartkloof and Argentolle are important because an imaginary line connecting the 

two on the grade vs. tonnage plot (fig. G6) divides the plot into 

two parts, one to the right and up from the line; the other to the 

left and down from the line. That line approximately marks a 

boundary between deposits that, as of late 2015, were viably 

economic (right and up) and those that were uneconomic (left 

and down, including Zwartkloof and Argentolle, themselves).

The approximate boundary formed by the line between 

A and Z should not be considered a permanent and rigid 

divider between economically favorable and unfavorable 

grades and tonnages of fluorspar deposits. For example, 

that line certainly expresses some aspects of mining history, 

showing, on that boundary, at least two deposits that were 

Carbonatite-related fluorspar deposits

Fluorspar deposits related to strongly

differentiated granites

Vein fluorspar deposits of all classes

Fluorspar mine that closed 

EXPLANATION

Fluorspar Figure G6

A

A

Z

Z

Argentolle

Zwartkloof

7

10

50

100

0.01 0.1 1 10 100 200 

Deposit tonnage, in million metric tons 

Grade, in weight percent CaF2

Figure G6. Plot of fluorite (CaF2) grade versus tonnage for 

fluorspar deposits related to strongly differentiated granites, 

for carbonatite-related fluorspar deposits, and for veins from all 

classes of fluorspar deposits. Two vein fluorspar mines that failed 

because of either insufficient grade (Zwartkloof; Z) or insufficient 

tonnage (Argentolle; A) are identified on the plot. Data are from 

table G1–1; deposits whose grade and tonnage are plotted are 

listed in the footnote.
Resources and Production G25

historically considered minable but that have become 

uneconomic in the modern fluorspar world market. Also, a 

deposit that is being explored may likely pass from lower 

left to upper right—from uneconomic to economic—by 

crossing the line horizontally (that is, becoming larger with the 

knowledge gained by additional drilling), although a deposit 

might cross the line vertically (that is, become higher grade as 

high-grade drilling results are achieved once the best part of 

the deposit is reached through exploration).

In any case, the plot provides a measure with which to 

compare the resources of new deposits or deposits still in 

exploration with typical fluorspar deposit grades and tonnages; 

the farther to the right and up, the better. The deposits in 

Las Cuevas, Mexico, and Vergenoeg, South Africa, are the two 

deposits that are farthest up and to the right of that imaginary 

line between A and Z.

Resources and Production

Identified Resources in the United States

The United States has numerous moderate-tonnage (that 

is, tens of thousands of metric tons to up to a few hundred 

thousand metric tons) but high-grade fluorspar deposits scattered across seven States (Idaho, Illinois, Kentucky, Montana, 

Nevada, New Mexico, and Utah). Some larger tonnage but 

lower grade deposits have also been identified in Alaska, 

Illinois, Nevada, and Tennessee.

Some moderate-tonnage high-grade fluorspar deposits 

and one large-tonnage low-grade deposit have been identified 

in the Illinois-Kentucky fluorspar district. The Anabelle Lee, 

the Denton, and the Minerva Number 1 Mines (Hardin 

County, Ill.) in the Illinois-Kentucky fluorspar district 

closed in 1996 with ore still in the faces of each mine and 

some limited tonnages of drilled resources identified at each 

(Richard Yancy, former Ozark-Mahoning Co. mine geologist 

at the three mines, oral commun., March 11, 2013). Since that 

time, exploration and development in the Illinois-Kentucky 

fluorspar district has taken place at the Klondike II project, 

which is located about 8 km southwest of Salem, Ky., in 

Livingston County. Production began in 2012 (W.H. Anderson, 

Kentucky Geological Survey, oral commun., October 22, 2014). 

The Klondike II deposit is located on the Pittsburgh Fault and 

constitutes a vein resource of at least 1.6 million metric tons 

of ore grading 60 percent CaF2 (Feytis, 2009). A decline has 

been driven from the surface to allow rubber-tire mechanized 

underground mining. Plans are in place to ramp up production 

to 200,000 metric tons per year during a period of several 

years. Other exploration is ongoing nearby and to the northeast 

in southern Crittenden County. Farther north, Hicks Dome 

Corp. (HDC) has continued to hold and explore the company 

property since obtaining its first leases in 1970.

Since the HDC property was first leased, several different 

partners have worked together in joint ventures with HDC, 

although each left the partnership within a few years. The last 

public announcement of tonnage and grade at Hicks Dome 

was reported by Kirkemo (1978) and totaled just 11.3 Mt at 

a grade of 11.5 percent CaF2. This estimate was made when 

only 12 exploration core holes had been completed, however. 

Since that time, tens of additional holes have been drilled, 

many of these encountering similar mineralization. Still, 

this is a relatively low-grade fluorspar deposit, although its 

tonnage as now known is doubtlessly significantly larger than 

that previous estimate of 11.3 Mt. With the known geology 

of Hicks Dome, production of fluorspar as a byproduct or 

coproduct of REE mining could possibly take place if the 

property comes into production (Kirkemo, 1978). Fluorite 

mineralization at Hicks Dome is found in breccias that also 

contain substantial concentrations of beryllium, niobium, 

REEs, thorium, and yttrium, as well as local concentrations 

of zinc, lead, and barite.

Lost River, Alaska, mentioned earlier, is the largest 

identified resource of fluorspar in the United States outside of 

the Illinois-Kentucky fluorspar district. Tin and fluorspar could 

possibly be produced as coproducts, along with tungsten and 

possibly beryllium as byproducts if a mine can operate profitably in that remote location and extreme climate. Additional 

resources in the Western United States have been identified 

in southwestern Utah, along the flanks of the Rio Grande Rift 

in New Mexico and Texas, in two small districts in Idaho, in 

one place in Montana, and in at least two places in Nevada. 

Among identified resources in the Western United States, the 

McCullough Butte deposit, which is located 19 km southwest 

of Eureka, Nevada, appeared to be in an advanced stage of 

exploration at the end of 2013. Several exploration programs 

had been conducted at the property in the past; the latest one 

was by Tertiary Minerals plc, which also owns the Storuman 

deposit in Lappland Province, Sweden, and the Lassedalen 

fluorspar deposit in Buskerud County, Norway. McCullough 

Butte is a skarn deposit in limestone along its contact with a 

quartzite adjacent to and extending away from dikes of Late 

Cretaceous two-mica granite porphyry (Barton, 1982). Tertiary 

Minerals has announced a Joint Ore Reserves Committeecompliant estimate of 6.1 million metric tons of indicated 

resources at a grade of 10.8 percent CaF2

, and 80.3 million 

metric tons of inferred resources at a grade of 10.7 percent 

CaF2 at a cutoff grade of 8 percent CaF2 (Tertiary Minerals plc, 

2015). The geology and topography of the McCullough Butte 

deposit would allow for open pit mining.

In the Eastern United States, low-grade resources 

have been identified in the Sweetwater (MVT) district in 

Tennessee. With the exception of the deposits in the IllinoisKentucky fluorspar district described above and possibly 

the McCullough Butte deposit, none of the resources in the 

United States would appear to be economic at this time, given 

competition with Mexican and South African imported HF 

and fluorspar.

The success of the project at the recently discovered 

Klondike II deposit in Kentucky demonstrates that there 

probably are still additional deposits to discover in old 

districts. The Klondike II deposit has been described as 
G26 Critical Mineral Resources of the United States—Fluorine

being 18 m thick, 91 m vertically from the surface to the 

bottom of the ore, and about 1,200 m along the strike of the 

vein, and as containing 3.6 million metric tons of proven 

ore (reported as 60 feet wide, 300 feet deep, more than 

4,000 feet long, and containing about 4 million short tons) 

(Anderson and Moodie, 2010). That an orebody of such size 

would have escaped discovery for some 140 years since 

the district was recognized as a major source of fluorspar is 

perhaps surprising, but it is not remarkable. The district has 

several tens of known veins and areas of bedding replacement ores. The aggregated total strike length of faults with 

known fluorspar mineralization along them is conservatively 

estimated to be 350 km. A scientist from the Kentucky 

Geological Survey (W.H. Anderson, Kentucky Geological 

Survey, oral commun., January 24, 2014) who has worked in 

the Illinois-Kentucky fluorspar district now and again since 

1989 says that exploration along those faults is generally 

thorough enough to preclude additional orebodies, but points 

to the following two areas that are thought to have remaining 

potential: (1) the area of the Rock Creek graben southwest of 

the Ohio River in Kentucky for new discoveries of bedding 

replacement deposits, and (2) the extension of the Tabb faults 

southeastward beyond the Senator, Eddie Crowder, and 

Williamson Mines that form the current-day southeastern limit 

of the district for new discoveries of vein ores.

Identified Resources in Other Countries

The Nui Phao fluorspar-tungsten-bismuth-copper-gold 

deposit in Thai Nguyen Province, Vietnam, about 80 km 

north-northwest of Hanoi, began operations at the end of 

April 2013 after an extended period of permitting, land 

acquisition, and other preparations. The deposit is being 

mined by open pit, and it is projected to provide nearly 

214,000 metric tons per year of CaF2 for 20 years, as well as 

about 5,620 metric tons per year of tungsten, 2,675 metric 

tons per year of bismuth, and 588,000 grams per year of gold. 

Paragenetically, early mineralization at Nui Phao is a tungsten 

skarn thought to be related to the intrusion of a Triassic biotite 

granite. Later hydrothermal overprint thought to be related to 

the intrusion of a Cretaceous two-mica granite is of pyrrhotitefluorite-albite greisen (Richards and others, 2003; Masan 

Group, 2013).

In South Africa, the Nokeng fluorspar project, which 

is located on a property that borders the Vergenoeg Mine 

in northern Gauteng Province, was scheduled to start mine 

construction in late 2012, with production targeted for the 

second quarter of 2014. As of February 2014, construction 

had apparently not yet started, and no further information that 

explains the reason for the delay was available. The project 

planned to mine two orebodies—the Plattekop deposit and 

the Outwash Fan deposit. Together, the deposits contain 

SAMREC-compliant reserves of about 12.2 million metric 

tons of 27.2 percent CaF2

 (Clay and de Wit, 2009). The 

Plattekop deposit is a conformable layer within the local 

Vergenoeg Suite stratigraphy that overlies the Vergenoeg 

fluorite-magnetite-fayalite-sulfide breccia pipe. The Plattekop 

deposit is a 6- to 27-m-thick layer of hematite-fluorite within 

a sequence of ferruginous agglomerates, tuffs, felsites, and 

ironstones. Its outcrop, which forms a hilltop, extends from a 

sharp contact with the pipe southeastward for about 1,200 m. 

The Outwash Fan deposit occurs 7.6 km south-southeast from 

Plattekop Hill and extends for another 3.5 km from there. 

The Outwash Fan ore is in a 1- to 9-m-thick layer of fluoritebearing conglomerate or sedimentary breccia intercalated 

among laminated fine grits, quartzites, bedded ironstones, 

cherts, and conglomerates. The orebodies are planned to be 

mined by separate open pits. Fluorspar is to be recovered by 

flotation to make acid-grade concentrates, and a production 

rate of 180,000 metric tons per year of CaF2 is expected 

(Clay and de Wit, 2009).

Also in South Africa, the Doornhoek deposit in 

North West Province is simply the eastward extension of 

the Witkop MVT dolomite-hosted fluorspar deposit in the 

Marico district. The Doornhoek owners “planned to initiate 

underground development, feasibility studies and design” 

(Roberts, 2012, p. 38.). No additional information from the 

company was available.

The Speewah fluorspar deposit in Western Australia, 

Australia (which contains 6.7 million metric tons grading 

24.6 percent CaF2 ), was purchased by a Chinese company, 

Jiangxi Yunfeng. The project was undergoing a preproduction 

feasibility study (Mining Atlas, 2016).

The Lassedalen fluorspar mine (which contains 4 million 

metric tons grading 24.6 percent CaF2 ) in Buskerud County, 

southwestern Norway, was scheduled by Tertiary Minerals 

to be reopened in 2012, but the latest available company 

releases suggest that prefeasibility study drilling would 

be conducted with the purpose of completing a Joint Ore 

Reserves Committee (JORC)-compliant estimate of reserves 

(Tertiary Minerals plc, 2014). The deposit is adjacent to one 

of the Permian peralkaline granites of the Oslo graben (Ineson 

and others, 1975). Tertiary Minerals’ Storuman deposit in 

Sweden was somewhat nearer to being ready to mine; the 

deposit is a replacement deposit along bedding in Cambrian 

sandstones. This has been called an MVT deposit, based 

on nearby sphalerite-galena-fluorite-calcite-bearing ores 

occurring in fractures within the Precambrian basement with 

strongly saline fluid inclusions that homogenize mostly from 

80 to 200 °C and also, in the fluorite, contain hydrocarbon 

inclusions (Billström and others, 2012). The class assignment 

is also based on analogy with the zinc-lead deposit at Laisval, 

located nearby, that formed in the same host rocks in the 

mid-Silurian Period, a time when thrust-belt compressional 

mountains were thrown up just tens of kilometers to the west 

(Sherlock and others, 2005). The JORC-compliant resource 

estimate for Storuman is 25.0 million metric tons of indicated 

resources grading 10.28 percent CaF2 and 2.7 million metric 

tons of inferred resources grading 9.57 percent CaF2. Tertiary 

Minerals planned to submit its mining lease application in the 

first quarter of 2014 and its environmental permit application 
Resources and Production G27

sometime in 2014. The company planned to begin 

development of the Storuman Mine in 2016 and to start 

production in 2017. The mine had a planned mine life of at 

least 18 years at a production rate of roughly 160,000 metric 

tons per year of CaF2. The deposit appears to be open in 

at least two directions, however, and additional drill holes 

indicate mineralization in both directions.

In China, a probably relatively large tungsten-tinbismuth-molybdenum-fluorite deposit was found in western 

Zhejiang Province within the past several years. It is called 

Bamianshan, and although there is considerable geologic 

literature on the deposit already, its exact location and grade 

and tonnage have apparently not been reported. The deposit 

is geologically interesting because it is clearly related to a 

strongly differentiated granite (Liu and others, 2012), and it 

is located close to several large epithermal deposits related 

to subalkaline volcanic rocks, including (in decreasing order 

of size) Yangjia-Wuyi, Houshu, Hushan, and Badu.

In Newfoundland, Canada, the St. Lawrence fluorspar 

project, which has had exploration ongoing for nearly a 

decade by Canada Fluorspar, Inc., appears to be becoming 

active again, after a period of pause and review. Arkema, 

Inc. of France became a joint-venture partner with Canada 

Fluorspar in May 2012, stimulating a review of the project’s 

costs that began in about September 2012. The project 

issued an updated preliminary feasibility study in 2013 

(Lecuyer and others, 2013). No new startup date had then 

been announced. The project planned to mine two large 

veins, the Blue Beach North and the Tarefare veins, which 

cut Late Devonian or Early Mississippian peralkaline granite 

(Van Alstine, 1944; Teng and Strong, 1976; Strong and 

others, 1984). This mining district was last active in 1978, 

but mining had been documented there for up to 45 years 

before that time, and some records indicate that it had 

been mined as early as 1870 by early French settlers. Total 

indicated and inferred resources on the Blue Beach North 

and Tarefare veins combined are 10.05 million metric tons 

grading 41 percent CaF2 (Lecuyer and others, 2013). Plans 

were in place to produce about 131,000 metric tons per year 

of acidspar.

In the United Kingdom, British Fluorspar was formed 

as a successor of Minmet UK, Ltd., and the company has 

reopened the Milldam Mine on Hucklow Edge Rake in the 

Southern Pennine Orefield (Hodge, 2012). British Fluorspar 

restarted mining in March 2013 and sent its first shipments 

in May 2013 (Lismore-Scott, 2013). Milldam Mine was part 

of the holdings of the former Glebe Mines, last a subsidiary 

of INEOS Flúor, and Glebe had been separated from an 

acquisition of INEOS Flúor’s fluorochemical business by 

Mexichem Flúor on April 1, 2010. That had led to an end, 

for a time, to fluorspar mining in the United Kingdom, 

as of the end of 2010, although Mexichem had continued 

to operate the fluorochemical production facilities of the 

former ICI/INEOS using acidspar imported from Mexico.

Summarizing the above, there are multiple new fluorspar mines being started up or reopened around the world, 

including a mine in China as well as at least one mine each 

in five other countries. Four of those five mines (excluding 

the Milldam Mine) planned cumulative production of about 

770,000 metric tons per year of fluorspar within a few years. 

Additional resources have been identified in Australia, South 

Africa, Tunisia, the United States, and probably a few other 

countries. Hodge (2012) was clearly correct in his conclusion 

that “. . . there should be no serious supply shortages providing 

a pragmatically sensible and ongoing rapport is established . . . 

between consumers and producers.”

Miller (2013) estimated the reserves of fluorspar of 

various countries for which information was available at 

that time. In units of 100 percent CaF2 and rounded to the 

nearest million metric tons, they are as follows, from most 

to least reserves among the top eight countries: South Africa, 

41 million metric tons; Mexico, 32 million metric tons; China, 

24 million metric tons; Mongolia, 22 million metric tons; 

Spain, 6 million metric tons; Namibia, 3 million metric tons; 

Kenya, 2 million metric tons; and Brazil, 1 million metric tons. 

Reserve data were not available for Kazakhstan, Morocco, 

Russia, and the United States. As discussed above, reserves 

from Vietnam and Canada and additional reserves from the 

United States can be expected to enter this list soon. Reserves 

in the remainder of the world were estimated to be 110 million 

metric tons totaling, together with the eight countries for 

which reserves were reported, about 240 million metric tons. 

The world’s total identified resources were estimated to be 

500 million metric tons of contained fluorspar (Miller, 2013).

Undiscovered Resources in the United States

A preliminary model for large, high-grade fluorspar 

deposits has been developed based on deposits in Mexico. 

Such deposits as Las Cuevas (San Luis Potosi State), Realito 

(Guanajuato State), and El Refugio (Guanajuato State) are 

localized along faulted and brecciated contacts between 

Cretaceous limestone and Tertiary rhyolite, east of the 

main volcanic arc that is slightly older than the rhyolite. 

The fluorspar-related rhyolites are post-convergent and 

extension-related and are not derived from the same arc (or 

any arc) (Steve Ludington, U.S. Geological Survey, retired, 

written commun., 2013). Deposits in Texas in a belt parallel 

to the Rio Grande from El Paso to Big Bend National Park 

share most if not all of these characteristics, but the deposits 

are too small to mine profitably (McAnulty, 1974). Modern 

exploration has not taken place in these areas. Elsewhere 

along the Rio Grande Rift are numerous fluorspar occurrences 

and deposits (Van Alstine, 1976), some of which are fairly 

large (Hansonburg district, New Mexico, for example). 

Modern exploration techniques have not been applied in 

those areas either.

Iron Mountain, N.M., is a fluorine-, beryllium-, and 

tungsten-bearing skarn related to a strongly differentiated granite 

(Jahns and Glass, 1944), and this type of mineralizing system 

has formed many large deposits around the world (fig. G4). 
G28 Critical Mineral Resources of the United States—Fluorine

The presence of additional strongly differentiated granites 

in southwestern New Mexico is predictable, perhaps, by the 

numerous occurrences of wood tin related to rhyolite flows 

and domes in that region (Maxwell and others, 1986). A oneto-one relation between rhyolites with wood tin and strongly 

differentiated granites has not been demonstrated, however.

The province of Late Cretaceous two-mica granites 

described by Barton and Trim (1991), including the granite 

porphyry related to the McCullough Butte deposit in Eureka 

County, Nev., could be examined closely to determine if 

others of those granites have related fluorite-bearing skarns. 

Likewise, the province of tin granites on the Seward Peninsula 

in Alaska (Hudson and Arth, 1983) might contain additional 

fluorspar skarn-and-greisen deposits similar to the Lost River 

deposits. Other areas of Alaska with tin greisen mineralization have been explored but with little attention paid to the 

associated fluorspar (Hudson and Reed, 1997).

The Basin and Range province of the United States in 

Nevada and parts of Arizona, California, Oregon, and Utah 

may contain lacustrine brines that have extractable fluorine. 

Taking Lake Magadi (Kenya) and Lake Natron (Tanzania) 

discussed earlier in this chapter as examples, the water in 

and beneath Alkali Lake in Lake County, Oreg., could be 

an analogous evaporitic brine from which fluorine might 

be recoverable. That playa lake has magadiite (monoclinic 

NaSi7O13(OH)3 • 3H2O) and trona precipitated as evaporites in 

the lake’s recent sediments (Rooney and others, 1969), which 

are the same two minerals that characterize the evaporite 

minerals of Lake Magadi and Lake Natron. Chemical data 

for water from several sites around the Alkali Lake basin 

from 1969 and 1970, however, show a maximum fluoride 

concentration of only 12.75 milligrams per liter (mg/L), 

which is about 100 times less than the largest concentrations 

at Lake Magadi (Newton and Baggs, 1971, p. 18–30). The 

highest fluorine value comes from a water sample with a 

pH of 10 and total dissolved solids of 74,850 mg/L, both 

also highest among the 13 water-sample analyses from the 

Alkali Lake area. Together, these water chemical values 

appear to indicate that some of the same evaporationdominated processes are operating at Alkali Lake as operate 

at Lake Magadi and Lake Natron. Other basin and range lakes 

with recent evaporitic precipitates of trona include Big Soda 

Lake located northwest of Fallon, Nev. (Vanderburg, 1940); 

Searles Lake, California (Smith, 1979); and Deep Springs 

Lake, Calif. (Jones, 1965). Water analyzed by Jones (1965, 

p. A30) from near Deep Springs Lake had concentrations up 

to 171 ppm fluorine. Although this sample was anomalous, 

the fluorine concentration is still roughly 10 times smaller 

than in Lake Magadi. Most of the water analyses reported in 

Smith (1979) at Searles Lake did not include fluorine among 

the analytes, but eight did. The highest fluorine concentration 

among those eight samples was 35 ppm, which is approximately 10 times the typical fluorine concentration in lowtemperature surface water or groundwater, but it is still far 

below a concentration that would be economically interesting 

for its fluorine content.

Undiscovered Resources in Other Countries

Conclusions from the analysis in the Identified Resources 

in Other Countries section are extensible to the subject of 

worldwide undiscovered fluorine resources; in short, the world 

likely has no shortage of deposits yet to be discovered. Many 

deposits probably remain to be discovered on every continent. 

Deposits of every type (fig. G3) almost certainly remain to 

be discovered, though some general additional guidelines for 

exploration can be helpful to supplement deposit-type models.

In the category of undiscovered resources, perhaps the 

most interesting is the possibility of producing a fluorine 

product from the high-fluorine, high-pH, sodium-carbonate 

brines of Lake Magadi (Kenya) and Lake Natron (Tanzania) 

in the East African Rift system. Also, apparently conformable 

fluorspar deposits in tuffaceous limy lacustrine sediments, 

such as those in Italy, are likely to occur in similar young 

alkalic volcanic settings elsewhere in the world.

Production and Consumption

Fluorspar mining began in the United States at various 

locations between 1820 and 1840. The first commercial 

production appears to have been in 1837 from a vein in 

Trumbull, Fairfield County, Connecticut. The existence of 

fluorspar in Illinois and Kentucky was known from the early 

19th century, but it was not until 1871 that the first commercial 

shipments from Kentucky to a glassworks in Baltimore were 

recorded (Ladoo, 1927, p. 2–3). Production was limited 

by demand to at most a few thousand metric tons per year 

until the late 1880s when demand for its use as a flux in 

steelmaking was bolstered by the introduction of the openhearth-furnace steelmaking process. The United States was 

the world’s leading producer of fluorspar until it was surpassed 

by Mexico in the mid-1950s (McDougal and Roberts, 1958).

U.S. domestic fluorspar production peaked during 

World War II at 375,000 metric tons in 1944 (Miller, 2002). 

In the post-World War II period, a booming economy 

and increasing consumer demand for products containing 

aluminum, fluorochemicals (especially chlorofluorocarbons), 

and steel continued to push U.S. consumption of fluorspar 

upward. Domestic production, however, was unable to keep 

up with rising demand and in the mid-1970s, the domestic 

fluorspar mining industry began to decline because of 

foreign competition. In 1996, primary fluorspar mining in the 

United States ceased, although small amounts continued to be 

recovered as a byproduct of quarrying activities in southern 

Illinois (fig. G7). Consumption of fluorspar peaked in the early 

1970s, which was, not coincidentally, also the peak period of 

U.S. steel production. Thereafter, decreasing steel production, 

a change in consumption practices by the steel industry, 

and the ban on chlorofluorocarbons when it was found that 

they were responsible for the “ozone hole” above Antarctica 

resulted in a general decrease in U.S. fluorspar consumption. 

From that time, consumption stabilized and then began to 
Resources and Production G29

increase slowly as replacement fluorocarbon compounds, such 

as hydrochlorofluorocarbons and hydrofluorocarbons, were 

introduced into the market (U.S. Bureau of Mines, 1972–96; 

U.S. Geological Survey, 1997–2016).

Despite substantial domestic production for much of 

the 20th century, the United States has been a net importer 

of fluorspar since the early 20th century. Sources of 

supply for the United States have changed over the years. 

Beginning during the period of peak U.S. consumption, the 

United States sources of supply have been dominated by a 

few countries—China, Mexico, and South Africa (fig. G7). In 

addition, between 1993 and 2006, the U.S. fluorspar supply 

was augmented by sales from the National Defense Stockpile. 

In 2012, annual world production and consumption was about 

7 million metric tons. The leading fluorspar producing countries were, in descending order, China, Mexico, Mongolia, 

and South Africa, which together accounted for 89 percent of 

estimated world production (Miller, 2016).

The global supply of fluorspar available for export 

decreased between 2001 and 2012 because of the substantial 

decreases in exports from China. China began exporting 

significant amounts of fluorspar in the 1960s, and the first 

exports to the United States took place in 1980 (fig. G7). 

China was the world’s leading fluorspar exporter from the 

early 1990s to 2008 or 2009. Beginning in the early 2000s, 

China’s fluorspar exports started to decrease as its domestic 

consumption increased. Producers outside of China were 

unable to increase production capacity to make up for 

the reduction in global supply. At the same time, global 

industrial activity increased, resulting in increased demand 

for fluorspar and higher prices (U.S. Bureau of Mines, 

1972–96; U.S. Geological Survey, 1997–2016).

In recent decades, the United States annual fluorspar 

consumption has ranged from 500,000 to 700,000 metric 

tons per year, although “fluorine” consumption is higher 

because the United States imports significant amounts of 

HF as well as aluminum fluoride and synthetic cryolite. In 

2012, the United States imported 133,000 metric tons of HF 

(equivalent to 293,000 metric tons of acid-grade fluorspar) 

and 58,100 metric tons of aluminum fluoride and synthetic 

cryolite (equivalent to an additional 85,000 metric tons of 

acid-grade fluorspar; Miller, 2016).

Fluorine, mostly in the form of fluorspar, is expected to 

continue to be important to the United States owing to its use 

in the manufacture of diverse materials and products necessary 

to the U.S. economy. The United States will continue to be 

import-reliant for most of its fluorspar needs, although in 

2012, mine production began at the Klondike II fluorspar mine 

in the Illinois-Kentucky fluorspar district, mentioned earlier 

(Feytis, 2009). An additional fluorine resource exists in the 

form of fluorosilicic acid, which is produced as a byproduct of 

phosphoric acid production in the phosphate fertilizer industry. 

In 2012, U.S. phosphate reserves were estimated to contain 

about 100 million metric tons of 100 percent CaF2-equivalent, 

and world reserves were estimated to contain 4.7 billion 

metric tons of 100 percent CaF2-equivalent. In 2012, about 

70,000 metric tons per year of 100 percent basis fluorosilicic 

acid was produced and used in the United States mostly for 

water fluoridation, although fluorosilicic acid can be used 

to produce aluminum fluoride, and the technology exists to 

convert it to HF (Miller, 2013).

The United States does have some remaining resources

of fluorspar, and those were discussed in the previous

section on Identified Resources in the United States.

Internationally, several new mines either have been opened

recently or are expected to open in the next few years, as

discussed in the previous section on Identified Resources

in Other Countries.

Figure G7. Graph showing sources of U.S. fluorspar supply from 1970 to 2011. The layers of the 

graph are placed one above the other, forming a cumulative total. Data are from U.S. Bureau of 

Mines (1972–96) and U.S. Geological Survey (1997–2016).

Year 

1970 1975 1980 1985 1990 1995 2000 2005 2010

Fluorspar Figure G8. 

0 

200 

400 

600 

800 

1,000 

1,200 

1,400 

1,600 

Fluorspar, in thousand metric tons 

Defense stockpile 

 shipments

Imports

 Others

 South Africa

 Mexico

 China

U.S. production 

EXPLANATION

U.S. source of supply
G30 Critical Mineral Resources of the United States—Fluorine

Exploration for New Deposits

None of the deposit types listed in figure G3 has been 

found in any oceanic geologic terrane, so apparently only 

continental rock terrains are favorable. Fluorite is commonly 

associated with end-member igneous differentiates, so 

exploration will profit from focusing on granites (all types, 

but particularly those not related to arc magmatism; that is, 

anorogenic granites), syenites, and carbonatites, and their 

volcanic equivalents. Fluorite is commonly found replacing 

limestone or dolostone. Thus, the optimal areas to explore 

are carbonates intruded by anorogenic granites, syenites, 

or carbonatites and carbonates in contact with anorogenic 

rhyolite or trachyte. The continental rifting association of 

salt-related carbonate-hosted fluorspar deposits could also be 

used to guide exploration; continental rifts have associated 

fluorspar deposits owing to their associated salt deposits and 

(or) their associated peralkaline to alkaline intrusive and 

volcanic rocks.

More specifically, provinces with strongly differentiated 

granites could be targeted for fluorspar skarn-and-greisen 

deposits. Although most of the belts of tin granites of the 

world have been identified, most of these belts have not been 

thoroughly explored for either tin or fluorspar. Some of the 

belts are located in places that are difficult to explore owing 

either to climate (for example, the Seward Peninsula tin 

granite province in Alaska and the Southeast Asian tin granite 

province) or other factors (for example, political risk in the 

Nigerian province). Taylor (1979, p. 6) has mapped 42 tin 

provinces worldwide, and provided consistently structured 

packages of information on 23 of these provinces (Taylor, 

1979, p. 491–511). Not all of the provinces that have tin 

deposits have tin granites, however.

No fluorspar deposits analogous to those of the Latium 

Region, Italy, have been discovered to date, but that they 

would be unique in the world seems unlikely. The broadest 

scale indicator for these types of deposits would appear to 

be young alkalic volcanic rocks with interlayered lacustrine 

sedimentary rocks, and such volcanic rocks are widespread 

across the globe. As an initial test, a keyword search of 

geologic literature conducted on the combination of terms 

“phonolite” and “maar,” and a second search conducted 

on “trachyte” and “maar” yielded 15 areas of the world (in 

addition to central Italy). They are (1) the Czech Republic, 

(2) southern Italy, (3) northern South Africa, (4) western 

Germany, (5) east-central Germany, (6) southern Germany, 

(7) Hungary, (8) central France, (9) the Canary Islands, 

Spain, (10) Slovakia, (11) Shandong Province, eastern China, 

(12) Hawaii, (13) Big Bend National Park, Tex., (14) Pinacate 

volcanic field, northwestern Mexico, and (15) western 

Cameroon (including maar Lake Nyos in North-West Region, 

where a catastrophic lake-water overturn released CO2 from 

depth in the lake and killed 1,700 people in 1986). Among 

these areas, the Canary Islands and Hawaii seem unlikely 

possibilities because no economic fluorspar deposits are 

known to occur on any oceanic islands. The Big Bend area 

of Texas, on the other hand, is firmly within the belt of 

fluorspar deposits in that region. Also, the Aguachile deposit 

in Coahuila State, Mexico, consists of limestone replacements 

adjacent to rhyolite porphyry ring dikes of a caldera; these 

replacement deposits may have formed beneath a caldera lake.

As mentioned previously, an unconventional source of 

fluorine could be brines from Lake Magadi and Lake Natron 

along the East African Rift, as well as other places where such 

alkaline evaporite brines may occur.

In the past, exploration for fluorspar deposits has not 

used methods comparably sophisticated with those used in 

exploring for such minerals as gold and copper. Most deposits 

have been found by simply following up occurrences of 

outcropping fluorite. It was common for mines to drill only 

enough to establish a few years’ worth of reserves ahead of the 

mining operation. This has probably begun to change since the 

advent of such requirements as those of the Canadian Government NI 43–101 reports and Australia’s JORC-compliant 

resource and reserve estimates. The requirements of such 

publicly released documents has encouraged the estimation 

of total deposit resources to be known before financing and 

various government permitting approvals are granted, or it 

has at least encouraged the definition of resources adequate 

for many years of mine life. Even just a few decades ago, 

geophysical methods were seldom used in exploration for 

fluorspar. Geochemical exploration methods were slightly 

more common, but still not heavily used. The future probably 

will see the use of increasingly sophisticated geophysical and 

geochemical methods of exploration.

Grogan (1960, 1964) was of the opinion that geochemical 

prospecting was not well adapted to finding fluorspar deposits 

because he thought that fluorite is too effectively dispersed, 

being soft and extremely cleavable. Geochemical prospecting 

for fluorspar deposits has succeeded in some areas, however. 

Van Alstine (1965) demonstrated effective results from 

sampling and analyzing rock or soil for fluorine in various 

areas within the Browns Canyon district in Chaffee County, 

Colo. In addition, geochemical prospecting was shown to be 

effective in an area near the Osor vein in the eastern Pyrenees 

in Spain’s Catalonia Autonomous Region (Schwartz and 

Friedrich, 1973). Near Osor, broad-scale prospecting using 

measured fluoride in surface water identified target areas 

of 3 to 10 km2. Dry stream-sediment samples narrowed the 

search to targets of less than 1 km2 and led geologists to 

mineralized outcrops. Then, fluoride in soil samples was used 

to map small areas and guide placement of drill holes. The 

work resulted in discovery of an area of fluorite veinlet stockwork. In addition, surface waters proved to be an effective 

medium for district-scale exploration. This is likely because 

of the relatively high solubility of fluorine at low temperatures 

and because of fluorite’s increasing solubility with decreasing 

temperature below about 60 °C (fig. G2).

Three different studies (Graham and others, 1975; 

Deering and others, 1983; and Ebbott and others, 1985) have 
Environmental Considerations G31

shown that groundwater sampled from wells revealed areas 

of high fluorine concentration broadly centered on areas of 

fluorite-bearing MVT mineralization. This method would 

not be applicable in areas of greenfield exploration because 

the distribution of water wells into a target aquifer would be 

insufficient in remote areas. The method is applicable in the 

agricultural areas of central North America, although only in 

defining an area of tens of square kilometers as anomalous.

An extensive rock geochemical study of the Taskaynar 

deposit related to a strongly differentiated granite in 

Kazakhstan demonstrated that several additional elements 

had primary dispersion patterns similar to that of fluorine, 

and, consequently, might be used as “pathfinders” in exploration, especially to guide exploration drilling. The Taskaynar 

deposit is hosted by sandstone that lies above limestone with 

nearby “porphyrite” and granodiorite. The elements Ag, As, 

Ba, La, Mo, Sr, Y, and Yb all had primary dispersion patterns 

that were as far-reaching from the ore as fluorine itself. The 

dispersion patterns of lead, zinc, and sodium+potassium were 

somewhat more restricted around the orebody, but also had 

abundance boundaries parallel to the fluorine boundaries, and 

both showed control by fracturing (Koplus and others, 1977).

In exploring for skarn-and-greisen fluorspar deposits 

related to strongly differentiated granites, several more 

sophisticated methods could be brought to bear. Gravity and 

magnetic geophysical studies could be used to determine 

the depths to crystalline rocks through sedimentary cover, 

with the magnetics also directly targeting magnetite-bearing 

skarn. For a deposit standing upright and arrayed around 

a granitic cupola on a larger batholith, detailed magnetic 

surveys could help identify magnetite-bearing skarn, probably 

forming a magnetic ridge in a partial ring around a magnetic 

low, the low produced by greisen with a nonmagnetic pyritemuscovite(-fluorite) assemblage. Fluorspar system-related 

but locally developed semimassive sulfide bodies typically 

rich in sphalerite-pyrrhotite-arsenopyrite can be detected by 

a number of different methods that show anomalies in rock 

conductivities. An airborne electromagnetic survey using the 

Barringer INPUT (induced pulse transient) method detected 

such a sulfide-rich body at the Ar Ridaniyah fluorite-bearing 

tin skarn system located 200 km west of Riyadh, Saudi Arabia, 

in 1967 (Canadian Aero Mineral Surveys, Ltd. and Arabian 

Geophysical and Surveying Company, 1967).

Most vein fluorspar deposits of various types are also 

detectable using methods that measure conductivities. 

Successful exploration of fluorite veins in granitic terrain 

using the combination of very low frequency (VLF) electromagnetics and audio-magnetotelluric (AMT) methods was 

reported by Zhang and others (2010). Vein fluorspar deposits 

are amenable to discovery by such methods despite the low 

conductivity of fluorite. Anomalous amounts of saline groundwater in fractured porous rock of fault zones mineralized by 

fluorite and (or) metallic sulfide minerals associated in minor 

concentrations with the vein fluorite are detectable. In China, 

VLF electromagnetics provided clear resolution of veins and 

accurate positioning of their location near the surface, and 

AMT methods provided accurate location of the veins at 

greater depth (Zhang and others, 2010). These methods could 

work effectively as long as the host rocks of the veins have 

high resistivities (that is, low conductivities) so that the vein 

zones contrast sufficiently.

Environmental Considerations

Sources and Fate in the Environment

Fluorine is chemically and biologically active. It readily 

forms hydrogen fluoride or combines with metals to make 

fluoride compounds, some of which are relatively stable, such 

as calcium fluoride (fluorite). The weathering of rocks and 

soils may release dissolved fluoride into the environment, and 

fluoride is added to drinking water in some countries, which 

is then released into surface water in municipal wastewater. 

Clay content, pH, temperature, and the concentrations of 

aluminum, calcium, magnesium, phosphorus, and hydroxides 

all influence the dissolved fluoride content of water. Fluoride 

compounds generally are most soluble and thus mobile under 

acidic conditions. In contrast, alkaline conditions in the 

presence of carbonates limit the mobility of fluorine in the 

environment; fluorine is fixed by reaction with carbonates to 

form stable fluorides. In the absence of carbonates, however, 

alkaline evaporitic conditions yield the greatest known 

solubility of fluorine at ambient temperature (table G2). Also, 

fluorine is easily sorbed by clay minerals (Kabata-Pendias and 

Mukherjee, 2007). As previously mentioned, fluorite tends to 

dissolve as temperature decreases (that is, it has “retrograde” 

solubility) under most earth-surface temperatures (60 °C 

to freezing, depending on the composition of the water; 

Richardson and Holland, 1979).

Fluorine is present in most natural waters. River waters 

commonly contain between 0.05 to 2.7 ppm fluorine, whereas 

seawater generally contains between 0.03 to 1.4 ppm with a 

median of approximately 1.3 ppm fluorine (table G2). High 

concentrations of fluorine have been shown to be present in 

waters in regions of volcanic activity or granitic bedrock, 

particularly in arid areas, and where fluorite or phosphates are 

mined and processed. China, East Africa, and India have the 

most extensive areas of high-fluorine groundwaters (Jacks and 

others, 2005). Table G2 includes examples of both naturally 

occurring and mine-influenced concentrations of fluorine in 

waters. One extreme example of naturally occurring concentrations of fluorine in surface waters is in Africa’s Eastern 

Rift Valley in Kenya and Tanzania where concentrations 

reached 437 ppm from the weathering of volcanic rocks; 

evaporation also concentrates the fluorine in the water, locally 

elevating the fluorine concentration to above 2,000 ppm in 

closed basin lakes in that setting (table G2; Kilham and Hecky, 

1973; Jones and others, 1977). An example of anomalous 

mine-affected waters is that within mine workings at the 
G32 Critical Mineral Resources of the United States—Fluorine

now-closed Lovozero niobium mine in Russia that were highly 

alkaline (pH 9.6 to 12) and contained extreme concentrations 

of fluorine (1 to 1.5 percent) and silica (1 to 1.3 percent) 

(table G2; Kraynov and others, 1969).

The amount of fluorine in soils is a reflection of its 

content in parent rocks as influenced by soil-forming 

processes, soil texture, and climatic conditions. Concentrations 

of fluorine in most soils range between 150 to 400 ppm 

(Kabata-Pendias and Mukherjee, 2007). For soils in the 

conterminous United States, fluorine concentrations average 

430 ppm and range from less than 10 to 3,700 ppm (Shacklette 

and Boerngen, 1984). The lowest fluorine content is found in 

sandy soils under humid climate conditions and the highest 

content is in clay soils and soil derived from mafic rocks 

(Kabata-Pendias and Mukherjee, 2007). High levels of 

fluorine (more than 1,000 ppm) have commonly been reported 

for uncontaminated soils derived from fluorine-rich bedrock 

(Fuge and Andrews, 1988; Ermakov, 2004). Contaminated 

soils surrounding aluminum processing plants, china clay 

production facilities, and phosphate-fertilizer factories have 

been reported to contain between 2,000 and 3,600 ppm 

fluorine (Kabata-Pendias and Pendias, 2001).

The major natural source of atmospheric fluorine in 

the environment is volcanic activity, with sea-salt spray and 

suspension of soil by wind as additional natural sources. 

Anthropogenic sources include mining and processing of 

phosphorites, apatites, and fluorite; production of cement, 

bricks, steel, aluminum and other metals; combustion of coal; 

and production and use of fertilizers, fluorine-bearing pesticides, and drugs, plastics, and refrigerants (Ermakov, 2004). 

Fluorine is removed from the atmosphere primarily by wet 

deposition either as particulate fluorides or by dissolving into 

atmospheric water. Fluorine concentrations in the atmosphere 

are variable. The ambient air concentration of gaseous fluoride 

varies from 0.01 to 2.1 parts per billion (0.01 to 2.1 micrograms per cubic meter) in the United States and Canada, 

with about 75 percent as hydrogen fluoride (Bourgeau and 

others, 1996).

The geochemical signatures in soils and waters associated 

with fluorspar deposits before mining are variable because 

fluorite is common in a wide range of geologic environments. 

The chemistry of soils, streams, and groundwaters near 

fluorspar deposits reflect the composition of the ore and host 

rock, the structure of the deposit, and the climatic conditions. 

Streamwater and groundwater near deposits may not display 

the same geochemical signature as soils and stream sediments 

because some of the enriched elements occur in relatively 

stable insoluble minerals. For example, soils and weathered 

rock associated with carbonatite deposits may be naturally 

enriched in Ba, Fe, Nb, P, REEs, Th, Ti, U, and Zr, and stream 

sediments downstream from deposits commonly contain 

anomalous abundances of Ba, Nb, Th, and REEs (Modreski 

and others, 1995). The content of such elements as Nb, REEs, 

Th, and U may not be anomalous in streamwaters, however, 

because they are present in relatively stable, insoluble 

minerals (Modreski and others, 1995). The pre-mining 

environment around alkaline- to peralkaline-related deposits 

may contain anomalous concentrations of some of the same 

elements as carbonatites, including Be, Li, Nb, REEs, and Ta. 

Streamwaters near the Yermakovskoye deposit in Russia 

contained naturally elevated concentrations of Cu, F, Fe, Hg, 

Mn, Mo, and Zn. In addition, native soils near the deposit 

contained anomalous concentrations of As, Ba, Be, and F 

(Kislov and others, 2010). Natural metal-rich acidic drainage 

could be significant from deposits associated with significant 

pyrite or other sulfides. Carbonates are nearly ubiquitous at 

fluorspar deposits and may neutralize the acid generated by 

sulfides by removing many trace elements from the solution. 

In MVT deposits, iron, lead, and zinc from the weathering of 

associated sulfides may occur in anomalous concentrations 

in soils and stream sediments; however, at most deposits, 

elevated concentrations of these metals are not present because 

carbonate rocks limit the dispersion of trace elements by 

precipitating them as carbonates (Leach and others, 1995).

In the United States, some fluoride compounds are 

recycled in industrial settings. For example, synthetic fluorite 

is recovered from uranium enrichment and, to a lesser extent, 

from petroleum alkylation and stainless-steel pickling, 

producing on the order of a few thousand metric tons per 

year. In addition, aluminum producers recycle hydrofluoric 

acid and fluorides from smelting operations, and hydrofluoric 

acid is recycled in the petroleum alkylation process (Miller, 

2013). New methods are being developed to recover fluoride 

from industrial wastewaters. One of these methods produced 

a nearly pure synthetic fluorite (>97 percent) that can be 

used for the manufacture of hydrofluoric acid (Aldaco and 

others, 2007).

Mine Waste Characteristics

Mine waste is the material generated during ore 

extraction and processing at the mine site that has no direct 

economic value. The volume and characteristics of the 

waste material depend on the type of deposit and its size, 

the commodity or commodities recovered, and the mining 

and processing methods used. A summary of the mineralogy, 

chemistry, and volume, if available, of mine waste generated 

at several fluorspar mines is presented in table G3.

In general, overburden may need to be removed or host 

rock separated from the fluorite-rich ores as the first step of 

ore extraction; the fluorite-poor material becomes waste rock 

and can be backfilled into mine workings or pits, disposed 

of in waste piles, or, depending on the composition, used or 

sold for reuse (for example, as aggregate for construction 

purposes). The next step is to concentrate the fluorite in the 

ore. This is accomplished using a variety of methods, which 

typically include crushing and grinding followed by gravity 

separation and (or) flotation. The gangue minerals become the 

major components of waste tailings; common gangue minerals 
Environmental Considerations G33

include calcite, dolomite, quartz, sulfides, and (or) barite, 

and, for carbonatite-related deposits, possibly apatite. Gravity 

separation is most often used to separate fluorite from calcite 

and silicates. Chemical flotation is used, however, when barite, 

celestite, or sulfide minerals are present, as they are denser 

than fluorite. Tailings generated from flotation are usually 

disposed of as an aqueous slurry into a tailings impoundment, 

where they are allowed to dry.

It is common for mine waste to contain a significant 

amount of fluorite as a result of inefficient or nonrecovery in 

years past; at some places, the fluorite content is high enough 

that the waste can be reprocessed using current methods 

(most commonly flotation) to recover most of the remaining 

fluorite. For example, mine waste and tailings from several 

locations in Kentucky (Crittenden and Livingston Counties) 

were reprocessed during World War II to meet the increased 

demand for fluorite (Fine, 1948). The tailings and waste rock 

contained a significant amount of fluorite (up to 59 percent), 

as well as calcite, chert, clay, and quartz; a minor amount of 

barite; and some lead and zinc sulfides (table G3). Another 

example of mine waste with a significant amount of fluorite 

(up to 17 percent total fluoride) is from fluorite and lead-zinc 

mines from the Northern Pennine Orefield and the Southern 

Pennine Orefield in the United Kingdom (table G3; Cooke 

and others, 1976; Fuge and Andrews, 1988). Modern recovery 

in fluorite operations is generally more efficient than in 

earlier operations.

Human Health Concerns

Fluorine is essential for some organisms but, like most 

substances, in excess, is highly toxic. Soluble forms of 

fluorine are absorbed into the body through the mouth lining, 

stomach, and intestines. Nearly all fluorine in the human 

body is in the form of fluorapatite, which is resident mostly in 

the bone, teeth, and cartilage (Ermakov, 2004). The primary 

chronic exposure route of concern for fluorine compounds is 

oral ingestion of drinking water, food, and fluoride-containing 

dental products. Inhalation and dermal exposure to HF is 

an acute occupational exposure that can occur (Agency for 

Toxic Substances and Disease Registry, 2003).

Fluorine is important for hard teeth in humans. In the 

United States, artificially fluoridated water commonly contains 

between 0.7 and 1.2 ppm fluorine (National Research Council, 

2006). Fluorine deficiency in humans (less than 0.3 to 0.5 ppm 

in drinking water) leads to tooth decay and osteoporosis. Also, 

animals with diets low in fluorine have exhibited slow growth, 

reproduction issues, and higher mortality (Ermakov, 2004). On 

the other hand, consumption of water with high concentrations 

of fluorine may lead to dental and skeletal fluorosis (that is, 

mottling of the tooth enamel, and pain and damage to bones 

and joints, respectively) starting at concentrations of 1.5 ppm 

and becoming severe at greater than 6 ppm (Reimann and de 

Caritat, 1998). Other health effects related to a high intake of 

fluorine include endocrine effects and carcinogenic changes, 

particularly in bone (National Research Council, 2006). 

According to the U.S. Environmental Protection Agency 

(2002), the no-observable-effect level for dental fluorosis 

based on children is approximately 1.0 ppm fluoride, 

whereas crippling skeletal fluorosis, the most severe stage 

of skeletal fluorosis, was not observed in adults at consumption levels of up to 4 ppm fluoride. In the United States, 

primary and secondary drinking water standards have been 

established for fluorine. The primary drinking water standard 

(maximum limit of 4 ppm fluorine) was established to 

prevent adverse health effects and the secondary standard 

(maximum limit of 2.0 ppm fluorine) was established to 

reduce adverse cosmetic consequences (U.S. Environmental Protection Agency, 2009). A recent review of the 

U.S. drinking water guidelines by the National Research 

Council (2006) suggests that the maximum contaminant 

level of 4 ppm should be lower. For comparison, the World 

Health Organization reports a guideline value of 1.5 ppm 

(World Health Organization, 2008).

Skeletal deformations and the crippling bone deformities of genu valgum (that is, knock knee) have been reported 

to occur in regions in India, Kenya, and Tanzania where high 

levels of fluorine (3 to 8 ppm) are present in drinking water 

(Ermakov, 2004). Also, surface water, groundwater, and mine 

water from Kerio Valley, Kenya, in Africa’s Eastern Rift 

Valley, reaches 305 ppm, with average values greater than 

25 ppm fluoride (table 2; Davies, 1994); tooth decay, dental 

and skeletal fluorosis, and osteochondral conditions are 

common in this area. Regional groundwater studies in China 

and Mexico also link high fluorine groundwaters to high 

fluorine rocks and fluorspar mines with some documentation 

of serious human health effects (Fuhong and Shuqin, 1988; 

Mahlknecht and others, 2004).

Although severe human health effects from high 

fluoride in drinking water are not commonly reported in the 

United States, naturally occurring fluoride concentrations in 

waters do exceed U.S. drinking water standards in several 

areas. In 1992, approximately 10 million people had public 

water supplies that were naturally fluoridated. The concentration of fluoride in 14 percent of those people’s drinking water 

supplies were between 2.0 and 3.9 ppm; about 2 percent 

of those people’s water supplies had concentrations equal 

to or greater than 4.0 ppm fluoride. The highest, but not 

necessarily typical, concentrations were found in Colorado, 

Oklahoma, New Mexico, and Idaho, with maximum values 

of greater than 10 ppm fluoride (National Research Council, 

2006). In Texas, where groundwater makes up about 

60 percent of the drinking water consumed, well water from 

a northwest region in the state exceeded the drinking water 

standard of 4 ppm fluoride in more than 50 percent of the 

samples (Agency for Toxic Substances and Disease Registry, 

2003). For those areas with fluoride greater than 4 ppm in the 

drinking water, the Agency for Toxic Substances and Disease 

Registry recommends buying bottled drinking water.
G34 Critical Mineral Resources of the United States—Fluorine

For soils, screening levels based on human health risk 

in the United States have been established based on the 

use of the land (residential versus industrial) to determine 

whether concentrations of elements or compounds at sites 

warrant further investigation or site cleanup. Screening 

levels for noncancer health effects are 3,100 ppm fluoride 

and 4,700 ppm fluorine (soluble fluoride) for residential land 

use and 41,000 ppm for fluoride and 61,000 ppm for fluorine 

(soluble fluoride) for industrial land use (U.S. Environmental 

Protection Agency, 2012). Natural soils in the United States, 

even those developed on fluorine-rich rocks, will almost never 

contain such concentrations, and if so, only in isolated small 

areas; however, such concentrations may be more common in 

soils located in industrial areas.

Other human health concerns related to fluorine are 

related to the mining and processing of fluorspar ores, 

which may expose workers to dust or, for some types of 

ores, radiation. Silica is associated with many types of ore, 

including fluorspar, and may be inhaled during mining 

and milling operations and can cause silicosis. In addition, 

inhaling high amounts of fluorspar dust during ore processing 

can lead to health issues that include gastric, intestinal, circulatory system, and nervous system problems, and, in the long 

term, anemia, weight loss, bone and teeth defects, pulmonary 

lesions, and bronchitis (Osinsky and Stellman, 1998). In 

the United States, the permissible exposure limits in air for 

fluorine in an 8-hour working day is 0.3 ppm (0.2 milligram 

per cubic meter) (Occupational Safety and Health Administration, 2013). Dust control should be carefully enforced 

for mining and processing of ores from any type of fluorspar 

deposit. In addition to dust, human health issues have been 

documented from exposure to radiation during the mining 

of deposits related to peralkaline granite in the St. Lawrence 

district in Newfoundland, Canada. In this district, higher 

incidences of lung cancer were reported for underground 

miners exposed to high levels of radon and its progeny during 

fluorspar mining; mining started in the 1930s and continued 

until 1978. In the early 1960s, mechanical ventilation was 

installed, which decreased radon to levels below the exposure 

limit (Morrison and others, 1998).

An additional human health concern associated with 

mining operations that process ores using flotation techniques, 

which include some fluorspar mines, is the failure of tailings 

impoundments. Although rare, collapse of tailings dams can 

destroy property and cause injuries, fatalities, and environmental damage. For example, a fluorspar mine in Italy that 

used flotation to process ores from 1961 to 1985 stored tailings 

and waste in two large tailings impoundments. In 1985, 

approximately 180,000 cubic meters (m3 ) of tailings slurry 

was inadvertently released and flowed into a valley, destroying 

a small town and causing destruction in another small town; 

this event caused 268 deaths and 100 injuries. The failure was 

owing to construction errors and a lack of operational monitoring during tailings deposition (Luino and De Graff, 2012). 

In more recent times, technical and scientific investigations on 

tailings impoundments and new legislation and regulations for 

stricter control and monitoring are helping to minimize tailings 

dam failures.

Table G3. Selected examples of the chemistry and mineralogy of mine waste generated at fluorspar mines.

[Elements: Ca, calcium; Cd, cadmium; Cu, copper; F, fluorine; Fe, iron; Ni, nickel; P, phosphorus; Pb, lead; S, sulfur; Si, silicon; Zn, zinc. Abbreviations and 

symbols: ppm, part per million; wt. %, weight percent; —, not determined or not reported; ?, identification uncertain]

Sample type Location Deposit type Minerals/compounds/components 

Tailings and mine waste 

(pre-World War II)

Several locations in Crittenden 

and Livingston Counties, 

Kentucky, United States

Mississippi 

Valley-type

Fluorite (13 to 59 wt. %), calcite (2 to 28 wt. %), quartz

(8 to 65 wt. %); usually with smaller amounts of clay, 

barite (0 to 2 wt. %), Pb, Zn, Fe sulfides (sphalerite, 

galena, pyrite)

Tailings Southern Pyrenees, France ?Salt-related, 

carbonatehosted

Quartz, shale, iron hydroxides, fluorite (1 to 5 wt. %), 

iron carbonate, iron/copper sulfides

Mine waste1 Hammam Zriba Mine, Tunisia Salt-related, 

carbonatehosted

Quartz, barite, calcite > K-feldspar, sylvite > plagioclase, 

fluorite, limonite, zeolites, opaque minerals

Mine waste Northern Pennine Orefield and 

Southern Pennine Orefield, 

United Kingdom

Mississippi 

Valley-type

Barite (16 to 23 wt. %), aluminosilicates (20 to 51 wt. %), 

calcium carbonate (15 to 32 wt. %), fluorite, among 

others

1

Chemistry is for a partial digestion procedure (that is, silicates were not completely digested).
Environmental Considerations G35

Ecological Health Concerns

Only small amounts of fluorine are taken up by plants 

from soils. Terrestrial plants from uncontaminated soils 

usually do not contain greater than 30 ppm fluorine and likely 

do not require fluorine (Kabata-Pendias and Mukherjee, 2007). 

Plants, in particular tea plants (Camellia sinensis), however, 

can accumulate up to 2,000 ppm fluorine in areas with high 

soil concentrations. Levels of fluorine in soils between 

200 and 2,000 ppm have been shown to result in toxicity to 

plants (that is, phytotoxicity) (Ermakov, 2004). Food plants 

commonly contain between 0.1 and 11 ppm (Kabata-Pendias 

and Mukherjee, 2007). Fluorine on and within the plant parts 

can enter the food chain by animals foraging on the plants, and 

toxic threshold values for fluorine in fodder has been reported 

to be in the range of 20 to 50 ppm (Ermakov, 2004). Fluorine 

toxicity in dairy cattle has resulted in lower reproduction and 

lower milk production (Swarup and others, 1998). Studies 

on minks and screech owls reported effects on fertility and 

reproduction success when fluorine concentrations reached 

hundreds of ppm in food (Aulerich and others, 1987; Pattee 

and others, 1988). Marine aquatic organisms have been 

shown to accumulate fluorine, and fluorine concentrations 

are commonly an order of magnitude higher (for example, 

500 ppm compared with 50 ppm) in sea organisms compared 

with freshwater fish and land organisms because of the generally higher fluorine levels in seawater compared to freshwater 

(Ermakov, 2004). Industrial releases of high amounts fluorine 

in waterways have had toxic effects on aquatic organisms 

(Davies, 1994; Agency for Toxic Substances and Disease 

Registry, 2003). Water-quality criteria and stream-sediment 

guidelines for the protection of aquatic life are generally 

lacking. Suter (1996), however, suggested an estimate of the 

highest concentration in surface water to which an aquatic 

community can be exposed briefly (that is, acute exposure) 

and indefinitely (that is, chronic exposure) without resulting 

in an unacceptable effect; these secondary acute and chronic 

aquatic ecosystem guidelines, established with fewer data 

than regulatory guidelines, are 19.2 and 1.18 ppm fluorine, 

respectively. These guideline concentrations are higher than 

those found in the vast majority of natural surface waters in 

the United States.

The mining and processing of ores may enrich waters and 

soils in potentially toxic elements and can cause ecological 

health concerns at some mine sites. Several examples from 

a few different fluorspar deposit types are discussed below. 

In general, ecosystem impact studies are more frequently 

conducted for sulfide-containing deposits, such as MVT 

deposits, because sulfides can generate acid and contribute to 

the release of some potentially toxic trace elements into the 

environment. Several studies have been conducted regarding 

the environmental impact of mining lead and fluorspar 

from MVT deposits in the Northern Pennine Orefield and 

Southern Pennine Orefield of the United Kingdom. The mine 

waste from the mines contains high concentrations of some 

trace elements as shown in table G3. Waters from one mine 

contained high concentrations of, in order of greatest environmental concern, zinc (up to 111 ppm), manganese (up to 

138 ppm), iron (up to 68 ppm), and sulfate (up to 2,501 ppm), 

and, although most waters were near neutral, some were 

acidic, likely from sulfides in shales and mudstones (table G2; 

Johnson and Younger, 2002). Fluorine concentrations in soils 

surrounding mines in the Northern Pennine Orefield reach up 

to 20,000 ppm; surface waters, up to 2.3 ppm fluorine, and 

Table G3. Selected examples of the chemistry and mineralogy of mine waste generated at fluorspar mines.—Continued

[Elements: Ca, calcium; Cd, cadmium; Cu, copper; F, fluorine; Fe, iron; Ni, nickel; P, phosphorus; Pb, lead; S, sulfur; Si, silicon; Zn, zinc. Abbreviations and 

symbols: ppm, part per million; wt. %, weight percent; —, not determined or not reported]

pH Ca

(wt. %)

F

(wt. %)

Fe

(wt. %)

P

(wt. %)

S

(wt. %)

Si

(wt. %)

Cd

(ppm)

Cu

(ppm)

Ni

(ppm)

Pb

(ppm)

Zn

(ppm) Reference

— — — — — 0.03 to 

2.2

— — — — <500 to 

8,200

1,000 to 

38,300

Fine

(1948)

— — — 3.5 to

7

— — 37 to

42

— — — — — European 

Commission 

(2009)

— 7.8 — 0.31 0.11 0.3 — 36 34 0.5 3,420 12,100 Yoshida 

and others 

(2002)

6.2 to 

7.7

— 2.3 to 

17.4

— 0.001 to 

0.06

— — 8 to

28

32 to

67

30 to

65

4,360 to 

30,800

2,020 to 

20,360

Cooke 

and others 

(1976)
G36 Critical Mineral Resources of the United States—Fluorine

grasses up to about 400 ppm fluorine (table G2; Fuge and 

Andrews, 1988). A study of plants growing on mine waste in 

both the Northern Pennine Orefield and the Southern Pennine 

Orefield revealed fluoride concentrations of up to 4,500 ppm 

in leaves and up to 10,000 ppm in seeds; the plants did not 

shown signs of phytotoxicity (Cooke and others, 1976). 

Some waste piles have been revegetated to create natural 

ecosystems, and small mammals living on and near the 

revegetated piles have been shown to contain significantly 

higher levels of cadmium, fluorine, and lead and somewhat 

higher levels of zinc than do small mammals living in uncontaminated areas (Wright and others, 1978; Andrews, Johnson, 

and Cooke, 1984, 1989a, b: Andrews, Cooke, and Johnson, 

1989). Geeson and others (1998) suggested health problems 

affecting young lambs living in the vicinity of old mines or 

smelters in the Southern Pennine Orefield may be related to 

the excess lead as well as fluorine in their diets.

Ecological health impacts have also been studied at a 

peralkaline to alkaline-intrusion-related fluorspar deposit in 

Russia that contains only a minor amount of sulfides. Waters 

from the open pit at the previously mined Yermakovskoye 

beryllium-fluorine deposit contained elevated concentrations 

of some elements, including Cu, Fe, Mg, Mn, Mo, Pb, and 

Zn, all of which exceeded local standards for water bodies 

used for fishery purposes. In addition, sediment at the bottom 

of the pit was contaminated with As, Be, Cd, Pb, and Zn, and 

the concentrations of Pb and Zn exceeded values above which 

effects to aquatic biota are likely. Mine waste piles contained 

concentrations of beryllium that were above the maximum 

allowable concentrations (Kislov and others, 2010). The 

natural background contents of some of the elements, such as 

Cu, F, Fe, Mn, Mo, and Zn, in local streams and melt waters 

are also elevated and above ecological-health guidelines, 

making it difficult to distinguish the contamination from 

mining from the pre-mining concentrations associated with the 

mineralized area. According to Kislov and others (2010), the 

pasture vegetation in the local area may contain high levels of 

beryllium from the deposition of contaminated dust, and the 

agricultural uses of the land should be limited.

The potential ecosystem health concern associated with 

the Moscona Mine, which is a salt-related carbonate-hosted 

fluorspar deposit in Asturias, Spain, is related to aqueous 

fluoride; acidic metal-rich mine drainage from the weathering 

of sulfides is not an issue because of neutralization by 

associated carbonates. Waters from within the underground 

mine and from a surface spring are saline (800 to 1,700 microsiemens per centimeter conductivity), have near neutral 

pH (pH 7.6 to 8.2), and contain elevated levels of fluoride 

(0.5 to 3.5 ppm), carbonate, and sulfate (table G2; Roqueñí 

and others, 2005). Also, each year approximately 200,000 m3

of wastewater from the mine is discharged into a nearby 

stream, which results in fluoride concentrations of up to 

4.8 ppm. The release of this water, some of which exceeds the 

secondary chronic surface water guideline for the protection 

of aquatic life as suggested by Suter (1996), may be a threat to 

the health of the local ecosystem: however, decant ponds were 

constructed to reduce the contamination of the nearby stream 

(Roqueñí and others, 2005).

Carbon Footprint

A new fluorine-based compound, hydrofluoroolefin, is 

recommended as a refrigerant for air conditioning in motor 

vehicles to replace an existing compound, chlorofluorocarbon, 

currently in use. The new compound has a lower global 

warming potential, which is a relative measure of the warming 

effects of greenhouse gases, than the fluorocarbon compound 

and has been approved by the U.S. Environmental Protection 

Agency as a replacement (U.S. Environmental Protection 

Agency, 2011). Another advantage of the replacement 

compound is that it may contribute to the reduction of ozone 

concentrations in the stratosphere.

Mine Closure

The post-mining landscape at mines, including fluorspar 

mines, depends on the deposit type and size, the character of 

the host rock, the mining and processing methods, the climate, 

and the regulations in place regarding mine closure. Open 

pits, shafts and adits from underground operations, processing 

facilities, mine and processing wastes, and drainage from 

the site may need to be addressed for mine closure. One 

example of a remediated mine site in the United States is the 

Burlington Mine, which was the largest fluorspar deposit in 

the Jamestown district in Boulder County, Colo.; this site 

was reclaimed through a Voluntary Clean Up action, which 

is a State cooperative program that encourages voluntary 

cleanup of contaminated industrial and commercial properties (O’Shea-Stone, 2011). The mine produced fluorspar, in 

addition to some silver, lead, and uranium, from 1920 to 1973; 

limited remediation was conducted at the time of mine closure, 

and a Voluntary Clean Up action was conducted within the 

past 10 years. Prior to the recent cleanup, the landscape 

included a large vegetation-free, acidic, waste pile, a mine adit 

and shaft openings, and subsidence features. In addition, water 

draining from the waste pile flowed into a nearby creek. The 

remediation efforts consisted of consolidating acid-generating 

waste rock, closing of the adit and shafts, filling subsidence 

pits, limiting surface water and groundwater interaction with 

waste materials and mine workings, adding soil amendments 

(including lime, soil, and compost) to waste rock piles and 

disturbed soils, and site-wide revegetation. In 2008, remediation efforts were considered successful by State and Federal 

agencies based on improved water quality, minimal erosion 

from the site, removal of safety hazards, and the establishment 

of a natural and functional self-sustaining plant community 

(O’Shea-Stone, 2011).

Other mines have not been reclaimed. For example, 

remediation efforts were not undertaken when production 
Acknowledgments G37

ceased in 1989 at the Yermakovskoye beryllium-fluorine 

deposit in Russia. The surface area of the open pit is 

277,000 square meters with a depth of at least 70 m, and it 

contains a few meters of water in the bottom (Kislov and 

others, 2010). The impact to the ecosystem from the mined 

landscape was discussed in the previous section. Future 

plans for the site include resuming mining and construction 

of a processing plant; construction was expected to begin in 

2013 (Jaskula, 2013). An estimated 8,500,000 m3 of rock will 

be extracted from an open pit operation, including ore and 

waste rock (Kislov and others, 2010). The combined volume 

of sulfide-rich and nonsulfide-rich tailings generated will 

be about 830,000 m3. Eventually, mine closure will involve 

stabilizing disturbed surfaces and reclaiming soils, waters, 

and stream sediments at and near the mine site.

Problems and Future Research

The Klondike II Mine (Livingston County, Ky.) began 

production in 2012, the Nui Phao Mine (Thai Nguyen Province, Vietnam) began production in 2013, and, until late 2013, 

the Nokeng Mine (Gauteng Province, South Africa) had been 

scheduled to start production in the second quarter of 2014. 

A number of other projects were projected to begin mining 

in the 2015–17 timeframe. With these new supplies, Chinese 

domination of the world fluorspar market is likely to lessen. 

Also, China’s reduced exports of fluorspar in response to the 

need for fluorspar by their own industries may be compensated 

for with production from other countries.

The possible extraction of fluorine from alkaline, 

NaHCO3-type evaporite brines is a subject deserving of 

additional research. Extraction of resources from brines has 

proven very cost effective for some mineral commodities, such 

as borax, lithium, potash, and soda ash. Lake Magadi (Kenya) 

and Lake Natron (Tanzania) within the Eastern Rift Valley 

contain large resources of surface and interstitial water with 

fluorine concentrations in excess of 1,000 ppm. Because the 

existing trona mining operation does not recover the villiaumite from the solid trona, it would seem a boon to remove 

fluorine from the brines, possibly allowing two new product 

streams—an evaporating pond product of fluorine-poor trona 

and a solid fluorine product of some type.

Exploration focused on discovery of new, apparently 

conformable fluorspar deposits in tuffaceous limy lacustrine 

sediments also could be fruitful. In the Latium Region of 

Italy, deposits were marginally economic because of difficulty 

recovering the fine-grained fluorite, but other regions may 

have deposits that do not have that problem. The Latium 

deposits are in a minimally eroded young volcanic terrane 

that is specifically ultrapotassic and alkaline. That set of 

parameters can be matched at multiple places in the world; 

for example, Italy’s Pantelleria Island in the Mediterranean 

Sea between Sicily and Tunisia. The deposits in Italy could be 

studied further, possibly to include exploration of the nearby 

maar lakes to see if mineralization processes are ongoing in 

their bottom sediments.

Three types of research in recent years on the composition of fluid inclusions in hydrothermal ore deposits have 

led to greatly increased understanding of the sources of the 

hydrothermal fluids. These are quadrupole mass spectrometry 

(Landis and Hofstra, 1991), noble gas mass spectrometry 

on irradiated fluid inclusions (Kendrick and others, 2002), 

and study of chlorine isotopes in inclusion fluids (Partey 

and others, 2009). Each of these methods has succeeded in 

defining magmatic contributions to hydrothermal fluids where 

magmatic involvement had previously been in question. These 

methods are especially effective when combined with determination of cation concentrations and halogen ratios (Na+

:Br –

and Cl–:Br –) in the same hydrothermal fluids. The ratios can 

help define any component of redissolved halite within the 

fluids. The combined methods on elemental composition of 

solutes in the hydrothermal fluid can separate the fluids of saltrelated carbonate-hosted systems from those of other systems 

with sedimentary brines as ore fluids (Sánchez and others, 

2009, for example). These modern methods are likely to prove 

valuable when extended to additional deposits and districts. 

For deposits whose relation to igneous rocks is equivocal, such 

studies would provide critical new genetic knowledge. The 

fluid inclusion studies are even more effective when combined 

with modern isotopic age determination (Chesley and others, 

1994; Sánchez and others, 2006; Munoz and others, 2005). 

These multifaceted methods have made clear the processes 

and tectonic setting that produced fluorspar and base-metal 

deposits in Western Europe of the class that here are called 

salt-related carbonate-hosted deposits. A promising place to 

apply this combination of methods is Las Cuevas in San Luis 

Potosi State, Mexico.

Acknowledgments

Our USGS colleagues John DeYoung, Steve Ludington,

Klaus Schulz, and Hendrik van Oss provided astute reviews

that improved both the presentation and content of this 

chapter. Dr. Brian Hodge, well known worldwide consultant 

on fluorspar deposits, provided an early critique of our classification of fluorspar deposit types by tectonic or magmatic 

environment (fig. G3B), but this should not be taken to 

indicate that he supports that classification. Dr. Hodge also 

provided many references to published tonnages and grades of 

fluorspar deposits, some of which appear in the compilation 

in appendix G1. With that information, he also provided 

congenial discussion of grades, tonnages, and information 

sources. As with figure G3, however, we the authors, not 

Dr. Hodge, are responsible for the tabulation of appendix G1, 

including any poorly discriminating choices of statistics that 

appear there. Keith Labay helped create the deposit location 

maps, and Suzanne Nicholson provided additional editorial 

comments after the technical review was complete.
G38 Critical Mineral Resources of the United States—Fluorine

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Appendix G1. Selected Fluorspar Districts, Deposits, and

Prospects of the World

The tabulations contained in this appendix do not include 

all world fluorspar deposits and districts, nor even all the 

largest of them. They instead include characteristic examples 

of fluorspar deposits and districts scattered throughout the 

world and provide examples of all fluorspar deposit classes 

in as many places as possible. Collectively, the tables include 

most large fluorspar deposits and districts. All the deposits 

that were compiled in examinations of the distributions of 

fluorspar vein deposit grades and tonnages by Orris (1992) 

are included. The grades and tonnages listed here have not 

been confirmed by the U.S. Geological Survey (USGS), 

and the USGS does not endorse or assure any of the grade 

and tonnage measurements. The grades and tonnages of 

deposits are listed as given in the general geologic and mining 

literature, unrevised by any additional knowledge. Some of the 

values given probably reflect only the resources known at a 

particular instant of time, although, where possible, the listed 

values are thought to be the total identified resources, and 

have included past production and unmined resources of the 

deposits or prospects.

The division between large and small deposits was made 

arbitrarily in a way that placed the largest 20 deposits in China 

from the tabulation of Kamitani and others (2007) in the large 

(L) class and all others in China in the small (S) class. The 

boundary between L and S classes by that procedure lies at 

about 1.5 million metric tons of total calcium fluoride (CaF2 ).

Where possible, the deposits were located on imagery 

from Google Earth, and where Google Earth is listed as the 

source of location data, the latitude and longitude provided 

are directly from Google Earth; these latitudes and longitudes 

have not been corrected to any of the accepted standard 

geodetic spheroids, such as World Geodetic System 1984 

(WGS 84). Latitude and longitude from Google Earth, 

in degrees, were rounded to four decimal places, which 

correspond approximately to the nearest 1-second of latitude 

or longitude. Where possible on Google Earth imagery, the 

latitude-longitude location is placed on the headframe of the 

shaft of the working or abandoned underground mine, or in the 

approximate center of the open pit mine. Any other placement 

of the location point is explained in the Comments column. 

Negative values for latitude indicate that the deposit is in the 

Southern Hemisphere; negative values for longitude indicate 

that the deposit is in the Western Hemisphere.

The table is sorted alphabetically by continent, then 

alphabetically by country, then alphabetically by deposit 

name. The complete reference citations corresponding to 

the reference callouts in the table appear at the end of this 

appendix. The grades and tonnages of selected deposits have 

been plotted in text figure G6; those that are included in that 

graph are listed at the end of this table.

The following is the key to the the identification numbers 

given in the deposit class column in the table:

1. Carbonatite-related

2. Alkaline-intrusion-related veins, skarns, and 

 carbonate replacements

3. Alkaline-volcanic-related epithermal veins, 

breccia-fills, and carbonate replacements

4a. Mississippi Valley-type

4b. Salt-related carbonate-hosted

5. Skarns, greisens, veins, and replacements related

 to strongly differentiated granites

6. Subalkaline-volcanic-related epithermal veins

 and replacements

7. Apparently conformable in tuffaceous, 

 limy lacustrine sediments

8. Deposit class uncertain or not known
G54 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Africa 

Kimwarer Kenya 0.33 35.63 Google Earth (2013), at 

the approximate center 

of the Kimwarer pit, as 

identified by Nyambok 

and Gaciri (1975, 

figs. 2 and 5)

8 Deposit class is 1, 2, or 3; 

resources are given at cutoff 

grade of 20% CaF2

 and 

minimum vein width of 5 m

L 

*El Hammam Morocco 33.56 –5.81 Google Earth (2013), Van 

Alstine and Schruben 

(1980) and Cheilletz and 

others (2010)

5 None L 

*Jebel Tirremi Morocco 34.44 –3.00 Google Earth (2014) 

and Van Alstine and 

Schruben (1980)

2 None S 

*Canxixe area Mozambique –17.58 34.31 Google Earth (2012); 

see comments

8 Location given is for the village

of the same name; no fluorspar

workings identifiable nearby 

on Google Earth

S 

*ChiocoDjanguire 

area

Mozambique –16.40 32.58 Van Alstine and Schruben 

(1980)

8 None S 

*Macossa Mozambique –17.93 34.20 Van Alstine and Schruben 

(1980)

8 None S 

Mount Muambe Mozambique –16.32 34.08 Woolley and Kjarsgaard 

(2008)

1 None S 

*Okorusu Namibia –20.07 16.77 Woolley and Kjarsgaard 

(2008)

1 None L 

*Buffalo South Africa –24.48 28.66 Google Earth (2013);

see comments

5 Location given is centered 

between 3 open pits

L 

*Hlabisa South Africa –28.13 31.90 Van Alstine and Schruben 

(1980)

8 None S 

Vergenoeg South Africa –25.26 28.58 Google Earth (2013) 5 Some interpret that the deposit 

class is 1

L 

Marico district South Africa –25.72 26.09 Google Earth (2013) 4a None L 
Appendix G1 G55

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Africa—Continued

17.1 Mt at 35.4% (indicated); 

and approximately 5 Mt 

at 35.3% (inferred); see 

comments 

Agnerian, 2010 Epithermal/mesothermal vein and 

replacement deposits in Precambrian 

crystalline marble and gneiss; some 

occupy faults of the Rift Valley system; 

veins in Miocene volcanic rocks; 

alkaline volcanic-related epithermal; or 

possibly carbonatite-related

Nyambok and 

Gaciri (1975)

Kimwarer

2.82 Mt at 35% to 45% (as of 

2011); 5 Mt at 57% (as of 

1976)

Managem, Inc. (2011);

Van Alstine and 

Schruben (1980)

Veins Cheilletz and 

others (2010); 

Jébrak and 

others (1988)

*El Hammam

1.5 Mt at 40% U.S. Bureau of Mines 

(1973)

Veins and replacements of limestone 

associated with alkaline intrusive rocks

Van Alstine and

Schruben (1980);

Bouabdellah and 

others (2014)

*Jebel Tirremi

0.16 Mt at >35% Van Alstine and 

Schruben (1980)

Veins n.a. *Canxixe area 

0.698 Mt at >35% Van Alstine and 

Schruben (1980)

Veins n.a. *ChiocoDjanguire

area

0.043 Mt at 35% Van Alstine and

 Schruben (1980)

Veins Van Alstine and 

Schruben (1980)

*Macossa 

1.63 Mt at 19% Globe Metals and

 Mining (2012)

Replacement masses in carbonatite(?) Globe Metals and 

Mining (2012)

Mount Muambe

6 Mt at 56% Notholt and others 

(1990, p. B76)

Replacement masses in carbonatite and 

carbonate wallrock

Kogut and others 

(1997)

*Okorusu

60 Mt at 16% Crocker and Martini 

(1976)

Dense stockwork Absolom (1986) *Buffalo

0.400 Mt at 35% Van Alstine and 

Schruben (1980)

Veins and stockworks Van Alstine and 

Schruben (1980)

*Hlabisa

122 Mt at 35 to 40% (mined 

2.03 Mt at 42.0% CaF2 in 

2005 through 2008)

For tonnage: Minersa 

Group (2012). For 

grade and period of

production: Metorex,

Ltd. (2008, p. 34)

Breccia pipe in Bushveld-related rhyolite 

and tuffaceous sedimentary rocks, 

including ironstones

Goff and others 

(2004)

Vergenoeg

100 to 150 Mt at 15% Martini (1976) Peneconcordant strata-bound replacement 

bodies in dolomites; Witkop Mine, 

Bulhoek Mine, Doornhoek prospect 

Ryan (1986) Marico district
G56 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Africa—Continued 

*Zwartkloof South Africa –24.86 28.17 Google Earth (2013) 5 None S 

*Jebel Semeih 

(J. Sumayh)

Sudan 12.73 30.84 Google Earth (2013); see 

comments 

8 Location given is atop the jebel 

(hill) —artisanal workings 

may be to the southeast along 

the base of the hill

S 

Hammam Zriba Tunisia 36.34 10.20 Google Earth (2014) 4b None L 

Asia 

Badu China 28.03 118.90 Kamitani and others (2007) 6 None L 

Chenlou China 33.77 112.08 Kamitani and others (2007) 8 None L 

Donggangshan China 29.18 112.98 Kamitani and others (2007) 6 None L 

Houshu China 28.92 119.70 Kamitani and others (2007) 6 None L 

Huahe China 31.40 114.57 Kamitani and others (2007) 2 None L 

Huayuan district China 28.59 109.47 Schneider and others 

(2002)

4a Location given is at the major 

intersection in Huayuan city

S 

Hushan China 28.67 118.93 Kamitani and others (2007) 6 None L 

Jianshan China 32.53 113.85 Kamitani and others (2007) 6 None L 

Mahuaping China 27.30 100.08 Kamitani and others (2007) 5 None L 

Maoniuping China 28.46 101.98 Woolley and Kjarsgaard 

(2008)

1 None L 

Shizhuyuan China 25.77 113.18 Google Earth (2014); 

see comments

5 Location given is at an adit, 

but that may be a different 

nearby mine

L 

Shuangjiangkou China 26.95 112.87 Kamitani and others (2007) 8 None L 
Appendix G1 G57

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Africa—Continued

7 Mt at 13.7% Brian Hodge, consultant 

on world fluorspar 

deposits, written 

commun., (2013), 

based on contemporary data published by 

the operating company

Veins Van Alstine and 

Schruben 

(1980)

*Zwartkloof

~0.2 Mt at about 26.5% Van Alstine and 

Schruben (1980)

Vein Van Alstine and 

Schruben (1980)

*Jebel Semeih 

(J. Sumayh)

5 Mt at 15 to 35%, or 6 Mt

at 50%

Bouhlel and others 

(1988) or Van Alstine

and Schruben (1980)

Peneconcordant replacement bodies in 

limestone; dolomitizatioin of wallrocks

Souissi and others 

(1997)

Hammam Zriba

Asia—Continued

5.1 Mt at 50.8%** Kamitani and others 

(2007)

n.a. n.a. Badu

2.5 Mt at 67.25%** Kamitani and others 

(2007)

n.a. n.a. Chenlou

5.97 Mt at 30.49%** (plus 

recoverable Pb and Zn)

Kamitani and others 

(2007)

n.a. n.a. Donggangshan

7.7 Mt at 46.6%** Kamitani and others 

(2007)

Vein n.a. Houshu

3.72 Mt at 60%** Kamitani and others 

(2007)

Veins Wen (1998) Huahe

n.a. n.a. Peneconcordant strata-bound replacement 

bodies in carbonates

Schneider and 

others (2002)

Huayuan district

8.5 Mt at 51.5%** Kamitani and others 

(2007)

n.a. n.a. Hushan

4.8 Mt at about 57.5%** Kamitani and others 

(2007)

Veins Chen and others 

(1996)

Jianshan

10.02 Mt at 22.96%** (plus 

recoverable W and Bi) 

Kamitani and others 

(2007)

Skarn and greisen veinlet stockwork Kamitani and 

others (2007)

Mahuaping

n.a. n.a. n.a. n.a. Maoniuping

46 Mt at 21.7% (plus W, Sn, 

Bi, and Mo)

Brian Hodge, consultant

on world fluorspar 

deposits (written 

commun., 2013); that 

grade and tonnage 

given also by Somerley 

Ltd. (2011, p. 35)

Large skarns, greisens, and stockworks in 

limestone adjacent to biotite granite

Mao and others 

(1996)

Shizhuyuan

4.03 Mt at 60.26%** Kamitani and others 

(2007)

n.a n.a. Shuangjiangkou
G58 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Asia—Continued 

Sumochagan 

Obo

China 43.12 111.27 Google Earth (2012) 6 Deposit class instead may be 7,

and the deposit geologically 

older and grain-coarsened 

than other 7s. Deposit class 

may be 5; grades to skarn in 

the same beds to the south

L 

Taolin China 29.37 113.45 Roedder and Howard 

(1988)

5 None S 

Yangjia-Wuyi China 28.99 119.89 Google Earth (2012); 

see comments

6 Location given is at Yangjia in 

an open cut along a vein

L 

Yinzishan China 30.58 119.87 Kamitani and others (2007) 6 Deposit class instead 

may be 5

L 

Zhuancum China 30.53 118.77 Kamitani and others (2007) 8 None L 

Amba Dongar India 22.00 74.06 Google Earth (2014) 1 None L 

Karadzhal Kazakhstan 49.87 78.04 Google Earth (2016); 

probable location; 

see comments 

5 Location given at left is 

170 km southwest of Semey, 

Kazakhstan, as given by Ulba 

Metallurgical Plant Joint 

Stock Co. (2015). Location 

given at left “lies on the 

northern exo-contact of the 

granite massif [of the Degelen 

Mountains]” on geology map 

of (Böttger and others, 1998, 

p. 127, p. 203). The location 

is marked as “Karadzhal” on a 

map in Vakulchuk and Gjerde 

(2014, p. 16). That location is 

interpreted as a large excavation on a low resolution image 

presented by Google Earth

L?

(resource 

at right is 

considered unlikely to 

be a total 

resource)

Taskainar 

(Taskaynar)

Kazakhstan 43.13 75.14 Seltmann and others (2012) 5 None L 

Taskainar 

(Taskaynar) 

South

Kazakhstan 43.15 75.10 Seltmann and others (2012) 8 None L 

Adag Mongolia 46.50 109.33 Kamitani and others (2007) 3 None L 

Berkh Mongolia 47.77 111.17 Google Earth (2013) 3 None S 
Appendix G1 G59

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Asia—Continued

20 Mt at 56.52% Kamitani and others 

(2007)

Peneconcordant strata-bound 

replacement(?) body in tuffaceous 

limestone, within section mixed with 

high-K calcalkaline volcanics, nearby 

peraluminous granite intrusive

Chen and others 

(1996)

Sumochagan 

Obo

6.1 Mt at 14.3% (plus recoverable Pb, Zn, Cd, Ge, and In) 

Somerly, Ltd. (2011) Veins Roedder and 

Howard (1988)

Taolin

3.38 Mt at about 47.5%** Kamitani and others 

(2007)

Veins Chen and others 

(1996)

Yangjia-Wuyi

5.1 Mt at 38.6%** Kamitani and others 

(2007)

Veins n.a. Yinzishan

3.45 Mt at about 50%** Kamitani and others 

(2007)

n.a. n.a. Zhuancum

12.0 Mt, or 11.6 Mt at 30% For tonnage: Indian Bureau of Mines (2015).

For tonnage and grade:

Palmer (1994)

Hydrothermal veins and replacement 

masses in carbonatite and potassium 

feldspar fenite host rocks

Palmer and 

Williams-Jones 

(1996)

Amba Dongar

1.55 Mt at 40.5% (calculated), 

or given grade at 25%

Tonnage and grade 

calculated from 

Interfax-Kazakhstan 

(2006). Given grade: 

Powell and Throop 

(2011)

Beryllium-bearing, fluorite-magnetitegarnet-skarnAleksandrov, 

(1998, table 8, 

p. 80). 

Karadzhal

>15.625 Mt at 32% Calculated from Seltmann

and others (2012) and 

Koplus and others (1977)

Mineralized breccia at limestone/siliciclastics contact associated with “porphyrite”

dikes or with granodiorite sill

Koplus and others 

(1977)

Taskainar 

(Taskaynar)

>5 Mt fluorite Seltmann and others 

(2012)

n.a. n.a. Taskainar 

(Taskaynar) 

South

4.6 Mt at about 33%** Kamitani and others 

(2007)

Adularia-sericite epithermal vein associated

with alkaline basalt-trachyte-rhyolite series

Lkhamsuren and

Hamasaki (1998)

Adag

1.6 Mt at 81.8%** Kamitani and others 

(2007)

Adularia-sericite epithermal veins 

associated with alkaline basalt-trachyterhyolite series

Lkhamsuren and 

Hamasaki 

(1998)

Berkh
G60 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Asia—Continued 

Bor Ondor Mongolia 46.27 109.44 Google Earth (2013) 3 None L 

MushugaiKhudag

Mongolia 44.37 104.16 Google Earth (2013) 1 None S 

Urgun (Örgön) Mongolia 44.71 110.74 Google Earth (2013) 3 None L? 

Zuum-TsagaanDel

Mongolia 46.37 110.04 Google Earth (2013) 3 None S 

Kalat district Pakistan 29.44 67.15 Durrani (1980) and 

Google Earth (2014)

4a None S? 

Auninsky Russia 54.90 113.12 Ocean Policy Research 

Foundation (2012)

2 None L 

Chailag-Khem Russia 51.91 92.40 Bolonin and others (2009) 1 None L 

Kalangui Russia 51.01 116.53 Google Earth (2013) 6 None L 

Karasug Russia 51.31 92.13 Transcribed from the 

map of Nikiforov and 

others (2005)

1 None L 

Naransky Russia 50.92 105.62 Ocean Policy Research 

Foundation (2012)

2 None L 
Appendix G1 G61

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Asia—Continued

8.4 Mt at 32.25%** Kamitani and others 

(2007)

Adularia-sericite epithermal veins 

associated with alkaline basalt-trachyterhyolite series

Lkhamsuren and 

Hamasaki 

(1998)

Bor Ondor

n.a. n.a. Apatite-magnetite-fluorite intrusive rock 

and later quartz-celestite-fluorite-bariteankerite-phlogopite-bastnaesite veins in 

alkaline ultramafics

Lkhamsuren and 

Hamasaki 

(1998)

MushugaiKhudag

?18.6 Mt at 61.3% (notation

is unclear in reference)

Kamitani and others 

(2007)

Adularia-sericite epithermal vein 

associated with alkaline basalt-trachyterhyolite series

Lkhamsuren and 

Hamasaki 

(1998)

Urgun (Örgön)

3.4 Mt at about 34%** Kamitani and others 

(2007)

Adularia-sericite epithermal veins 

associated with alkaline basalt-trachyterhyolite series

Lkhamsuren and 

Hamasaki 

(1998)

Zuum-TsagaanDel

n.a. n.a. Veins and peneconcordant replacement 

bodies in carbonates; Dilband and at 

least 5 other mineralized areas

Durrani (1980) Kalat district

n.a. n.a. (?)Stockworks and skarns of fluorite-bertrandite

or phenakite veinlets associated with 

aegerine- or reibeckite-granites

By spatial 

association with 

Yermakovskoye

Auninsky

See Karasug; Chailag-Khem is 

part of the same carbonatite 

“province” as Karasug 

Nikiforov and others 

(2005)

See Karasug Nikiforov and 

others (2005)

Chailag-Khem

10.5 Mt at 60% Rodionov and others 

(2010)

Epithermal veins Rodionov and 

others (2010)

Kalangui

At least tens of Mt of about 

10% CaF2; tonnage and 

grade not better known. 

Mean compositions of 

18 weathered-zone samples 

of siderite carbonatite 

(CaF2

-Fe-barite-Sr ore) 

from Karasug field:

9.73% CaF2

;

10.26% siderite;

18.15% hematite;

19.84% goethite;

15.74% barite;

11.47% baritocelestite;

8.92% quartz and other 

fragments; 4.47% calcite; 

0.98% REE minerals;

0.43% apatite

Nikiforov and others 

(2005)

Breccias from plug-like to dike-like, 

to partial ring dikes; ?intrusive- or 

?hydrothermal-matrix-breccia

Nikiforov and 

others (2005)

Karasug

Tonnage not known

at 31%

Rodionov and others 

(2010)

Swarm of parallel veins Rodionov and 

others (2010)

Naransky
G62 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Asia—Continued 

Suran I & 

Suran II

Russia Possibly 

53.84

Possibly 

57.81

Google Earth (2014); 

see comments

8 Deposit class is most likely 

6. Location is uncertain. 

That given by Ocean Policy 

Research Foundation (2012) is 

inconsistent with description 

in Ellmies and others (1999)

L 

Voznesenka and 

Pogranichny 

(the

Yaroslavskoye

or Yoroslavsky 

Mine)

Russia 44.17 132.19 Google Earth (2013); 

center of the 

Voznesenka pit

5 Grade is rounded to the nearest 

5% from the mean of seven 

ore samples reported by 

Kupriyanova and Shpanov 

(1997, p. 385) and four 

ore samples reported by 

Ryazantseva (1998, p. 22). 

Tonnage includes B and 

C1 categories, which are 

approximately equivalent to 

“measured” and “indicated” 

resources, respectively

L 

Yermakovskoye 

(Yermakovska, 

Yermakovsky, 

Ermakovsky, 

Ermakovskoye)

Russia 51.67 109.56 Google Earth (2013) 2 None L 

Takob Tajikistan 38.84 68.94 Google Earth (2012); 

see comments

8 Location given is at the village 

of the same name

S? 

Ban Hong Thailand 18.21 98.82 Google Earth (2015) 8 Deposit class could be 6; tin 

deposits in same area but 

fluorite deposit is epithermal

L 

Doi Tao district Thailand 17.88 98.78 Gardner and Smith (1965) 8 None S? 

Yang Hak 

district

Thailand 13.37 99.47 Gardner and Smith (1965) 8 None S? 

Kizilcaören 

(also known 

as Beylikahir)

Turkey 39.62 31.37 Woolley and Kjarsgaard 

(2008)

Probably

 1

None L 

Suppatash Uzbekistan 41.21 70.53 Seltmann and others (2012) 8 None S 
Appendix G1 G63

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Asia—Continued

n.a. n.a. Veins Ellmies and others 

(1999)

Suran I & 

Suran II

22 Mt at about 40% (Note 

that this tonnage disagrees 

dramatically from that given 

in Nokleberg and others 

(1996). Calculations of 

tonnage using deposit maps 

and sections with reasonable 

assumptions of average ore 

density yield results more 

than an order of magnitude 

smaller than the tonnage 

given by Nokleberg and 

others (1996) [450 Mt])

Tonnage: United 

Company RUSAL 

(2012). Grade: see 

comments 

Fluorite skarn and greisen-veinlet 

stockwork in limestone enveloping 2 

separated greisenized granite cupolas

Kupriyanova 

and Shpanov 

(1997); 

Ryazantseva 

(1998); Sato 

and others 

(2003); and 

Obolenskiy and 

others (2010)

Voznesenka and 

Pogranichny 

(the

Yaroslavskoye

or Yoroslavsky 

Mine)

0.875 Mt at 17.7% (plus 

0.97% BeO) (reserves as 

of 2007 for the mine’s 

proposed re-opening; past 

production not available)

Kislov and others, (2010) Small skarns and stockworks Lykhin and others 

(2010)

Yermakovskoye 

(Yermakovska, 

Yermakovsky, 

Ermakovsky, 

Ermakovskoye)

n.a. n.a. n.a. n.a. Takob

>3.55 Mt at 46.6 % Gardner and Smith 

(1965)

Veins cutting metasedimentary rocks Gardner and 

Smith (1965)

Ban Hong 

n.a. n.a. Veins cutting porphyritic granite Gardner and 

Smith (1965)

Doi Tao district

n.a. n.a. Veins cutting granite, phyllite,

and quartzite

Gardner and 

Smith (1965)

Yang Hak district

12.4 Mt at 37.4% (plus barite, 

REE [ƩREE=3%], and 

Th [0.2%])

Engin (1988) Veins, breccia-fills, and replacements in 

alkalic (aegerine-bearing) tuffs near 

phonolite plugs. REE minerals, including 

bastnaesite, brockite, florencite, and 

monazite, suggest a relation to carbonatite

Kirikoglu (2002) Kizilcaören (also 

known as 

Beylikahir)

5.147 Mt at 27.5% ITE-Uzbekistan (2012?) Epithermal vein in volcanic host rocks State Committee 

of the Republic 

of Uzbekistan 

for Geology 

and Mineral 

Resources 

(2011?)

Suppatash
G64 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Asia—Continued 

Nui Phao Vietnam 21.65 105.67 Richards and others (2003) 

and Google Earth (2014)

5 Other commodity grades for

the 52.5 Mt: W, 0.21%;

Bi, 0.10%; Cu, 0.21%;

Au, 0.22 g/t 

L 

Australia 

Moina Australia –41.49 146.07 Google Earth (2013); 

see comments

5 Location given is for the village 

of the same name

L 

Speewah Australia –16.41 127.98 Gwalani and others (2010) 1 None L 

Europe 

*Argentolle France Possibly 

46.97

Possibly 

4.08

Google Earth (2013); 

see comments

8 Location uncertain. The given 

location is at an industrial 

site between the villages of 

Les Chaux and Crot Morin, 

Burgundy, France. The two 

villages are 1.74 km distant 

from one another.

S 

*Escaro France 42.54 2.31 Google Earth (2013) 8 Either class 4b or 2 S 

Montroc France 43.82 2.38 Google Earth (2013) 8 Either class 4b or 2 S 

*Le Barlet France 45.05 3.50 Google Earth (2013) 8 Either class 4b or 2 S 

Le Burc France 43.87 2.41 Google Earth (2013); 

see comments

8 Either class 4b or 2; 

location uncertain

S 

Rossignol France 46.42 1.31 Google Earth (2013) 8 Either class 4b or 2 S 

*Pakozd Hungary 47.22 18.51 Google Earth (2013) 5 Location is among a group

of exploration trenches

S 

Bruncu 

Molentinu

Italy 

(Sardinia)

39.44 9.48 Google Earth (2013) 8 None L(?) 

*MuscadroxiuGennaTres 

MontisS’Acqua Frida 

(Silius Mine)

Italy 

(Sardinia)

39.52 9.26 Google Earth (2013) 8 Either class 4b or class 5 L 

Castel GiulianoPianciano

Italy 42.07 12.14 Google Earth (2013); 

see comments 

7 Location given is at outcrop 

#5 of Mastrangelo (1976, 

p. 44–45), fairly central to

the fluorspar body

L 

Santo Maria

di Sala 

Italy 42.58 11.75 Google Earth (2013); 

see comments 

7 Location given is at outcrop #1 

of Mastrangelo (1976, p. 41), 

towards the north-west of the 

fluorspar body

L 
Appendix G1 G65

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Asia—Continued

52.5 Mt at 8.0% (plus 

recoverable W, Cu, Au, and 

Bi) (proven and probable)

Masan Group (2012) Skarn and greisen veinlet stockwork Richards and 

others (2003)

Nui Phao

Australia—Continued

24.6 Mt at 15.7% 

(plus Sn, W, Fe)

Bucci and others (2012) Skarn Kwak and Askins 

(1981)

Moina 

6.7 Mt at 24.6% Gwalani and others 

(2010)

Vein/dike of “replacement fluorite rock” 

and carbonatite cutting dolerite

Alvin and others 

(2004)

Speewah

Europe—Continued

0.313 Mt at 48% Bouladon (1989) Vein cutting Mississippian volcanics and 

fine-grained granite basement rocks

Marchand and 

others (1976)

*Argentolle

1.5 Mt at 50% Clarke (1980) Peneconcordant strata-bound replacement 

body in carbonates

Bouladon (1989) *Escaro

1.3 Mt at 50% Deloule (1982) Vein Deloule (1982) Montroc

1.2 Mt at 45% Bouladon (1989) Vein Bouladon (989) *Le Barlet

0.48 Mt at 75% Bouladon (1989) Vein Bouladon (989) Le Burc

0.8 Mt at about 55% Sizaret and others (2004) Vein “venting to distal” stratiform barite 

body with minor amounts of fluorite

Sizaret and others 

(2004)

Rossignol

0.0187 Mt at 46% Morvai (1982) Veins in schist near contact with 

peraluminous granite

Morvai (1982) *Pakozd

n.a. n.a. Vein Cortecci and 

others (1987)

Bruncu 

Molentinu

8.0 Mt at ~42.5% Fulton and Montgomery 

(1994)

Veins Boni and others 

(2009)

*MuscadroxiuGennaTres 

MontisS’Acqua Frida 

(Silius Mine)

10 to 15 Mt at 33% to 50% 

(plus barite)

Matteucci (1976) Stratiform body in tuffaceous lacustrine 

silt/mudstone/marl within peralkaline 

volcanic pile

Mastrangelo 

(1976)

Castel GiulianoPianciano

40 to 50 Mt at about 12%

(plus barite)

Matteucci (1976) Stratiform body in tuffaceous lacustrine 

silt/mudstone/marl within peralkaline 

volcanic pile

Mastrangelo 

(1976)

Santo Maria 

di Sala 
Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

Europe—Continued 

Asturias region Spain 43.43 –5.62 Google Earth (2012); 

see comments 

4a Location given is at the minedout La Collada Mine in 

La Collada (middle) sector 

of the mining region

L 

*Osor Spain 41.95 2.59 Google Earth (2013) 8 Deposit class is most likely 4b S 

Bétic Cordillera 

region: 9 

small-to 

medium-sized 

districts in 

Almeria and 

Granada 

Provinces, 

Andalusia 

region

Spain 36.89 –2.63 Location as given by 

Hellerman (2015)

8 Deposit class is most likely 4b. 

Location given is for the Mina 

Del Carmen located 3.2 km 

northeast of Felix, Almeria, 

Andalusia, Spain, in the 

Sierra de Gádor

L? 

Northern 

Pennine 

Orefield 

United 

Kingdom

54.79 –2.17 Google Earth (2013); 

see comments 

4a Location given is for Frazer’s 

Hush Dib (dib = an inclined 

shaft) of the Frazer’s Grove 

mine in the northeastern part 

of the Orefield

L 

Southern 

Pennine 

Orefield 

United 

Kingdom

53.27 –1.69 Google Earth (2012); 

see comments 

4a Location given is for Stony 

Middleton Mill about 1.7 km 

north of Longstone Edge 

vein system

L 

North America 

*Fission 

(Richardson) 

Mine

Canada 45.06 –78.19 ***MRDS (U.S. 

Geological Survey, 

2012), ID no. I005021

1 None S 

*St. Lawrence 

district 

Canada 46.92 –55.41 Google Earth (2013); 

see comments

2 Location given is for the 

Blue Beach North vein

L 

G66 Critical Mineral Resources of the United States—Fluorine
Appendix G1 G67

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

Europe—Continued

15 Mt at about 38.5% Sánchez and others 

(2009) and Garcia 

Iglesias and Loredo 

(1994)

(1) Veins; (2) Peneconcordant strata-bound 

replacement bodies in calcite-cemented 

sandstone capped by marl-clay; and 

(3) Peneconcordant strata-bound bodies

in “red breccia/conglomerate” immediately overlying basement of limestone 

and slate, also capped by marl-clay

Garcia Iglesias 

and Loredo 

(1994)

Asturias region 

2 Mt at 45% Lipperheide Wicke and 

Barrenechea Guimon 

(1983)

Vein Piqué and others 

(2008)

*Osor

Single deposit; for example, 

Mina Lújar, Granada, as of 

1987: 2.275 Mt proven and 

probable at 34% (plus 

1.85% Pb)

Marina and Vazquez 

Guzmán (1987)

Veins and peneconcordant strata-bound 

replacement bodies in carbonates

Martin and others 

(1987)

Bétic Cordillera 

region: 9 

small-to 

medium-sized 

districts in 

Almeria and 

Granada 

Provinces, 

Andalusia 

region

n.a. n.a. Veins and lesser peneconcordant 

strata-bound replacement bodies

in carbonates

Dunham (1983) Northern Pennine 

Orefield 

Estimated, for example, 

*Sallet Hole No. 1 Adit 

deposit on Longstone Edge: 

2.5 Mt at 40% to 50%; This 

grade and tonnage have not 

been used in figure G6

Calculated tonnage 

estimate uses the 

predicted rate of 

production for 1990 

from Bramley (1990) 

projected for the full 

mine life—1965 to 

1999. This method 

of estimation was 

suggested by 

Dr. Brian Hodge

Veins and lesser peneconcordant 

strata-bound replacement bodies

in carbonates

Dunham (1983); 

Mason (1974)

Southern Pennine 

Orefield 

North America—Continued

0.27 Mt at 26.7% (plus U) Canada Department of 

Energy, Mines, and 

Resources (1984)

Calcite-fluorite-apatite vein/dikes in the 

Bancroft carbonatite province

***MRDS

(U.S. Geological 

Survey, 2012), 

record no. 

I005021

*Fission 

(Richardson) 

Mine

10.04 Mt at 41.%* Agnerian and others 

(2011)

Large veins cutting aegerine granite Van Alstine (1944)

and Strong and 

others (1984)

*St. Lawrence 

district 
G68 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

North America—Continued 

Aguachile Mexico 29.26 –102.54 Google Earth (2013) and 

McAnulty and others 

(1963)

3 None L 

El Refugio Mine Mexico 21.57 –100.18 Google Earth (2015) 6 El Realito Mine with similar 

geology located just 6 km 

northwest

L 

EncantadaBuenavista

Mexico 28.54 –102.50 Google Earth (2012); 

see comments

4a(?) Location given is for Las Sabinas

Mine near southeastern 

extreme of district, which 

stretches more than 30 km 

northwestward from there

L 

Las Cuevas Mexico 21.94 –100.58 Google Earth (2013) 6 None L 

Parral district Mexico 26.85 –105.86 Google Earth (2014); 

see comments

8 Location given is for the

Frisco Mine

L 

La Azul deposit, 

Taxco district

Mexico 18.59 –99.54 Google Earth (2014); 

see comments

8 Location in the open pit S? 

Brown’s Canyon 

district 

United States 38.63 –106.07 Google Earth (2015); 

see comments

8 Deposit class is most likely 6; 

location is for the mouth of 

the Colorado-American adit

S 

Crystal 

Mountain

United States 46.01 –113.89 Google Earth (2013) 8 Deposit class is most likely 5,

or possibly 1

S 

Daisy Mine United States 36.89 –116.68 Google Earth (2013)

and ***MRDS 

(U.S. Geological Survey, 

2013), ID no. M241867

3 None S 
Appendix G1 G69

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

North America—Continued

For example, >0.6 Mt at about 

70% (produced from 1966 

to 1972) (plus Be, 0.3%)

Pickard (1974) 250 m-high, 50 m-wide replacement bodies

in carbonates along their contact with 

peralkaline rhyolite-porphyry ring dike 

McAnulty and 

others (1963) 

Aguachile

about 2.6 Mt at 85% Pickard (1974) Breccia pipe (chimney) replaced by 

fluorite at contacts between limestone 

and intrusive rhyolite

Pickard (1974) 

and Fraga 

(1991)

El Refugio Mine

For example, Las Sabinas 

concession: 13 Mt at 60%

Mexichem (2012) and 

Hodge (2012)

Peneconcordant strata-bound replacement 

bodies (mantos) in carbonates

Gonzalez-Partida 

and others 

(2003)

EncantadaBuenavista

45 Mt at 84% Tonnage: Mexichem 

(2012); grade: 

Hodge (2012)

Breccia body along the faulted contact 

between Jurassic limestone and 

Tertiary rhyolite

Ruiz and others 

(1980)

Las Cuevas

n.a., plus Ag, Pb, and Zn Pickard (1974) Quartz-sulfide veins with calc-silicate 

wallrock alteration cutting shale and 

andesite and some localized along 

intrusive rhyolite dikes; fluorite earlier 

discarded as gangue and formed a 40 Mt 

tailings resource at 15% to 20% CaF2

Pickard (1974); 

Grant and Ruiz 

(1988)

Parral district

0.2 to 0.5 Mt at 63% Pickard (1974) Manto—Peneconcordant replacement 

body in limestone associated with 

fault contact between host Cretaceous 

limestone and Eocene-Oligocene 

felsic volcanics

Tritlla and others 

(2007)

La Azul deposit, 

Taxco district

n.a. n.a. Veins Van Alstine 

(1969)

Brown’s Canyon 

district 

Unknown tonnage at 97.2% 

(plus trace U, Th, Nb, 

and REE)

Sahinen (1962) Massive, sill-like erosional remnants lying 

relatively flat atop (Idaho batholith) 

biotite granite; one body enveloped 

in coarse quartz; contains thortveitite 

(monoclinic [Sc,Y]Si2

O7), a mineral 

elsewhere found in pegmatites, in one 

tin skarn, in a U-Be-Zr-mineralized 

breccia related to granite intrusion, 

and at Franklin, New Jersey, as well as 

fergusonite (tetragonal [Ce, Nd, La, Y]

[Nb, Ti]O4

), a mineral elsewhere found 

with carbonatites or pegmatites

Sahinen (1962) 

and Parker and 

Havens (1963)

Crystal 

Mountain

n.a. n.a. Fault-controlled replacement body in 

limestone

Cornwall and 

Kleinhampl 

(1961)

Daisy Mine
G70 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

North America—Continued 

*Great Eagle United States 32.73 –108.68 Google Earth (2013)

and ***MRDS 

(U.S. Geological Survey, 

2013), ID no. D009524

3 None S 

Hicks Dome United States 37.52 –88.35 Trace and Amos (1984) 1 Igneous carbonatite has not 

yet been identified at and 

published for Hicks Dome, 

but mineralogy and elemental 

signatures suggest this class

L 

*Huckleberry 

Mine

United States 33.32 –108.82 Google Earth (2013)

and ***MRDS

(U.S. Geological Survey, 

2013), ID no. US35038

3 None S 

IllinoisKentucky 

fluorspar 

district 

United States 37.30 –88.24 Trace and Amos, 1984;

see comments

4a Location point is approximately 

at the district centroid; 

the centroid is located 

approximately 1 mile north

of the Eagle-Babb Mine, 

north of Salem, Kentucky

L 

Indian Peak 

district 

United States 38.22 –113.83 Google Earth (2013),

and ***MRDS 

(U.S. Geological Survey, 

2013), ID no. W006953

3 Location point is at the 

Cougar Spar Mine

S 

Jamestown 

district 

United States 40.13 –105.40 Google Earth (2013) and 

Goddard (1946)

2 Location point is at the 

Burlington Mine

S? 

Lost River United States 65.48 –167.16 MRDS (U.S. Geological 

Survey, 2013), ID 

no. A016009

5 Location point is at the 

“New No. 3 Adit” of Plate 2 

of Sainsbury (1964)

L 

Meyers Cove United States 44.85 –114.50 Google Earth (2013) and 

Cox (1954)

8 Location point is at Big Lead 

Mine. Deposit class is most 

likely 6

S 

Northgate 

district 

United States 40.92 –106.27 Google Earth (2013) and 

Steven (1960)

8 Location point is at Pender 

Mine. Deposit class is most 

likely 3

L 

Rome 

occurrence

United States 42.81 –117.72 Google Earth (2013) and 

Sheppard and Gude (1969)

7 None S 

Spor Mountain 

district (Thomas Range)

United States 39.70 –113.17 Google Earth (2013) and 

Staatz and Osterwald 

(1959)

3 Location point is at the 

Ben Hill Mine, pit 1

S 

Sweetwater 

district 

United States 35.60 –84.46 Google Earth (2013); 

see comments 

4a No actual mine or exploration 

workings. Location given is for 

the center of the town square 

of Sweetwater, Tennessee

L 
Appendix G1 G71

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

North America—Continued

0.191 at 43.1% McAnulty (1978) Vein Gillerman (1964) *Great Eagle 

>>11.3 Mt at 11.5% (plus 

potentially recoverable 

REE, Nb, Y, Be, barite, Zn, 

and Pb)

Kirkemo (1978) (tens of 

additional exploratory 

holes have been drilled 

since 1978)

Variably hydrothermally mineralized 

breccia body(s). The dome, though 

spatially northwestern in the greater 

Illinois-Kentucky fluorspar district 

(below), is that district’s thermal center

Taylor and others 

(1992); Joseph 

Porter (oral 

commun., 

1990)

Hicks Dome

0.085 Mt at 35% McAnulty (1978) Vein McAnulty (1978) *Huckleberry 

Mine

30 Mt at 30%; an example 

single vein deposit is the 

Klondike II vein: 1.6 Mt

at 60% 

Grade and tonnage: 

Kogut and others 

(1997). Klondike II 

grade and tonnage: 

Feytis (2009)

Veins and peneconcordant strata-bound 

replacement bodies in carbonates 

(“bedding replacement deposits”). 

Hicks Dome (above), which is 

considered probably related to 

carbonatite, is a thermal center for 

this areally extensive district)

Shawe (1976) IllinoisKentucky 

fluorspar 

district 

n.a. n.a. n.a. n.a. Indian Peak 

district 

>0.7 Mt at 60 to 85% (plus

Pb, Ag, and Au)

Goddard (1946) Veins and breccia pipes, lenticular in 

plan and nearly vertical

Goddard (1946) Jamestown 

district 

23.5 Mt at 16.4% (plus 

recoverable Sn and W)

Watts, Griffiths, and 

McQuat, Ltd. (1972), 

as referenced by Hudson

and Reed, 1997

Skarn and stockwork of exogreisen 

veinlets

Dobson (1982); 

Hudson and 

Reed (1997)

Lost River

n.a. n.a. Veins Cox (1954) Meyers Cove

n.a. n.a. Veins Steven (1960) Northgate 

district 

Typically 0.84 m thickness at 

3.98% 

Sheppard and Gude 

(1969)

Conformable layer in tuffaceous lacustrine 

silt/mudstone within basalts

Sheppard and 

Gude (1969)

Rome 

occurrence

0.033 at 45% (plus Be, U) Bullock (1976) Breccia pipes Staatz and 

Osterwald 

(1959)

Spor Mountain 

district 

n.a. n.a. Breccia bodies involving several 

tens of meters of stratigraphy and 

peneconcordant replacement bodies 

in carbonates

Misra and others 

(1989)

Sweetwater 

district 
G72 Critical Mineral Resources of the United States—Fluorine

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; 

W, tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Deposit

name1 Country

Latitude Longitude Source of

location data

Deposit 

class Comments Deposit size 

(decimal degrees)

North America—Continued 

*White Eagle United States 32.61 –107.75 Google Earth (2013),

and ***MRDS

(U.S. Geological Survey, 

2013), ID no. W006929

3 None S 

South America 

Cerros Negros Argentina –32.50 –64.79 Google Earth (2014) 

and Coniglio and 

others (2000)

6 None S 

Mato Preto Brazil –24.75 –49.18 Google Earth (2014) 

and Jenkins (1987); 

see comments

1 Location given is for the 

approximate center of the 

Clugger orebody

L 

Santa Catarina 

district 

Brazil –26.62 –49.26 Google Earth (2014), 

and Bastos Neto and 

others (1991)

8 Deposit class is most likely 2, 

but might be 4b. Location 

given is at Shaft 3 on the 

Segunda Linha Torrens vein

L 

1

Grade and tonnage information from the following deposits have been plotted in text figure G6.

Fluorspar vein or vein-and-replacement deposits

1. El Hammam, Morocco

2. Canxixe, Mozambique

3. Chioco-Djanguire, Mozambique

4. Macossa, Mozambique

5. Zwartkloof, South Africa

6. Jebel Semeih, Sudan

7. Houshu, China

8. Huahe, China

9. Hushan, China

10. Jianshan, China

11. Yangjia-Wuyi, China

12. Yinzishan, China

13. Adag, Mongolia

14. Berkh, Mongolia

15. Bor Ondor, Mongolia

16. Zuum-Tsagaan-Del, Mongolia

17. Kalangui, Russia

18. Ban Hong, Thailand

19. Suppatash, Uzbekistan

20. Speewah, Australia

21. Argentolle, France

22. Montroc, France

23. Le Barlet, France

24. Le Burc, France

25. Rossignol, France

26. Muscadroxiu-Genna Tres MontisS’Acqua Frida, Sardinia, Italy

27. Osor, Spain

28. Fission (Richardson), Canada

29. Blue Beach North vein,

St. Lawrence district, Canada

30. El Refugio Mine, Mexico

31. Las Cuevas Mine, Mexico

32. Great Eagle, New Mexico,

United States

33. Huckleberry Mine, New Mexico,

United States

34. Klondike II deposit, Illinois-Kentucky

fluorspar district, United States

35. White Eagle, New Mexico,

United States

36. Cerros Negros, Argentina

Carbonatite-related fluorspar deposits

1. Mount Muambe, Mozambique

2. Okorusu, Namibia

3. Amba Dongar, India

4. Speewah, Australia

5. Mato Preto, Brazil

Skarn, greisens, and replacement fluorspar 

deposits related to strongly differentiated granites

1. Buffalo, South Africa

2. Vergenoeg, South Africa

3. Mahuaping, China

4. Shizhuyuan, China

5. Taskainar (Taskaynar), Kazakhstan

6. Voznesenka I & Pogranichny, Russia

7. Nui Phao, Vietnam

8. Moina, Tasmania, Australia

9. Lost River, Alaska, United States
Appendix G1 G73

Table G1–1. Selected fluorspar districts, deposits, and prospects of the world.—Continued

[*Deposit is included in the Orris (1992) grade and tonnage modeling of fluorite veins. **Grade and tonnage data are from Kamitani and others (2007); for these 

entries, it is not clear whether these are the total resources of each deposit or just a resource currently available to the mine that may be subject to updating with 

further exploration. ***Location data are from the Mineral Resources Data System (MRDS). The key to the identification numbers in the deposit class column 

is on page G53. Deposit size: L, identified resources are greater than 1.5 million metric tons of CaF2; S, identified resources are less than 1.5 million metric tons 

of CaF2; n.a., not available. Elements and compounds: Ag, silver; Au, gold; BeO, beryllium oxide; Bi, bismuth; CaF2, calcium fluoride;Cd, cadmium; Cu, copper; 

Fe, iron; Ge, germanium; In, indium; Mo, molybdenum; Nb, niobium; Pb, lead; REE, rare-earth element; Sn, tin; Sr, strontium; Th, thorium; U, uranium; W, 

tungsten; Y, yttrium; Zn, zinc. Units of measure: g/t, gram per metric ton; km, kilometer; m, meter; Mt, million metric tons; %, percent]

Published resource 

(tonnage and grade, in weight 

percent CaF2

)

Source of 

resource data

Predominant geometry

of orebody(ies)

Source of geologic descriptionDeposit

name 

North America—Continued

0.179 Mt at 54.3% McAnulty (1978) Vein McAnulty (1978) *White Eagle

South America—Continued

0.27 Mt at 56% Coniglio and others 

(2000)

Vein Coniglio and 

others (2000)

Cerros Negros

2.65 Mt at 60% Jenkins (1987) Lensatic hydrothermal masses replacing 

calcitic carbonatite and lesser syenite, 

enveloping phonolite dikes intruded 

along the contact between carbonatite 

and nepheline syenite

Woolley (1987) 

and Jenkins, 

(1987)

Mato Preto

n.a. n.a. Veins Bastos Neto and 

others (1991)

Santa Catarina 

district 
G74 Critical Mineral Resources of the United States—Fluorine

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for the Second Conference on Humanitarian Impact of Nuclear

Weapons in Nayarit, Mexico, February 13–14, 2014.]

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the Poncha Springs NE quadrangle, Chaffee County, 

Colorado, with a section on Fluorspar mines and prospects, 

by R.E. Van Alstine and D.C. Cox: U.S. Geological 

Survey Professional Paper 626, 52 p., 6 pls. [Also available 

at http://pubs.er.usgs.gov/publication/pp626.]

Van Alstine, R.E., and Schruben, P.G., 1980, Fluorspar 

resources of Africa: U.S. Geological Survey Bulletin 1487, 

25 p., 1 pl. [Also available at http://pubs.er.usgs.gov/

publication/b1487.]

Watts, Griffiths, and McQuat, Ltd., 1972, Preliminary feasibility

report on the Lost River fluorite-tin-tungsten: Toronto, 

Ontario, Canada, Lost River Mining Company, Ltd., 291 p.

Wen Lu, 1998, Fluorspar, in Wen Lu, Chinese industrial 

minerals (Griffiths, Joyce, ed.): Worchester Park, Surrey, 

United Kingdom, Industrial Minerals Information, Ltd., 

p. 68–75.

Woolley, A.R., 1987, Alkaline rocks and carbonatites of the 

world—Part 1—North and South America: London, United 

Kingdom, British Museum of Natural History, 216 p.

Woolley, A.R., and Kjarsgaard, B.A., 2008, Paragenetic types 

of carbonatites as indicated by the diversity and relative 

abundances of associated silicate rocks—Evidence from 

a global database: Canadian Mineralogist, v. 46, no. 4, 

p. 741–752. [Also available at http://dx.doi.org/10.3749/

canmin.46.4.741.]
Gallium

By Nora K. Foley, Brian W. Jaskula, Bryn E. Kimball, and Ruth F. Schulte

Chapter H of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–H

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter H

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................H1

Introduction..................................................................................................................................................H1

Strategic and Critical Resource Issues .........................................................................................H2

Uses and Applications ......................................................................................................................H2

Geology.........................................................................................................................................................H6

Geochemistry......................................................................................................................................H6

Mineralogy........................................................................................................................................H12

Gibbsite-Diaspore-Boehmite ................................................................................................H14

Sphalerite .................................................................................................................................H15

Deposit Types ...................................................................................................................................H15

Bauxite Deposits .....................................................................................................................H15

Sediment-Hosted Lead-Zinc Deposits ................................................................................H18

Clastic-Dominated Lead-Zinc Type .............................................................................H19

Mississippi Valley Lead-Zinc Type ..............................................................................H19

Kipushi Polymetallic Type .............................................................................................H20

Resources and Production ......................................................................................................................H21

Identified Resources .......................................................................................................................H21

Undiscovered Resources ...............................................................................................................H21

Production.........................................................................................................................................H21

Exploration for New Deposits .................................................................................................................H22

Environmental Considerations ................................................................................................................H22

Sources and Fate in the Environment ..........................................................................................H22

Mine Waste Characteristics ..........................................................................................................H23

Human Health Concerns .................................................................................................................H24

Ecological Health Concerns ...........................................................................................................H24

Carbon Footprint...............................................................................................................................H25

Mine Closure.....................................................................................................................................H25

Problems and Future Research ..............................................................................................................H26

References Cited.......................................................................................................................................H26

Suggested citation:

Foley, N.K., Jaskula, B.W., Kimball, B.E., and Schulte, R.F., 2017, Gallium, chap. H of Schulz, K.J., DeYoung, J.H., Jr., 

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. H1– H35, 

http://dx.doi.org/10.3133/pp1802H.
Gallium

By Nora K. Foley, Brian W. Jaskula, Bryn E. Kimball, and Ruth F. Schulte

Abstract

Gallium is a soft, silvery metallic element with an 

atomic number of 31 and the chemical symbol Ga. Gallium is 

used in a wide variety of products that have microelectronic 

components containing either gallium arsenide (GaAs) or 

gallium nitride (GaN). GaAs is able to change electricity 

directly into laser light and is used in the manufacture of 

optoelectronic devices (laser diodes, light-emitting diodes 

[LEDs], photo detectors, and solar cells), which are important 

for aerospace and telecommunications applications and 

industrial and medical equipment. GaAs is also used in 

the production of highly specialized integrated circuits, 

semiconductors, and transistors; these are necessary for defense 

applications and high-performance computers. For example, 

cell phones with advanced personal computer-like functionality 

(smartphones) use GaAs-rich semiconductor components. 

GaN is used principally in the manufacture of LEDs and laser 

diodes, power electronics, and radio-frequency electronics. 

Because GaN power transistors operate at higher voltages and 

with a higher power density than GaAs devices, the uses for 

advanced GaN-based products are expected to increase in the 

future. Gallium technologies also have large power-handling 

capabilities and are used for cable television transmission, 

commercial wireless infrastructure, power electronics, and 

satellites. Gallium is also used for such familiar applications 

as screen backlighting for computer notebooks, flat-screen 

televisions, and desktop computer monitors.

Gallium is dispersed in small amounts in many minerals 

and rocks where it substitutes for elements of similar size and 

charge, such as aluminum and zinc. For example, gallium 

is found in small amounts (about 50 parts per million) in 

such aluminum-bearing minerals as diaspore-boehmite and 

gibbsite, which form bauxite deposits, and in the zinc-sulfide 

mineral sphalerite, which is found in many mineral deposits. 

At the present time, gallium metal is derived mainly as a 

byproduct of the processing of bauxite ore for aluminum; 

lesser amounts of gallium metal are produced from the 

processing of sphalerite ore from three types of deposits 

(sediment-hosted, Mississippi Valley-type, and volcanogenic 

massive sulfide) for zinc. The United States is expected to 

meet its current and expected future needs for gallium through 

imports of primary, recycled, and refined gallium, as well as 

through domestic production of recycled and refined gallium. 

The U.S. Geological Survey estimates that world resources 

of gallium in bauxite exceed 1 billion kilograms, and a 

considerable quantity of gallium could be present in world 

zinc reserves.

Introduction

The existence of gallium was first predicted in 1871 by 

Dmitri Mendeleev, a Russian chemist who published the first 

periodic table of elements (Mendeleev, 1871). Mendeleev 

noted a gap in his table and named the missing element 

“eka-aluminum” because he determined that its location 

was one place away from aluminum in the table (fig. H1). 

Mendeleev thought that the missing element (gallium) would be 

Figure H1

Al

13

26.98

Si

14

28.09

Ge

32

72.64

Ga

31

69.72

Zn

30

65.39

Cd

48

112.4

Sn

50

118.7

In

49

114.8

Figure H1. Photograph of gallium metal with an inset showing 

its position on the periodic table of elements. The gallium 

crystals shown in the photograph grew in a tank of liquid 

gallium. The drop in the foreground fell off the crystal as the 

liquid slowly cooled below a temperature of about 30 degrees 

Celsius. Photograph courtesy of Afshin Yaghmaei, Simulated 

Planetary Interiors Lab, Department of Earth and Space 

Sciences, University of California, Los Angeles.
H2 Critical Mineral Resources of the United States—Gallium

very much like aluminum in its chemical properties, and he was 

right. The French chemist Paul-Emile Lecoq de Boisbaudran 

discovered gallium in sphalerite in 1875 using spectroscopy. 

Lecoq de Boisbaudran isolated gallium metal by electrolysis 

of gallium hydroxide in a potassium hydroxide solution. 

He named the element “gallia” after the Latin name for his 

native land of France (Gallia, or Gaul).

Solid gallium has a low melting temperature (about 

29 degrees Celsius [°C]) and an unusually high boiling 

point (about 2,204 °C) (table H1), so that gallium is a liquid 

across a wider range of temperatures than any other element. 

Because of these properties, the earliest uses of gallium were 

in high-temperature thermometry and in metal alloys that 

melt easily. The development of gallium arsenide as a direct 

band-gap semiconductor in the 1960s led to what are now 

some of the most well-known uses of gallium—in feature-rich, 

application-intensive, third- and fourth-generation smartphones and in data-centric networks.

Strategic and Critical Resource Issues

Gallium first became recognized as a strategic and critical 

metal during World War II because it was a necessary component of a stable gallium-plutonium alloy that was developed 

as part of the Manhattan Project and used to make a reliable 

atomic bomb. Gallium is currently important for U.S. manufacturing because of its expanding uses for novel thin-film 

photovoltaics, especially those of the copper-indium-gallium-

(di)selenide (CIGS) semiconductor type, and it is particularly 

important for clean energy technologies. Gallium is currently 

produced as a byproduct of the mining of other mineral 

commodities—mainly zinc, copper, and aluminum—and the 

mining technologies used are optimized for the extraction of 

those primary mineral commodities. The small market for 

gallium as a specialty metal creates little incentive for refiners 

of zinc, copper, and aluminum ores to invest in improvements 

to increase byproduct recovery of gallium. Because of the 

likelihood of rapid growth in the areas of photovoltaics and 

clean energy technologies, a potential exists for bottlenecks in 

the gallium supply pipeline. Production of gallium is limited 

by market factors that influence the production of the principal 

mineral commodity, whether it is zinc or aluminum.

Uses and Applications

A wide variety of products have microelectronic components that contain either gallium arsenide (GaAs) or gallium 

nitride (GaN). GaAs is able to change electricity directly into 

laser light and is used in the manufacture of optoelectronic 

devices (laser diodes, light-emitting diodes [LEDs], photo 

detectors, and solar cells), which are important for aerospace 

and telecommunications applications and industrial and 

medical equipment. GaAs is also used in the production of 

highly specialized integrated circuits, semiconductors, and 

transistors; these are necessary for defense applications and 

high-performance computers. For example, smartphones (cell 

phones with advanced computer-like functionality) use GaAsrich semiconductor components (fig. H2A). GaN is employed 

principally in the manufacture of LEDs and laser diodes, power 

electronics, and radio-frequency (RF) electronics. Because 

GaN power transistors operate at higher voltages and with 

a higher power density than GaAs devices, the applications 

for advanced GaN-based products are expected to increase 

in the future. The CIGS direct band-gap semiconductor was 

developed to make lightweight, flexible, and durable thin-film 

photovoltaics that have a high absorption coefficient; these 

attributes have potential applications in the manufacture of 

efficient solar cells (fig. H2B). Because of its high boiling 

temperature, gallium is used to make thermometers designed 

to measure very high temperatures, and because of its silvery 

color and ability to form metal alloys, gallium is used to 

make brilliant mirrors. The fundamental properties of thermal 

convection in liquid gallium metal are employed to study 

aspects of planetary and astrophysical magnetohydrodynamics.

Gallium can be replaced by other materials in some 

applications, but for others, it has no effective substitutes. 

Liquid crystals made from organic compounds are used in 

visual displays as substitutes for LEDs. Researchers also are 

working to develop organic-based LEDs that may compete 

with GaAs in the future. Indium phosphide components can 

be substituted for GaAs-based infrared laser diodes in some 

specific-wavelength applications, and helium-neon lasers 

can be used in place of GaAs lasers in visible laser diode 

applications. Silicon is the principal competitor with GaAs 

in solar-cell applications. GaAs-based integrated circuits are 

used in many defense-related applications because of their 

unique properties, and there are no effective substitutes for 

GaAs in these applications. GaAs in heterojunction bipolar 

transistors is being substituted by silicon-germanium in 

some applications.

In 2012, imports of gallium and GaAs wafers, which 

were valued at about $32 million, continued to satisfy almost 

all U.S. demand for gallium. GaAs and GaN electronic 

components represented about 99 percent of domestic gallium 

consumption (fig. H3). About 70 percent of the gallium was 

used in integrated circuits (both analog and digital), including 

those in defense applications, high-performance computers, 

and telecommunications. Optoelectronic devices for use in 

aerospace and telecommunications applications, consumer 

goods, industrial equipment, medical equipment, and so forth, 

as well as a small amount for research and development, 

represented the remaining 30 percent of gallium consumption.

Gallium prices decreased throughout 2012 when 

significant increases in gallium production from 2007 

to 2012 (fig. H4) exceeded the declining demand from 

LED producers. Gallium production capacity in China 
Introduction H3

expanded tremendously from 2008 to 2012 on the expectation 

of a strong LED-based backlighting market, which failed to 

materialize. In January 2012, the price for low-grade gallium 

metal (with a purity of 99.099 percent compared with a purity 

of 99.99999 for the higher grade) averaged $580 per kilogram. 

By July, the average price for low-grade gallium metal had 

decreased to $320 per kilogram, and by early October, to 

$280 per kilogram.

Markets continued to expand for GaAs- and GaN-based 

products in 2012. GaAs demand mainly results from cell 

phones and other high-speed wireless applications. In 2012, 

GaAs demand increased owing to the rapid increase in the 

use of feature-rich, application-intensive, third- and fourthgeneration smartphones, which contain up to 10 times the 

amount of GaAs compared with standard cellular handsets. 

Smartphones accounted for 37 percent of all handsets sold 

in 2012. Owing to the increased amounts of GaAs in smartphones and the increased use of GaAs-based LEDs in general 

lighting and automotive applications, the value of sales in the 

GaAs market is forecast to reach $650 million by 2017 at an 

annual growth rate of nearly 11 percent (Jaskula, 2013b).

Owing to the large power-handling capabilities, highswitching frequencies, and higher voltage capabilities of 

GaN technology, GaN-based products, which historically 

have been used in defense and military applications, have 

begun to gain acceptance in the cable television transmission, 

commercial wireless infrastructure, power electronics, and 

satellite markets. The value of sales for the GaN power device 

market was expected to reach $178 million by 2015 at an 

annual growth rate of nearly 29 percent (Jaskula, 2013b).

In 2012, the worldwide LED market, which was a 

significant driver for GaN-based technologies, increased 

by only 1.5 percent in revenues from that of 2011 owing 

to slower-than-expected growth in the LED backlighting 

market and to lower LED prices, as production was outpacing 

demand. Although televisions were increasingly being built 

with LED backlighting in 2012, improvements in technology 

required up to 50 percent fewer LEDs. LED-backlit televisions 

accounted for 63 percent of the television market in 2012 

and were forecast to account for 93 percent of market sales 

in 2013. By 2014, the strongest market segments for sales of 

LEDs were expected to be in general lighting, followed by 

signs and automotive applications.

Sustained high prices for energy continued to spark 

interest in solar energy in 2012. Sales of CIGS semiconductors 

for use in the production of high-efficiency solar cells was 

slower than expected, however, owing to a complicated 

manufacturing process that has impeded commercial mass 

production of CIGS panels. Decreased prices of silicon-based 

solar cells also affected demand for the more expensive 

CIGS technology. These two factors resulted in a large 

oversupply of CIGS modules that caused prices to decrease 

by 20 percent in 2011 and to remain low throughout 2012. In 

an effort to keep CIGS technology viable and competitive, 

CIGS manufacturers began in 2011 to trim production costs, 

improve module conversion efficiencies, and increase the use 

of CIGS-based photovoltaics in commercial rooftops.

The following sections of this chapter provide an overview of the principal types and global distribution of domestic 

and global resources of gallium (fig. H4). This information 

updates the report by Weeks (1973).

A

B

Gallium Figure 3. Figure H2. Photographs showing examples of some current 

uses for gallium. A, Cell phone that has advanced computer-like 

functionality (that is, a smartphone), which contains galliumbased semiconductor components and, consequently, up to 

10 times the amount of gallium arsenide (GaAs) than is used in 

standard cell phones. B, Experimental three-wheel car displaying 

high-efficiency, triple-junction GaAs solar cells. Such solar cells 

have many experimental applications; this three-wheel car, the 

Continuum, which was built at the University of Michigan, weighs 

220 kilograms and is covered with about 2,500 aerospace-grade 

GaAs-type solar cells. Photograph of experimental car courtesy 

of the University of Michigan.
H4 Critical Mineral Resources of the United States—Gallium

B. Estimated world primary gallium production capacity

A. Estimated world primary gallium production

Gallium Figure 4

C. U.S. gallium consumption U.S. gallium consumption by end use, 2012

Refined production (includes

 primary production until 2011)

Primary production

EXPLANATION

300

400

200

100

0

2007

2007

2008 2009 2010 2011 2012

2008 2009 2010 2011 2012

2007 2008 2009 2010 2011 2012

100

80

60

40

20

0

Analog 

integrated

circuits

60%

LEDs and

laser diodes

30%

Digital 

integrated circuits

8%

Solar cells

2% Research and

development

<0.5%

Year

Primary gallium 

 producers:

China

Germany

Hungary

Japan

Kazakhstan

Republic of Korea

Russia

Ukraine

Refined gallium producers:

 China

 Japan

 United Kingdom

 United States

Gallium content, in metric tons Gallium content, in metric tons Gallium content, in metric tons

300

400

200

100

0

China Other producers

No

estimate

Figure H3. Charts showing U.S. and world gallium production and consumption from 2007 to 2012. A, Estimated world 

primary gallium production and refined gallium production from 2007 to 2012. B, Estimated world primary gallium production 

capacity for China and the rest of the world from 2008 to 2012. C, Gallium consumption in the United States from 2007 to 2012, 

and major end uses of gallium as a percentage of U.S. consumption in 2012. Percentages in the pie chart have been rounded 

and consequently do not add to 100. LEDs, light-emitting diodes
Introduction H5

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Russian province

Chinese

province

Australian province

Indian

province

Southern

European

province

Guinea Shield

province

Guyana Shield

province

Caribbean

province

Brazilian Shield

province

Cameroon

province

Arkansas

province

Apex

Kipushi

Ruby Creek

Red Dog 

Viburnum Trend

Nanisivik

Upper 

Mississippi

Valley

Austinville

AU: The text mentions Mount Isa, and the Cooley and Ridge Mines as being Kipushi-type deposits also, but they are not shown on this map?

Gallium figure 1.

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

EXPLANATION

Gallium sources

Bauxite depositsSediment-hosted  zinc depositsKipushi-type deposits

Not all the Kipushi-type 

osits. The selected bauxite deposits are 

 both karst and lateritic types. The sediment-hosted zinc deposits and districts named in the text (Austinville, 

World map showing locations of selected mineral deposits, by type; gallium has been produced from these types of dep

 (See Taylor and others [2009] for names and details regarding other zinc deposits.) 

s major bauxite provinces and include

rend) are labeled. iburnum T

Figure H4.located within the world’Nanisivik, Red Dog, Ruby Creek, and Vdeposits mentioned in the text appear on this map.
H6 Critical Mineral Resources of the United States—Gallium

Geology

Gallium occurs in low amounts in a diverse group of 

geologic environments and is found most commonly in association with deposits of aluminum and zinc (fig. H5), which 

are its immediate neighbors in the periodic table of elements 

(fig. H1). The average abundance of gallium in Earth’s crust 

is generally less than 19 parts per million (ppm) (table H2). 

Because of the rarity of gallium in most geologic environments, minerals containing gallium as an essential structural 

component are also rare (table H3). Because gallium occurs 

as a minor or trace element in many minerals, there are a large 

number of rock types and tectonic settings for mineral deposits 

that are known to contain gallium in potentially extractable 

amounts (table H4). At the present time, the majority of the 

world’s supply of newly mined gallium metal comes from 

bauxite and sediment-hosted lead-zinc deposits (fig. H4). 

Gallium is currently derived as a byproduct of the processing 

of bauxite ore for aluminum, with lesser amounts produced 

from residues resulting from the processing of sphalerite ore 

for zinc.

Geochemistry

Gallium is a soft, silvery post-transition metal with an 

atomic number of 31 (table H1; fig. H1) that is very much 

like aluminum in its chemical properties. The chemical 

similarities between gallium and aluminum, which include a 

trivalent oxidation state, a similar atomic radius, tetrahedral 

or octahedral coordination in minerals, and amphoterism 

(that is, they can act as both acid and base in reactions), 

result in their generally synchronous geochemical behavior. 

Gallium is found primarily in the trivalent oxidation state; 

mixed-oxidation state compounds containing both Ga1+ and 

Ga3+ (ionic radius of about 0.62 angstrom [Å]) occur but are 

not common (Eagleson, 1994, p. 438). In compounds, gallium 

exhibits a chemical similarity to Al3+ (ionic radius of about 

0.54 Å), Fe3+ (ionic radius of about 0.69 Å) and Zn2+ (ionic 

radius of about 0.72 Å) (Shannon, 1976) and can substitute 

for these elements in the common rock-forming minerals 

(Burton and Culkin, 1978). Other elements typically found 

in geochemical association with gallium include cadmium, 

indium, germanium, silicon, and tin (fig. H1). Under certain 

geochemical conditions, these elements can be separated in 

near-surface environments—for example, by changes in pH 

or redox conditions—because their acid/base and electronic 

properties are sufficiently different (table H1).

The aqueous geochemistry of gallium was studied by 

Wood and Samson (2006, and references therein) based on a 

number of earlier works. Trivalent gallium is also the most 

common oxidation state of gallium metal in aqueous solution 

and, owing to the high charge, the hydrated ions of gallium are 

Pearson hard acids (Pearson, 1963), which favor the formation 

of complexes with hard ligands, such as fluoride, hydroxide, 

phosphate, and sulfate. In aqueous solutions, gallium is 

predicted to occur mainly in fluoride and hydroxide complexes 

(Wood and Samson, 2006). Solubility products of most 

solid phases containing gallium are generally not available; 

however, the limited available solubility data constrain the 

maximum concentrations of gallium to be expected in aqueous 

fluids to the subparts-per-million to subparts-per-billion (or 

micrograms per liter) range (Wood and Samson, 2006), and 

this is evident in the values seen for hydrothermal fluids, 

river waters, and seawater (table H2). The hydrolysis of 

GaOOH(s) (solid gallium oxyhydroxide) occurs at low pH, 

and its speciation is strongly dominated by Ga(OH)4

–

 in the pH 

range of natural solutions at 25 °C. At increasing temperature, 

the amount of Ga(OH)2

+

 and Ga(OH)3° becomes significant, 

although at 300 °C, Ga(OH)4

–

 remains the predominant 

species. A comparison of the geochemical behavior of gallium 

and aluminum in natural fluids indicates that Ga(OH)4

–

exhibits a chemical behavior very similar to that of Al(OH)4

−

(Wood and Samson, 2006). The low solubility of α-GaOOH(s) 

(in the absence of fluoride or strong organic ligands at low 

temperatures) is consistent with the immobility of gallium 

relative to most other elements, except aluminum, during 

weathering and alteration processes. This immobility leads 

to the accumulation of gallium in some bauxites and kaolins, 

and in the high-alumina hydrothermal alteration halos that 

form around many mineral deposits (for example, epithermal, 

porphyry, and volcanogenic deposits).

Figure H5. (page H7) Cross sections showing the general 

geologic environments for the types of mineral deposits with 

which gallium is most commonly associated and from which 

gallium is typically extracted. A, Karst bauxite deposits, which 

form by deep weathering and transport of material in carbonate 

rock-dominated sequences; unlabeled patterns show detritus. 

B, Lateritic bauxite deposits, which form mainly by deep 

weathering of material in silicate rock-dominated sequences. 

A and B modified from Retallack (2010). C, Sediment-hosted 

lead-zinc deposits, which form in passive margin environments. 

D, Clastic-dominated-type lead-zinc deposits, which may form by 

brine reflux in shales and deep-water carbonates. E, Mississippi 

Valley-type lead-zinc deposits, which form by replacement in 

platform carbonate sequences. C, D, and E modified from Kelley 

and others (1995); Bradley and Leach (2003); Leach, Sangster, 

and others (2005); and Leach, Taylor, and others (2010). 
Geology H7

Bedrock

Bauxitic pedolith

Bauxitic pedolith

Bauxitic paleosol

Bauxitic paleosol

Soil

Younger limestone

D. Clastic-dominated-type lead-zinc deposits

Reflux model for passive continental margin 

Open marine environment Evaporative basin

Brine pool Brine discharge Brine generation

Carbonate rocks

Clastic rocks—Massive (solid pattern) and 

 laminated (striped pattern)

Lead-zinc ores in clastic rocks

Carbonate turbidites

Fault 

Brine influx

Dolomitization 

Metal extraction 

C. Sediment-hosted lead-zinc deposits

Clastic-dominated

type host

environment 

Sea level

Mississippi Valleytype host

environment 

Mafic and ultramafic rocks Shale, deep-water carbonates

EXPLANATION (part C) EXPLANATION (part D)

Carbonate and minor clastic rocks of the

 passive margin platform

Rift facies rocks

Basement rocks

B. Lateritic bauxite deposits 

Lateritic paleosol

bauxite

Saprolite

Lateritic

pedolith

Bedrock

Lateritic

cuirasse

Bauxite

pallid zone

Mottled zone

Soil

Surface gradient towards craton

Central 

Missouri

Viburnum

North End

Northern 

Arkansas Tri-State

Filled basins

Precambrian basement rocks

Shale, deep-water carbonates

Graywacke turbidites, molasse facies rocks

EXPLANATION

Carbonate and minor 

 clastic rocks of the

 passive margin platform

Fault Location of ore

Underlying rocks

E. Mississippi Valley-type lead-zinc deposits

ALL FIGURES NOT TO SCALE

Gallium Figure 5

A. Karst bauxite deposits
H8 Critical Mineral Resources of the United States—Gallium

Table H1. Selected properties of gallium, a Group 13 posttransition metal.

[Sources: Instytutu Problem Materialoznavstva (1968) and Lide (2005, 

p. 14–39). Å, angstrom; °C, degree Celsius; cm, centimeter; eV, electronvolt; g/cm3

, gram per cubic centimeter; GPa, gigapascal; J/mol-K, joule per 

mole kelvin; K, kelvin; kJ/mol, kilojoule per mole; nΩ-m, nano ohm-meter; 

μm/m-K, micrometer per meter kelvin; W/m-K, watt per meter kelvin]

Property Description

Symbol Ga

Atomic number 31

Atomic weight 69.723

Isotopes 24 radioisotopes; 69Ga, 71Ga 

are stable isotopes; 67Ga has 

a half-life of about 3.3 days

Density at 293 K (g/cm3

) 5.91

Melting point (°C) 29.7646

Boiling point (°C) 2,204

Heat of fusion (kJ/mol) 5.59

Heat of vaporization (kJ/mol) 254

Molar heat capacity (J/mol-K) 25.86

Hardness (Mohs scale) 1.5

Brinell hardness (GPa) 0.06

Electrical resistivity at 20 °C 

(nΩ-m)

270

Thermal conductivity (W/m-K) 40.6

Thermal expansion at 25 °C 

(μm/m-K)

18

Young’s modulus (GPa) 9.8

Crystal structure Orthorhombic

Magnetic ordering Diamagnetic

Electron configuration [Ar]4s2 3d10 4p1

Ionic radius (Å) 0.62

First ionization potential (eV) 5.999

Second ionization potential (eV) 20.51

Third ionization potential (eV) 30.71

Common valance states Primarily +3, also +2, +1

Gallium is the 34th most abundant element found in 

Earth’s crust (Emsley, 2001), and it is widely distributed in 

low amounts in many rock types (table H2). The average 

values for gallium in igneous rocks vary considerably, from 

a low of about 5 ppm gallium or less in ultramafic rocks to a 

range of 10 to 22 ppm for most mafic to intermediate rocks, 

and to about 16 to 35 ppm gallium for granitic rocks (Burton 

and Culkin, 1978). Alkaline rocks, particularly nepheline 

syenites, granites, and related pegmatites, have generally 

higher ranges, from 20 to 70 ppm gallium (Burton and Culkin, 

1978). Values of gallium in shale, sandstone, and carbonates 

are 19 ppm, 12 ppm, and 4 ppm, respectively (Mielke, 1979), 

and gallium in loess averages about 14 ppm (McLennan and 

Murray, 1999).

Gallium contents of rocks may be increased by greisenization, albitization, hydrothermal alteration, and weathering 

processes, including saprolitization and lateritization. In 

general, the bulk content of gallium in a given rock type 

corresponds to the composition and bulk content of feldspar 

and mica in the rock. Alteration processes that selectively 

remove mobile elements from aluminosilicate minerals 

result in relative increases in the concentrations of gallium, 

aluminum, and other less mobile elements in all rock types. 

The chalcophile behavior of gallium contributes to its 

concentration in sulfide ores, which can contain hundreds 

of parts per million gallium (Burton and Culkin, 1978). In 

hydrothermal systems related to sulfide mineralization and in 

residual deposits formed by surficial weathering, the lithophile 

and chalcophile tendencies of gallium work in opposition in 

terms of enrichment, as sulfide and aluminosilicate minerals 

compete for gallium.

The value of the ratio of aluminum to gallium (the 

Al:Ga ratio) has been used as an indicator of gallium enrichment because it is thought to be relatively constant in most 

types of crustal rocks (Burton and Culkin, 1978). The value 

of the Al:Ga ratio in various common rock types can range 

from 5,000 to 40,000. It is lower, in the range of 1,000 to 

10,000, in alkaline igneous rocks and in sedimentary rocks 

produced by intense weathering (Burton and Culkin, 1978; 

Tervek and Fay, 1986). Rytuba and others (2003) evaluated the 

geochemical behavior of gallium in igneous rocks using the 

worldwide GEOROC (Geochemistry of Rocks of the Oceans 

and Continents) database (Max Planck Institute, 2002). They 

found that volcanic rocks from convergent-margin environments with about 8 to 30 ppm gallium had values of Al:Ga that 

increased from about 3,000 to 8,000 with increasing aluminum 

content. They concluded that gallium is not enriched by 

petrologic processes in more evolved silicic volcanic rocks, 

such as peralkaline rocks, because the Al:Ga ratio did not 

correlate with increasing quartz (SiO2) content or alkalinity. 

A comparison of the data for volcanic rocks worldwide, hydrothermally altered rocks of the McDermitt caldera near the 

McDermitt Mine in Humboldt County, Nevada (from Rytuba 

and others, 2003), and globally distributed bauxite deposits 

shows that gallium concentration is higher in rocks that have 

lower ratios of Al:Ga owing to hydrothermal alteration and 

(or) intense weathering (fig. H6). Because the Al:Ga ratio can 

be quite variable in natural waters (table H2) owing to differences in the mobility of gallium and aluminum in aqueous 

solutions, the Al:Ga ratio has been used effectively to model 

the behavior of trace metals as a result of surficial weathering 

processes (Hieronymus and others, 1990; Shiller and Frilot, 

1996) and in ocean waters (Orians and Bruland, 1988; Shiller, 

1998) and river waters (Shiller, 1998; Shiller and Frilot, 1996).
Geology H9

Table H2. Gallium concentrations in rocks, ore, coal, soils, and natural waters.—Continued

[FOREGS, Forum of European Geological Surveys, now EuroGeoSurveys; Ga, gallium; ICP-MS, inductively coupled plasma-mass spectrometry; XRF, 

X-ray fluorescence. Units of measure: cm, centimeter; m, meter; mm, millimeter; ppb, part per billion; ppm, part per million; μg/L, microgram per liter;

μm, micrometer. —, no data]

Environment

and (or) location

Gallium concentration 

(value or range) Unit Type Notes Reference(s)

Typical Average

Rocks

Upper continental crust — 17, 17.5 ppm Continental crust Average Taylor and McLennan 

(1995); Rudnick and 

Gao (2003)

Bulk continental crust — 18 ppm Continental crust Average Taylor and McLennan 

(1995)

Lower continental 

crust

— 18 ppm Continental crust Average Taylor and McLennan 

(1995)

Volcanic rocks — About 8 to 30 ppm Convergent 

margins

2,929 analyses Rytuba and others (2003)

High-sulfidation 

epithermal

— Up to 320 ppm Alteration None Melfos and Voudouris 

(2012)

High-sulfidation 

epithermal

<9 to 38 ppm Alteration None Khashgerel and others 

(2008)

All types 8 to 812 41 ppm Bauxite None Schulte and Foley (2014)

Arkansas (lateritic 

bauxite)

— 86 ppm Bauxite None Burton and Culkin (1978)

Bosnia — 46 ppm Bauxite None Schulte and Foley (2014)

France — 63 ppm Bauxite None Schulte and Foley (2014)

India 5 to 122 60 ppm Bauxite None Burton and Culkin (1978)

Karst 10 to 812 41 ppm Bauxite None Schulte and Foley (2014)

Lateritic 8 to 136 56 ppm Bauxite None Schulte and Foley (2014)

Malaysia 22 to 47 33 ppm Bauxite None Burton and Culkin (1978)

Russia 10 to 170 54 ppm Bauxite None Burton and Culkin (1978)

Turkey — 68 ppm Bauxite None Schulte and Foley (2014)

Indiana—Danville 

member

1.7 to 8.9 5.06 ppm Coal Pennsylvanian Mastalerz and Drobniak 

(2012)

Indiana—Springfield 

member

1.4 to 12.3 3.39 ppm Coal Pennsylvanian Mastalerz and Drobniak 

(2012)

USGS COALQUAL 

database

About 3 to 7 — ppm Selected domestic 

coal-bearing 

regions, averages

Averages Finkelman (1993)

General <6 to 66 ppm Kaolin None Hieronymus and others 

(1990)

Georgia 27 to 100 — ppm Kaolin None This study; Bárdossy (1982); 

Bárdossy and Aleva (1990)

Haile kaolin deposit, 

South Carolina

<6 to 31 19 ppm Kaolin None This study

Malaysia 20 to 25 23 ppm Kaolin None Burton and Culkin (1978)

Geothermal fields 8 to 144 — ppm Muds and sinter None Crump (1994)
H10 Critical Mineral Resources of the United States—Gallium

Table H2. Gallium concentrations in rocks, ore, coal, soils, and natural waters.— Continued

[FOREGS, Forum of European Geological Surveys, now EuroGeoSurveys; Ga, gallium; ICP-MS, inductively coupled plasma-mass spectrometry; XRF, 

X-ray fluorescence. Units of measure: cm, centimeter; m, meter; mm, millimeter; ppb, part per billion; ppm, part per million; μg/L, microgram per liter;

μm, micrometer. —, no data]

Environment 

and (or) location

Gallium concentration 

(value or range) Unit Type Notes Reference(s)

Typical Average

Rocks—Continued

Alkaline

Carbonate rock

Granitic

Loess

Mafic to intermediate

Peralkaline

Shale

Sandstone

Ultramafic

Apex Mine, Utah

Ruby Creek Mine, 

Alaska

Cadmium-lead-zinc 

ore, general

Kipushi ore, general

Mississippi Valleytype ore, general

Sulfide ore, bulk 

Callahan Mine, Maine

Red Dog Mine, 

Alaska

—

—

—

—

—

—

—

—

—

320 to 1,480

—

<1 to 1,600

<1 to 2,500

<1 to 500

20 to 1,000

—

—

About 20 to 70

4

About 16 to 35

14

About 10 to 22

About 28 to 39

19

12

About 5 

—

<<1

60

10s to 100s

57

100?

25

26

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

Rock

Rock

Rock

Rock

Rock

Rock

Rock

Rock

Rock

Sulfide

Sulfide

Sulfide

Sulfide

Sulfide

Sulfide

Volcanogenic 

massive sulfide, 

bulk sulfide

Clastic-dominated,

ore concentrate 

 None

Average

 None

Average

 None

 None

Average

Average

 None

 None

 None

 None

 None

 None

 None

 None

 None

Burton and Culkin (1978)

Mielke (1979)

Burton and Culkin (1978)

McLennan and Murray 

(1999)

Burton and Culkin (1978)

MacDonald and Bailey 

(1973)

Mielke (1979)

Mielke (1979)

Burton and Culkin (1978)

Bernstein (1986)

Bernstein and Cox (1986)

Schulte and Foley (2014)

Schulte and Foley (2014)

Schulte and Foley (2014)

Burton and Culkin (1978)

Ayuso and others (2013)

Alaska Department 

of Environmental 

Conservation (2005)

Soils

World

Western United States

Eastern United States

Poland

Poland

Poland

Poland

Poland

Sweden

—

—

—

—

—

—

—

—

—

17

16

9.3

100 ± 5

219 ± 5

38 ± 4

93 ± 7

268 ± 19

7 to 13

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

ppm

Soil, bulk

Soil

Soil

Soil

Soil

Soil

Soil, proximal

to a former 

Zn-Pb mine

Soil, proximal to 

Zn-Pb ore and a 

bauxite mine

Soil developed 

on quartzite 

and gneiss

None

Mean for 20 cm 

depth

Mean for 20 cm 

depth

Agricultural 

Agricultural 

Agricultural 

Proximal to

a former 

Zn-Pb mine

Proximal to

Zn-Pb ore and a

bauxite mine

Profile range

Koljonen (1992)

Shacklette and Boerngen 

(1984)

Shacklette and Boerngen 

(1984)

Połedniok and others (2012)

Połedniok and others (2012)

Połedniok and others (2012)

Połedniok and others (2012)

Połedniok and others (2012)

Tyler (2004)
Geology H11

Table H2. Gallium concentrations in rocks, ore, coal, soils, and natural waters.— Continued

[FOREGS, Forum of European Geological Surveys, now EuroGeoSurveys; Ga, gallium; ICP-MS, inductively coupled plasma-mass spectrometry; XRF, 

X-ray fluorescence. Units of measure: cm, centimeter; m, meter; mm, millimeter; ppb, part per billion; ppm, part per million; μg/L, microgram per liter;

μm, micrometer. —, no data]

Environment 

and (or) location

Gallium concentration 

(value or range) Unit Type Notes Reference(s)

Typical Average

Soils—Continued

FOREGS samples, 

median concentration

FOREGS samples, 

median concentration

FOREGS samples, 

median concentration

—

—

—

0.9

13.5

13.8

ppm

ppm

ppm

Humus, median 

concentration

Topsoil

Subsoil

< 

< 

< 

2.0 mm fraction, 

n=367, Total 

(ICP-MS)

2.0 mm fraction, 

n=843, Total 

(ICP-MS)

2.0 mm fraction, 

n=790, Total 

(ICP-MS)

Salminen (2015)

Salminen (2015)

Salminen (2015)

Waters

Atlantic Ocean

African rivers

North American rivers

South American rivers

California streams

Idel River

Japan

Wairakei, New Zealand

Broadlands, New 

Zealand,

Bath, United Kingdom

Taiwan

World rivers

FOREGS samples, 

median concentration

FOREGS samples, 

median concentration

FOREGS samples, 

median concentration

—

—

—

—

—

—

—

—

—

—

—

18.1

—

—

—

0.0003 to 0.004

0.02 to 0.11

0.001 to 0.012

0.003 to 0.12

0.0001 to 0.006

0.009

2.5

About 0.2

About 0.2

0.63

0.01 to >10.00

 —

0.011

12.0

11.0

μg/L

μg/L

μg/L

μg/L

μg/L

μg/L

ppb

ppb

ppb

ppb

μg/L

ppm

μg/L

ppm

ppm

Seawater

River

River

River

River

River

Geothermal

Geothermal

Geothermal

Geothermal

Groundwater

Suspended 

sediment

Water

Stream 

sediment 

Flood plain 

sediment

Dissolved and colloidal (<0.4 μm); 

surface to 5,000 m 

deep

Dissolved load 

(<0.2 μm)

Dissolved load 

(<0.2 μm)

Dissolved load 

(<0.2 μm)

Dissolved and colloidal load (<0.4 μm)

Dissolved load 

(<0.2 μm)

 None

 None

 None

 None

 None

Average

Filtered < 0.45 micrometer fraction

< 2.0 mm fraction, 

n=852, Total (XRF)

< 2.0 mm fraction, 

n=747, Total (XRF)

Shiller (1998)

Gaillardet and others 

(2003)

Gaillardet and others 

(2003)

Gaillardet and others 

(2003)

Shiller and Frilot (1996)

Gaillardet and others 

(2003)

Uzumasa and Nasu (1960)

Goguel (1988)

Goguel (1988)

Riley (1961)

Chen (2006)

Viers and others (2009)

Salminen (2015)

Salminen (2015)

Salminen, 2015
H12 Critical Mineral Resources of the United States—Gallium

Mineralogy

Gallium is an essential structural component of only 

a few minerals that form in nature. Gallium is dispersed, 

however, in low amounts in many minerals and rocks, 

where it substitutes for elements of similar size and charge 

(table H3). Because gallium minerals are too rare in nature to 

serve as a primary source of the element or its compounds, 

most gallium is produced from minerals containing less than 

100 ppm gallium (table H3), where the gallium may occur as 

inclusions, as stoichiometric or nonstoichiometric substitutions, or as adsorbed ions or molecules.

Four gallium-rich minerals have been recognized 

(approved) by the International Mineralogical Association, 

Commission on New Minerals, Nomenclature and 

Classification (table H3). These include the sulfide gallite, 

which has about 35 weight percent gallium; gallobeudantite, 

which is an alunite of variable gallium content; and the 

hydroxide minerals sohngeite and tsumgallite, both of which 

contain about 60 weight percent gallium. These minerals have 

been identified in association with carbonate-hosted copperlead-zinc deposits at the Tsumeb Mine in the Oshikoto Region, 

Namibia (Frimmel and others, 1996, and references therein). 

Gallite occurs as a primary sulfide mineral and is essentially 

a common copper sulfide, chalcopyrite (CuFeS2

), in which 

gallium replaces iron in the crystal structure. Typical occurrences of gallite are as inclusions in sphalerite (ZnS). Because 

gallite has a chalcopyrite structure, it may occur as either 

epitaxial overgrowths or inclusions as “chalcopyrite disease” 

(Barton and Bethke, 1987) in sphalerite. Gallobeudantite, 

sohngeite, and tsumgallite are supergene minerals resulting 

from surficial oxidation of other gallium-bearing minerals. 

Gallium also substitutes in many minerals, including silicates, 

sulfides, and hydroxides, in concentrations ranging from 

parts per million to weight percent (table H3). For example, 

significant amounts occur in krieselite (about 20 weight 

percent gallium), which is the germanium analog of topaz, and 

in gallium-rich plumbogummite, where it may fill one-third 

of the aluminum sites (about 25 weight percent gallium). 

Major rock-forming minerals, however, contain significantly 

less than 1 percent gallium, and most minerals have less 

than 100 ppm gallium. Feldspars and micas are the dominant 

host minerals for gallium in most igneous and sedimentary 

rocks. For example, gallium has been reported at relatively 

100

1,000

10,000

20,000

30,000

1 10 30 100

Ga content, in parts per million

Al:Ga ratio

McDermitt Mine

 <30 ppm gallium

 ≥30 ppm gallium

y=–3,580Ln(x)+15,300

R2

=0.53

y= –1,200Ln(x)+6,100

R2

=0.40

y= –1,200Ln(x)+5,800

R2

=0.16

Normal range

of Al:Ga ratios in

volcanic rocks is

3,000 to 8,000

Bauxite deposits—15 to 90 ppm gallium

Worldwide volcanic rocks 

 (n=2,928)

EXPLANATION

Gallium figure 6 Figure H6. Plot of the ratio of aluminum to gallium (Al:Ga, log scale) versus gallium content (in parts per 

million [ppm], log scale) for bauxite deposits, volcanic rocks, and hydrothermally altered rocks of the McDermitt 

caldera, Nevada. Results from the McDermitt Mine, Nev., are included for comparison; see text for details. 

Equations define slope of lines drawn through datasets for comparison; R2

 value is a measure of goodness of 

the fit of the line to the data. Modified from Rytuba and others (2003) with the addition of data for various globally 

distributed bauxite deposits from Schulte and Foley (2014).
Geology H13

Table H3. Selected gallium-bearing minerals—Formula, content, and occurrence.

[Compiled from Burton and Culkin (1978), Bernstein (1986), Dutrizac and others (1986), Tervek and Fay (1986), Frimmel and others (1996), Schlüter and 

others (2003), Ye and others (2011), and Schulte and Foley (2014). b.d.l., below detection limit; ppm, part per million; ~, approximately. Elements: Al, aluminum; 

Cu, copper; Fe, iron; Ga, gallium; Pb, lead; Zn, zinc]

Mineral name Chemical formula Ga content Group Deposit model, occurrence type

Minerals with essential gallium1

 (weight percent)

Gallite CuGaS2 35.32 Sulfide Carbonate-hosted Cu-Pb-Zn; Pb-Zn-Cu 

Tsumeb Mine in Namibia

Gallobeudantite PbGa3

[(AsO4),(SO4)]2(OH)6 14.55, variable Alunite Carbonate-hosted Cu-Pb-Zn; supergene, 

Tsumeb Mine in Namibia

Sohngeite Ga(OH)3 57.74 Hydroxide Carbonate-hosted Cu-Pb-Zn; supergene, 

Tsumeb Mine in Namibia.

Tsumgallite GaO(OH) 59.93 Hydroxide Carbonate-hosted Cu-Pb-Zn; supergene, 

Tsumeb Mine in Namibia

Minerals with nonessential gallium 

Krieselite (Al,Ga)2

(GeO4)(OH)2 20.3 weight percent Topaz group, 

germanate

Ge analog of topaz

Sphalerite ZnS b.d.l. to 2,500 ppm Sulfide Volcanogenic massive sulfide, sedimentary 

exhalative, and Mississippi Valley-type 

base-metal deposits

Eyselite FeGe3

O7(OH) 6,900 ppm Hydroxide Carbonate-hosted Cu-Pb-Zn; Pb-Zn-Cu 

Tsumeb Mine in Namibia

Ovamboite Cu20(Fe,Cu,Zn)6

W2Ge6S32 4,600 ppm Sulfide, germanite 

group

Massive sulfide base-metal ores

Calvertite Cu5

Ge0.5S4 3,500 ppm Sulfide Carbonate-hosted Cu-Pb-Zn; Pb-Zn-Cu 

Tsumeb Mine in Namibia

Ferrohogbomite2N2S (Fe,Zn,Mg,Al)6

Al14(Ti,Fe)2

O30(OH)2

1,600 ppm Oxide High grade Fe-oxide deposits

Maikainite Cu20(Fe,Cu)6

Mo2Ge6S32 1,500 ppm Sulfide, germanite 

group

Volcanogenic massive sulfide, sedimentary 

exhalative, and Mississippi Valley-type

base-metal deposits. (1) Maikain deposit,

Kazakhstan, and (2) Tsumeb deposit, 

Namibia

Plumbogummite, 

Ga-rich 

PbH(Al,Ga)3

(PO4)2(OH)6 ~25 weight percent Crandallite subgroup, alunite 

supergroup

Carbonate-hosted Cu-Pb-Zn; Pb-Zn-Cu 

Tsumeb Mine in Namibia

Gibbsite Al(OH)3 About 50 ppm Oxide, hydroxide Bauxite-silicate; bauxite-karst

Diaspore α-AlO(OH) About 50 ppm Oxide, hydroxide Bauxite-silicate; bauxite-karst

Boehmite γ-AlO(OH) About 50 ppm Oxide, hydroxide Bauxite-silicate; bauxite-karst

Kaolinite Al2

Si2O5(OH)4 About 50 to 200 ppm Silicate Kaolins; bauxite-silicate; bauxite-karst

Halloysite Al2

Si2O5(OH)4 About 50 to 200 ppm Silicate Kaolins; bauxite-silicate; bauxite-karst

Jarosite KFe(SO4

)2(OH)6 0.7 weight percent Sulfite Supergene

Limonite FeO(OH) •nH2

O 2 weight percent Hydroxide Supergene

Goethite FeO(OH) 200 ppm Hydroxide Supergene

Renierite (Cu,Zn)11(Ge,As)2

Fe4S16 0.04 to 400 ppm Sulfide, germanite 

group

Massive sulfide base-metal ores

Germanite Cu26Fe4

Ge4S32 0.09 to 900 ppm Sulfide, germanite 

group

Massive sulfide base-metal ores

Feldspar group NaAlSi3

O8, CaAl2Si2O8, 

KAlSi3

O8

b.d.l. to 100 ppm, rarely 

up to 4,000 ppm

Silicate Felsic igneous rocks

1

Approved by the International Mineralogical Association, Commission on New Minerals, Nomenclature and Classification.
H14 Critical Mineral Resources of the United States—Gallium

high concentrations (0.05 weight percent) in feldspars from 

the Nechalacho layered nepheline-aegirine syenite suite at 

Thor Lake in the Northwest Territories, Canada (Tervek 

and Fay, 1986); however, most feldspar contains much less 

than 200 ppm gallium (table H3). Gallium is retained in clay 

minerals, such as kaolinite and smectite, and in aluminum 

hydroxide minerals during weathering, which accounts for 

the higher gallium contents in many clay-bearing rocks and 

bauxites (table H2). At the present time, gallium is derived 

mainly from aluminum hydroxide minerals, including 

boehmite, diaspore, and gibbsite (Al(OH)3), which form in 

deeply weathered deposits of bauxite (fig. H7A) (Dutrizac 

and others, 1986; Bernstein and Waychunas, 1987; Moskalyk, 

2003). Significant amounts of gallium are also produced from 

zinc sulfide deposits where sphalerite (fig. H7B) is the main 

carrier (Moskalyk, 2003). Gallium (and germanium) can 

also be enriched in other sulfide minerals, such as bornite, 

enargite, galena, and luzonite (Hörmann, 1978). Weathering of 

gallium-bearing sulfide minerals can lead to gallium-enriched 

sulfate and hydroxide minerals, and in oxidized environments, gallium commonly occurs in plumbojarosite and 

hydrogoethite (Burton and Culkin, 1978). At the Apex Mine 

in Summit County, Utah, jarosite (a potassium-iron-sulfate 

mineral) reportedly contains concentrations of up to 1 weight 

percent gallium (Bernstein, 1986).

Gibbsite-Diaspore-Boehmite

Gibbsite forms in weathering environments by a series 

of reactions generated during the hydrolysis of feldspar and 

other aluminosilicate minerals in rocks undergoing saprolitization (Foley and Ayuso, 2013, and references therein). The 

proportion of gibbsite-diaspore-boehmite increases at the 

expense of kaolinite in more intensely weathered rocks, such 

as bauxites and laterites (Soler and Lasaga, 1996). In these 

environments, as a result of burial diagenesis or low-grade 

regional metamorphism, gibbsite transforms to boehmite 

(γ-AlO(OH)) by dehydration, and turns into diaspore 

(α-AlO(OH)) after compaction. Gallium in these minerals 

may be found as substitutions for trivalent aluminum or it 

may form as a separate phase (table H3). Sohngeite, which 

is the gallium analog of gibbsite, has a similar stability, and 

it transforms readily into tsumgallite (GaO(OH)) in aqueous 

solution (table H3). The unit cell parameters of tsumgallite 

mostly resemble diaspore compared with gibbsite or boehmite 

(Schlüter and others, 2003). Gallium hydroxide minerals 

rarely form as discrete phases; rather, they generally occur in 

supergene parts of carbonate-hosted polymetallic replacement 

deposits, such as at the Tsumeb Mine in Namibia (Frimmel 

and others, 1996). Diaspore and tsumgallite are isostructural, 

have similar crystal chemical characteristics, and are the most 

Gallium Figure 7.

A B

Figure H7. Photographs showing samples of A, bauxite ore, and B, sphalerite ore, which are the primary 

mineralogical sources of gallium. The three samples of bauxite are from U.S. Geological Survey (USGS) collections 

and the powdered bauxite ore sample was provided by E.L. Bray, USGS. The sphalerite ore is from the Creede district, 

San Juan County, Colorado.
Geology H15

stable hydroxides of aluminum and gallium, respectively. 

Thus, gallium, in the form of GaO(OH), is more likely to be 

enriched in diaspore than in the other aluminum hydroxides.

Sphalerite

Sphalerite (fig. H7B) is an important source of gallium 

in spite of the generally low concentrations in natural samples 

(table H3) because it is abundant in a wide variety of mineral 

deposits (table H4). Gallium enrichment in sphalerite under 

hydrothermal conditions results from its weak chalcophile 

behavior, which is similar to that of zinc. Gallium is thought 

to be incorporated in sphalerite by a coupled substitution of 

two Zn for Ga plus either Ag, Cu, Ge, In, or Sn to maintain 

a balanced charge (Johan and others, 1983; Johan, 1988; 

Cook and others, 2009). The highest known concentrations 

of gallium in sphalerite (2.1 to 3.7 weight percent gallium 

occur in sulfide nodules in the Qingzhen (EH3) chondrite 

(Rambaldi and others, 1986), which is a range that matches 

experimental determinations (Krämer and others, 1987; 

Ueno and Scott, 2002). Sphalerite from gallium-rich low 

temperature carbonate-hosted lead-zinc deposits of Namibia 

(Khusib Spring Mine in the Otjozondjupa Region and the 

Tsumeb Mine) show a range of from hundreds to more than 

3,000 ppm gallium (Melcher and others, 2006). In general, 

however, concentrations in most sphalerites rarely exceed 

100 ppm gallium. For example, sphalerite from clasticdominated lead-zinc deposits of the Red Dog Mine in the 

Northwest Arctic Borough of Alaska contains an average 

of 26 ppm gallium (Alaska Department of Environmental 

Conservation, 2005), whereas zinc ores from U.S. Mississippi 

Valley-type deposits average about 50 ppm gallium (Moskalyk, 

2003). Gallium contents can vary widely in volcanogenic 

massive sulfide deposits, even within a single district or within 

individual sphalerite grains from a deposit. For example, some 

individual deposits within the Bonnifield mining district in 

Alaska (such as the Anderson Mountain deposit in Denali 

Borough and the Fosters Creek zone) can contain anywhere 

from 20 to 275 ppm gallium in sphalerite, whereas others (the 

West Tundra Flats deposit in Southeast Fairbanks Borough) 

have less than 1 ppm gallium in sphalerite (Foley and others, 

2008; Dusel-Bacon and others, 2011; Dusel-Bacon and 

others, 2012). Sphalerite from the Callahan Mine in Hancock 

County, Maine (Ayuso and others, 2013), shows a range of 

from less than 3 ppm to up to 105 ppm gallium in single, 

1- to 5-millimeter (mm)-wide crystals (table H3).

Deposit Types

The principal deposit types (table H4) and significant 

global and domestic deposits (fig. H4) from which gallium 

metal is currently being obtained are emphasized here; however, 

other deposit types that have gallium in potentially extractable 

amounts are included in table H4. Although gallium is found 

in a diverse group of mineral deposit types, current sources 

are limited to a small number of deposit types because gallium 

is produced mainly as a byproduct. More than 80 percent of 

the world’s refined gallium is obtained from bauxite deposits 

(Jaskula, 2013b). The balance is derived from operations that 

collect zinc residues from a variety of zinc deposit types; 

the residues are combined prior to the gallium extraction 

cycle. Among zinc deposits, gallium concentrations tend 

to be high in carbonate-hosted replacement deposits and 

Mississippi Valley-type deposits, and variably enriched in 

clastic-dominated systems and volcanogenic massive sulfide 

deposits, whereas skarns are universally low in gallium (Cook 

and others, 2009). Gallium and other metals produced from 

zinc-refining residues are derived largely from such deposits 

as the Red Dog zinc mine in Alaska (Tervek and Fay, 1986). 

Other sources include volcanogenic massive sulfide deposits; 

for example, the Tizapa Mine in Mexico (Dowa Holdings 

Co., Ltd., 2013). Carbonate-hosted lead-zinc deposits of the 

Kipushi-type—for example, the Apex Mine in Utah—have 

the potential to produce gallium (and germanium) as a 

primary product. The two major types of gallium sources—

bauxite deposits and sediment-hosted lead-zinc deposits—

are described below.

Bauxite Deposits

Bauxite is mined for a number of products that have 

commercial application in abrasive, cement, chemical, 

metallurgical, and refractory industries. Approximately 

70 to 80 percent of the world’s dry bauxite production is 

processed into aluminum metal (Freyssinet and others, 2005), 

and these are the ores from which most gallium metal is 

currently derived. Bauxite is a rock that consists mostly of 

fine-grained aluminum minerals, notably boehmite, diaspore, 

and gibbsite, in mixtures with lesser amounts of iron oxides, 

typically goethite and hematite, kaolinite, and small amounts 

of anatase (Bárdossy and Aleva, 1990). Some bauxites form as 

accumulations of clayey remnants within paleokarst disconformities in limestone sequences; these are referred to as karst 

bauxites (fig. H5A). Karst bauxite is typically black to gray in 

color because of the presence of admixed organic matter and, 

in some instances, minerals that form in chemically reduced 

environments, such as pyrite (Bárdossy, 1982). Another type 

of bauxite, lateritic bauxite, occurs in parts of thick altered 

profiles formed by lateritization of aluminosilicate rocks 

(fig. H5B) (Retallack, 2010). Lateritization is a complex 

integration of surficial biological, chemical, and physical 

processes that result in thick sequences of intensely weathered 

saprolite and soil. Laterites are a form of ferricrete that can 

occur as horizons within a soil profile. By definition, periods 

of lateritization record exceptionally intense, regionally 

widespread weathering under tropical temperature regimes 

with ample water (Bárdossy and Aleva, 1990); thus, lateritic 

bauxites occur in thick profiles formed by deep tropical 

weathering. A combination of climate, tectonics, geomorphology, and hydrogeological factors are required to produce 

deposits that make up productive bauxite ore districts and 

provinces (fig. H4; Bogatyrev, Zhukov, and Tsekhovsky, 2009). 
H16 Critical Mineral Resources of the United States—Gallium Table H4. Significant global and domestic deposit types from which gallium is obtained or is potentially extractable.—Continued [VMS, volcanogenic massive sulfide; Elements: Au, gold; Ag, silver; Al, aluminum; As, arsenic; Cd, cadmium; Co, cobalt; Cu, copper; Ga, gallium; Ge, germanium; Hg, mercury; In, indium; Pb, lead; Zn, zinc] Petrologic association Principal commodity (byproduct) Tectonic/deposit character Description Primary references Example deposits/occurrences Gallium host mineralogy

Volcanic-hosted, hydrothermal

Hydrothermal, 

hot springs

Au Active epithermal gold-depositing 

environments

Taupo Volcanic Zone, Rotokawa geothermal 

system, New Zealand

Muds and sinters of 

geothermal fields; alunite, other aluminous phosphate minerals

Crump (1994)

High-sulfidation 

epithermal gold

Au Aluminum-phophate-sulfate minerals 

occurring in the advanced argillic zone of alteration

Hugo Dummett porphyry copper-gold deposit,

Mongolia; Agios Philippos Mine, Kirki, Greece; Lavrion intrusion-related deposit, Plaka deposit, Greece

Aluminosilicate minerals, 

alunite, other aluminous phosphate minerals, aluminum hydroxide

Khashgerel and others 

(2008); Melfos and Voudouris (2012)

Epithermal 

mercury

Au, Ag, Hg Tertiary age andesite and rhyolite 

volcanic flows

Gold Canyon’s Cordero Gallium Project, 

McDermitt, Nevada, United States

Alunite, other aluminous 

phosphate minerals

Rytuba and others 

(2003)

Volcanogenic 

massive sulfide 

Zn, Cu, Pb 

(Au, Ag)

Submarine volcanoes of backarc basins Tizapa zinc mine, Mexico; Bonnifield 

district, Alaska; Bald Mountain, 

Maine, United States

Sphalerite Shanks and Thurston 

(2012); DuselBacon and others (2012)

Sediment-hosted

Clasticdominated 

lead-zinc

Zn, Pb, Cu, 

Ag (Ga, In, 

Ge)

Clastic-dominated lead-zinc ores, which are 

hosted in shale, sandstone, siltstone, or mixed clastic rocks, or occur as carbonate replacement, within a clastic-dominated sedimentary rock sequence

Red Dog zinc mine, Howards Pass, Su-Lik, 

and Anarraaq, Alaska, United States; 

Sullivan, British Columbia, Canada; Broken Hill, Australia; Balmat-Edwards, 

New York, United States; Rammelsberg, 

Meggen, Germany 

Sphalerite Leach, Bradley, and 

others (2010); Leach, Taylor, and others (2010); Taylor and others (2009)

Mississippi 

Valley-type

Pb, Zn, Cu, 

(Ge, Ag, Ga, 

As, Cd)

Carbonate-hosted lead-zinc deposits, 

which are typically found in platform carbonate sequences in passive-margin tectonic settings

Tri-State area in the United States (Southeastern Missouri, Eastern and Central Tennessee, 

Northern Arkansas); Pine Point, Canada; 

Polaris, Nanisivik, Alaska, United States; 

Navan, Ireland 

Sphalerite Leach, Bradley, and 

others (2010); Leach, Taylor, and others (2010); Taylor and others (2009)

Carbonatehosted 

Cu±Pb±Zn (alt. Tsumeb or Kipushi)

Cu, Pb, Zn 

(Co, Ge, Ag, 

Ga, As, Cd)

Carbonate-hosted (karstic) polymetallic 

deposits found in intracratonic platform and rifted continental margin sedimentary sequences; typically folded and locally faulted; often overlain disconformably by oxidized sandstone-siltstone-shale units. Largest deposits are epigenitic replacements within thick sedimentary sequences

Grinnell and Kanuyak Island, Northwest 

Territories, Canada; Gortdrum, Ireland; M’Passa, Republic of the Congo; Timna, Israel; Nifty, Australia; portions of the Dongchuan deposits, China; Ruby Creek 

and Omar deposits, Alaska, United States; Tsumeb and Kombat, Namibia; Kipushi, Democratic Republic of the Congo; Apex Mine, Utah, United States 

Sphalerite Söhnge (1964); 

Intiomale and Oosterbosch (1974); Bernstein (1986)
Geology H17

Table H4. Significant global and domestic deposit types from which gallium is obtained or is potentially extractable.—Continued[VMS, volcanogenic massive sulfide; Elements: Au, gold; Ag, silver; Al, aluminum; As, arsenic; Cd, cadmium; Co, cobalt; Cu, copper; Ga, gallium; Ge, germanium; Hg, mercury; In, indium; Pb, lead; Zn, zinc]

Petrologic association

Principal 

commodity 

(byproduct)

Tectonic/deposit character Description Primary references

Example deposits/occurrences Gallium host mineralogy

Weathering-related

Bauxite, karstic Al (Ga) Formed by deep, in situ weathering, transport, 

and accumulation of intercalated clays or by clay dissolution residues of limestone or other carbonate rocks, karst setting, accumulation of aluminum hydroxide minerals

Europe (France, Croatia, Italy); Jamaica; 

bauxite mine in Arkansas, United States; China

Gibbsite, diaspore, 

boehmite

Freyssinet and others 

(2005); Retallack (2010); Bárdossy (1982), Bogatyrev, Zhukov, and Tsekhovsky (2009)

Bauxite, lateritic Al (Ga) Formed by lateritization of various predominately silicate rocks such as granite, gneiss, 

basalt, syenite, and shale. Intense mineral hydrolysis resulting in kaolinite dissolution and precipitation of aluminum hydroxide minerals

Major bauxite provinces; mostly in the 

countries of the tropics (current and paleo): (1) the South American province. includes 

the Guiana shield (Guyana, French Guiana, Suriname) and the central Brazilian shield; 

(2) the West African province, includes Guinea, Ivory Coast, Burkina Faso, Sierra Leone, 

Ghana, and Mali; (3) the Indian province, includes both the western and eastern coasts of India; (4) the South East Asian province, includes Vietnam, Laos, and Cambodia; and 

(5) the Austalian province. Minor provinces 

include the southern United States and the southern Ural Mountains (Russia and Kazakhstan)

Gibbsite, diaspore, 

boehmite

Freyssinet and others 

(2005); Retallack (2010); Bárdossy and Aleva (1990); Bogatyrev, Zhukov, and Tsekhovsky (2009)

Supergene, 

gossan 

Ge, Ga Gossans and other oxidized zones of copperrich, arsenic-bearing sulfide deposits, 

particularly those of the carbonate hosted polymetallic Kipushi-type rocks

Apex Mine, Utah, United States; see 

Carbonate-hosted Cu±Pb±Zn, 

Kipushi-type

Jarosite, Fe-hydroxide, 

Ga-hydroxides

Bernstein (1986); 

Dutrizac and others (1986)

Coal Coal (Ga?) Coal seams and waste products from coalfired powerplants; gallium is enriched

in fly ash, sometimes 6 to 10 timesmore than in the feed coal

Jungar Coalfield, Inner Mongolia; 

West Virginia, United States; Canada 

Coal, fly ash, kaolinite, 

boehmite

Finkelman (1993)
H18 Critical Mineral Resources of the United States—Gallium

The formation of large and superlarge deposits is thought to 

require more than 10 million years or the superposition of 

several intense weathering epochs.

Various models for the origin of laterites and lateritic 

bauxites (for example, McFarlane, 1976; Brimhall and others, 

1991; Retallack, 2010) have included both (a) elemental 

removal by hydrolysis and chemical leaching, and 

(b) elemental addition by mineral precipitation or fluvial or 

atmospheric deposition (for example, eolian dust) to explain 

the chemical compositions and volume and mass losses 

in bauxites. Bauxites usually form in the pallid soil zone 

(fig. H5B) but can occur in the B or C horizons of exceptionally 

thick profiles (Retallack, 2010). Lateritic bauxites typically have 

distinctive pisolitic textures and red colors (fig. H7A). Karst 

bauxites that form by transport and redeposition of lateritic 

bauxite clays and rubble in paleokarst depressions may retain 

the laterite characteristics of color and texture (Retallack, 

2010). The extreme weathering conditions that contribute to 

bauxite formation remove virtually all the alkalis and alkaline 

earths, leaving only the relatively immobile component of the 

original rock; as a result, paleosols within bauxites and soils 

that form on bauxite are classified as oxisols (U.S. Natural 

Resources Conservation Service, Soil Survey Staff, 2000).

Bauxite deposits form in a diverse array of geologic and 

tectonic environments over a wide range of geologic time 

(for example, Bárdossy, 1982; Bárdossy and Aleva, 1990; 

Retallack, 2010). An extensive study of well-characterized 

and dated bauxites (Retallack, 2010) showed that although the 

oldest known bauxites formed on the Arabian Shield at about 

3,500 mega-annum (Ma, or million years ago), bauxites are 

most abundant in rocks formed since about 100 Ma owing to 

better preservation of the rock record (Wilkinson and others, 

2009). Bauxites show an uneven distribution through time that 

is consistent with evidence supporting their correlation with 

paleoclimatic events of intense weathering, possibly related 

to greenhouse gas emissions, which lasted for geologic time 

intervals of less than about 100,000 years (Valeton, 1983; 

Retallack, 2010). Bauxite formation during these events 

had an expanded distribution from the Quaternary (between 

the Tropics of Cancer and Capricorn) to encompass areas 

near the Arctic and Antarctic Circles (Retallack, 2010). As a 

consequence, greenhouse gas events of the past, which are 

shown to have influenced the distribution of aluminum ores 

(Retallack, 2010), are also associated with gallium ores.

Bauxite, lateritic bauxite, and karst bauxite are rocks that 

have concentrations of aluminum exceeding about 45 percent 

by weight, by definition (Valeton, 1972). This reflects a high 

proportion of hydroxide minerals with aluminum contents 

on the order of about 60 or more percent alumina by weight. 

Bauxite generally is defined as rock with 20 percent or less 

total iron, whereas lateritic bauxite, in addition to a high 

percentage of aluminum, can have greater than 20 percent 

total iron, which is generally ferric owing to oxidation unless 

reducing conditions prevail locally. The limited mobility of 

aluminum and gallium in near-surface environments accounts 

for gallium enrichment in bauxite and related materials 

(soils, kaolins, and ball clays) derived from deep weathering 

of crustal rocks. Overall, bauxite contains concentrations in 

the range of 8 to 800 ppm gallium, with most values in the 

range 40 to 80 ppm gallium (table H2). The world average 

concentration of gallium in bauxite was initially estimated 

to be 52 ppm (Burton and Culkin, 1978), and additional 

data have not changed that value significantly. For example, 

carbonate-hosted bauxites from central and southern 

Europe display average concentrations of about 50 ppm 

gallium (Schroll, 1999); bauxites from Arkansas range from 

50 to 100 ppm gallium (average 86 ppm); and bauxite from 

India has a similar range of 5 to 122 ppm gallium (table H2). 

High gallium values are reported in bauxite originating from 

alteration of alkali rocks (for example, nepheline syenite of 

Russia ranges up to 170 ppm gallium) and in some unusual 

karst deposits (table H2). In general, however, there appear 

to be no significant differences in gallium concentrations of 

karst- or lateritic-type bauxites (Schulte and Foley, 2014).

Sediment-Hosted Lead-Zinc Deposits

Sediment-hosted lead-zinc deposits are historically 

the most significant sources of zinc, and these deposits are 

mined throughout the United States and the world (Leach, 

Bradley, and others, 2010). The largest and most economically 

important sediment-hosted deposits for zinc production are 

the clastic-rock-dominated and carbonate-rock-dominated 

(Mississippi Valley-type) subtypes (Leach, Marsh, and others, 

2004; Leach, Bradley, and others, 2010; Leach, Taylor, and 

others, 2010); these are the main types of sulfide deposits 

from which gallium metal is obtained at present because they 

are currently the predominant sources of zinc ore. Carbonatehosted polymetallic deposits of the Kipushi (or Tsumeb) type 

are of minor importance for zinc production; however, they 

are considered separately here because they have high gallium 

content relative to other deposit types of the sediment-hosted 

class and are historic gallium producers for the United States. 

Much of the following discussion is summarized from a 

number of seminal review studies of sediment-hosted leadzinc deposits, including Leach and Sangster (1993), Leach, 

Viets, and others (1995); Leach, Marsh, and others (2005); 

Leach, Sangster, and others, (2005); Groves and Bierlein 

(2007); Leach, Bradley, and others (2010); and Leach, Taylor, 

and others (2010).

Sediment-hosted lead-zinc deposits form in rocks 

of widely varying geologic age and tectonic environment 

(figs. H5C through H5E). In general, these deposits are hosted 

by large-scale sedimentary systems consisting of a variety of 

siliciclastic and carbonate rocks and have no direct genetic 

association with igneous activity. Many deposits in the 

sediment-hosted class display transitional features, and the 

distinctions among some of the deposit types are subjective. 

Most sediment-hosted deposits occur in sedimentary rock 

sequences that were originally deposited in extensional rift 

or passive-margin settings. There is an apparent uneven 

distribution of sediment-hosted deposits through time that is 
Geology H19

attributed to their genesis in large-scale sedimentary systems 

and to continued recycling of the sedimentary rock record 

within the evolving tectonic and geochemical systems of Earth 

through geologic time. Tectonic setting, depositional environment, and host-rock lithology influence the mineral chemistry 

of the contained sulfide minerals, including their trace element 

characteristics. The ores consist mainly of galena, sphalerite, 

and generally lesser amounts of iron sulfides, which may 

be marcasite or pyrite. In addition to lead and zinc, gallium, 

germanium, and silver can be important constituents; copper 

is generally quite low, except in the Kipushi-type deposits, 

but can be economically important in all these deposits. 

Laminated ore textures and a layered morphology are 

generally characteristic of clastic-dominated deposits, whereas 

Mississippi Valley-type deposits generally exhibit abundant 

evidence for epigenetic and replacement ores. Nonetheless, 

some clastic-dominated ores show replaced sediments in 

early or burial diagenetic environments and some Mississippi 

Valley-type deposits display laminated ore textures.

Clastic-Dominated Lead-Zinc Type

The most definitive characteristic of clastic-dominated 

deposits as described by Leach, Marsh, and others (2005) 

is their occurrence in clastic rock-dominated sedimentary 

sequences formed mainly in passive margins, continental 

rifts, continental sag basins, and back-arc basins (figs. H5C

and H5D). The orebodies are typically hosted in (in order 

of predominance) shale, sandstone, siltstone, carbonate, or 

mixed clastic rocks and their metamorphosed equivalents. 

They may also occur as carbonate replacement ores within 

a clastic-rock-dominated sedimentary sequence.

The vast majority of clastic-dominated deposits 

globally occur in Proterozoic basins that formed between 

2.02 billion years ago—the age of the earliest known deposit 

of these ores—and about 1.85 to 1.58 billion years ago, 

which reflects a major period of clastic-dominated lead-zinc 

deposition in Australia and India. The appearance of clasticdominated deposits in the Precambrian period corresponds to 

a time of formation of large and long-lived basins on stable 

cratons that followed the Great Oxidation Event at about 

2.4 to 1.8 billion years ago. The formation of clastic-dominated 

deposits is related to enhanced oxidation of sulfides in the crust 

(which released sulfate to the hydrosphere and lead and zinc to 

sediments); the development of major redox and compositional 

gradients in the oceans; the first formation of significant sulfatebearing evaporates, red beds, and oxidized aquifers; and the 

evolution of sulfate-reducing bacteria (Leach, Bradley, and 

others, 2010, p. 593). A number of clastic-dominated lead-zinc 

deposits also formed during the Carboniferous Period along 

passive margins in sedimentary-evaporite belts of Pangea and 

were deposited after the second oxidation event.

Mineralogical and geochemical evidence suggests that 

clastic-dominated ore fluids were principally hot metalliferous basinal brines (for example, Badham, 1981; Lydon, 

1983; Cooke and others, 2000). For example, the brines that 

deposited sphalerite at the Red Dog deposit in Alaska had 

temperatures in the range of 100 to 200 °C and salinities 

from 14 to 19 weight percent NaCl equivalent (Leach, Marsh, 

and others, 2004). The dominant sulfide minerals present 

are sphalerite (which is generally the most common sulfide 

mineral) and lesser amounts of galena and the iron sulfides, 

marcasite, pyrite, or pyrrhotite; chalcopyrite and sulfosalt 

minerals may be present in minor amounts (Large, 1981, 1983; 

Lydon, 1983; Large and others, 2005). The sulfide minerals 

form lens-like bodies that are interbedded with fine-grained 

dark clastic and chemical sedimentary rocks. Carbonate 

minerals (ankerite, calcite, dolomite, and siderite), barite and 

quartz, siliceous shale, and low-grade silica-replaced rocks 

are common and can be gradational into massive ore. Ore 

from these deposits (table H4) generally contains less than 

100 ppm gallium, with a typical range of 5 to 50 ppm gallium 

(table H2). For example, the Red Dog deposit (fig. H4) is 

hosted by organic-rich mudstone and shale of the Mississippian 

Kuna Formation (Kelley and others, 2004), and zinc concentrate from the deposit has 26 ppm gallium (Alaska Department 

of Environmental Conservation, 2005). Other elements in 

association with clastic-dominated lead-zinc deposits include 

Ag, Al, As, Au, B, Ba, Cd, Co, Cu, Fe, Ge, Hg, Mn, Mo, Ni, 

and Sb (Briskey, 1986; Kelley and others, 1995).

Mississippi Valley Lead-Zinc Type

Carbonate-hosted lead-zinc deposits are found throughout 

the world, and large famous examples occur in central North 

America; hence, they are generally referred to as Mississippi 

Valley-type deposits. The most important characteristics of 

Mississippi Valley-type deposits (Sangster, 1990; Leach and 

Sangster, 1993; Leach, Sangster, and others, 2005) are that 

they occur globally in dolostone and limestone that formed in 

platform carbonate sequences. They are commonly related to 

extensional domains inboard of contractional tectonic belts or 

continental extensional basins transitional to passive margin 

basins (figs. H5C and H5E). The ores are usually located 

at flanks of basins, orogenic forelands, or foreland thrust 

belts inboard of the clastic-rock-dominated passive margin 

sequences. Numerous lines of evidence have established that 

the metals were carried in fluids derived mainly from evaporated seawater and driven within the carbonate sequences by 

large-scale tectonic events.

More than 80 percent of Mississippi Valley-type deposits 

are hosted by rocks formed during the Phanerozoic Eon; 

less than 20 percent occur in Precambrian rocks (Sangster, 

1990; Leach, Bradley, and others, 2001; Leach, Sangster, and 

others, 2005; Kesler and Reich, 2006). Increased oxygenation 

of the oceans following the second oxidation event of the 

atmosphere in the late Neoproterozoic Era led to an abundance 

of evaporites, oxidized brines, and a dramatic increase in 

the volume of coarse-grained and permeable carbonates in 

Paleozoic carbonate platforms. Most Phanerozoic-hosted 

Mississippi Valley-type deposits formed in the carbonate 

platform sequences during Devonian to Permian time as a 
H20 Critical Mineral Resources of the United States—Gallium

result of tectonic events generated by the assimilation of 

Pangea. A lesser number formed during Cretaceous to Tertiary 

time when microplate assimilation affected the western margin 

of North America and Africa-Eurasia.

These ores consist mainly of galena, sphalerite, and 

generally lesser amounts of iron sulfides. Copper and silver are 

important constituents in some of these deposits (for example, 

deposits of the Irish subtype and some deposits of the 

Viburnum Trend). Gangue minerals are dominantly carbonates 

(ankerite, calcite, dolomite, and siderite), with typically minor 

amounts of barite. Silicification of the host rocks (or quartz 

gangue) is generally minor, but may be abundant in a few 

deposits. Minor minerals associated with Mississippi Valleytype mineralization differ widely between individual districts, 

and minor and trace metals also vary (Lavery and others, 

1994). Minor elements and element associations can include 

Ag, As, Au, Ba, Bi, Cd, Co, Cu, Ga, Ge, In, Fe, Hg, Mn, Mo, 

Ni, Sb, Sn, and Tl (Hall and Heyl, 1968; Hagni, 1983; Viets 

and others, 1992; Leach, Taylor, and others, 2010).

In Mississippi Valley-type deposits, gallium is primarily 

in sphalerite, which typically can contain from 0.02 to more 

than 500 ppm gallium (table H4). The overall content of 

gallium in studied deposits averages about 58 ppm; however, 

averages for individual deposits can vary widely, from 

6 to 130 ppm gallium.

Kipushi Polymetallic Type

Kipushi-type deposits are relatively rare, epigenetic, 

discordant deposits of Cu–Zn–Pb –Ag–As–Sb–Ge–Ga-bearing 

ores that occur in carbonate platform rock sequences and are 

typically associated with karst, solution collapse breccias, 

and related features, such as salt diapirs (fig. H5F; Hitzman 

and others, 2005). The most important characteristics of these 

deposits are their geologic setting and distinctive polymetallic 

mineralogy. Considerable controversy remains concerning 

whether these deposits have an igneous affinity; however, 

most are classified as sediment-hosted deposits. Deposits of 

the Kipushi type (fig. H4) include world-class polymetallic 

orebodies at Tsumeb and Khusib Springs in Namibia (Söhnge, 

1964), and at Kipushi (in the Democratic Republic of the 

Congo) in the Zaire-Zambia copper belt (Intiomale and 

Oosterbosch, 1974). Domestic deposits of this type include 

orebodies at the Ruby Creek Mine in Alaska (Hitzman, 1986; 

Bernstein and Cox, 1986; Selby and others, 2009) and at the 

Apex Mine in Utah (Bernstein, 1986). Other large deposits 

included in this group in the past are, in Australia, the Mount 

Isa Mine in the State of Queensland, and the Cooley and Ridge 

deposits of the McArthur River Mine in the Northern Territory 

(Bernstein and Cox, 1986; Hitzman, 1986; Selby and others, 

2009); more recent studies, however, suggest that these deposits 

are better classified as clastic-dominated-type deposits because 

they are mainly lead-zinc orebodies in dolomitic siltstones, 

mudstones, and sandstones formed in a continental sag basin 

(Cooke and others, 2000; Leach, Marsh, and others, 2005; 

Leach, Bradley, and others, 2010; Taylor and others, 2009).

Features of Kipushi-type deposits that are similar to 

aspects of clastic-dominated-type and Mississippi Valley-type 

deposits include the general platform carbonate setting, a lack 

of spatially associated igneous intrusions, regional-scale ore 

fluid migration (as evidenced by the widespread occurrence 

of mineralization over thousands of square kilometers in the 

largest deposits), and ore localization in solution collapse 

breccias. The orebody at Kipushi-type deposits forms a 

highly irregular subvertical pipe at the faulted boundary 

of dolomitic shale and marine dolomite, commonly within 

breccia zones, and prospecting for copper-rich ores is 

focused on paleokarst structures that are possibly related to 

salt diapirs (Intiomale and Oosterbosch, 1974). The Khusib 

Springs deposit, which is a lens-shaped sulfide orebody up to 

10 meters in thickness and elongated parallel to the bedding, 

replaces locally brecciated limestone of the lower Tsumeb 

subgroup. The origin of the breccia pipes is probably related 

to the erosion and karstification of the Hüttenberg Formation 

(Hughes and others, 1984; Hughes, 1987); however, there is 

some evidence that hydrothermal brecciation and silicification 

occurred during mineralization (Chetty and Frimmel, 2000; 

Kamona and Günzel, 2007). Fluid inclusion studies have also 

established that mineralizing fluids for Kipushi-type deposits 

are comparable to or higher than those of typical Mississippi 

Valley-type deposits and clastic-dominated lead-zinc deposits 

(see above). Ore-stage minerals contain highly saline fluids 

(17 to 43 weight percent NaCl equivalent) in inclusions 

with homogenization temperatures that vary from about 

100 to 380 °C; the fluids were generated in postorogenic, 

extensional settings (Misiewicz, 1988; Chetty and Frimmel, 

2000; Selby and others, 2009; Haest and others, 2009; 

van Wilderode and others, 2012).

Kipushi-type deposits are distinguished from Mississippi 

Valley-type deposits by having significantly higher trace 

element concentrations and more diverse mineralogy 

(Frimmel and others, 1996). Primary minerals include 

enargite, galena, and pyrite, and a zinc-rich tennantite; 

locally abundant zones of bornite, chalcocite-digenite, and 

chalcopyrite; galena, pyrite (cobaltiferous), pyrrhotite, and 

iron-poor sphalerite. Late-stage phases include carrollite 

and copper-thiogermanate minerals, pearceite-polybasite, 

native silver, tetrahedrite, and silver-tennantite. Additional 

minerals identified in some deposits include covellite, a 

tungsten-bearing germanite, germanium-bearing stannoidite, 

and stromeyerite. Gangue minerals include barite and bariumsilicate minerals, calcite, chlorite, dolomite, fluorite, sparse 

quartz, and siderite. Ore minerals in Kipushi-type deposits 

can have gallium contents that range from hundreds of parts 

per million gallium to about 60 weight percent gallium 

(tables H2 and H3). Compared with other sediment-hosted 

ore systems, Kipushi-type deposits are characterized by 

marked enrichments in germanium and rhenium (Schneider 

and others, 2007). Many weathered deposits of this type 

contain gallium-bearing secondary minerals formed largely 

through supergene processes (table H3). For example, at the 

Apex Mine in Utah, the primary ore is inferred to be similar 
Resources and Production H21

to bornite-chalcopyrite-tennantite-galena-sphalerite ore of 

a Kipushi type, although the Apex orebodies are almost 

completely oxidized to germanium-bearing hydrogoethite and 

gallium-bearing (up to about 1 percent gallium) jarosite.

Resources and Production

Identified Resources

The U.S. Geological Survey (USGS) estimates that world 

resources of gallium in bauxite exceed 1 billion kilograms 

and that a considerable quantity of gallium could be present 

in world zinc resources (Bray, 2013). Using USGS information 

coupled with data on global bauxite provinces (Bogatyrev 

and Zhukov, 2009), comparable estimates of geologically 

available gallium in bauxite reserves and resources (of nearly 

1.4 million metric tons and 4 million metric tons, respectively) 

have been suggested (Dittrich and others, 2011). Analogous 

estimates of world resources of gallium in zinc resources have 

not been made owing to insufficient data. The United States 

is expected to meet its current and expected future needs 

for gallium through imports of primary gallium from these 

resources, from imports of recycled and refined gallium, and 

from domestic production of recycled and refined gallium.

Identified bauxite resources of the world are about 

55 to 75 billion metric tons and occur in Africa (32 percent), 

Oceania (23 percent), South America and the Caribbean 

(21 percent), Asia (18 percent), and elsewhere (6 percent) 

(Bray, 2012). Lateritic bauxite deposits make up approximately 

90 percent of global bauxite resources, and aluminum production is mainly from these deposits (Freyssinet and others, 

2005). Karst bauxite deposits are more scattered in distribution 

and are mined mainly in eastern Europe and the Caribbean 

islands (Bárdossy, 1982). Domestic bauxite deposits consist 

mainly of subeconomic resources that are not generally suitable for alumina production owing to their high silica content; 

thus, recovery of gallium from these deposits is unlikely. 

Although domestic resources of bauxite are inadequate to meet 

either short- or long-term U.S. demand, the United States has 

essentially inexhaustible subeconomic resources of aluminum 

in materials other than bauxite. These are primarily clay 

deposits, which also contain gallium in amounts comparable to 

those of bauxite (30 to 50 ppm gallium) (table H2). Identified 

zinc resources of the world are about 1.9 billion metric tons 

(Tolcin, 2013). Mississippi Valley-type deposits, which are 

widespread in North America, account for about 38 percent 

of the global resources of lead and zinc in sediment-hosted 

ore deposits; significant zinc resources are also contained in 

clastic-dominated deposits of Alaska (fig. H4; Leach, Marsh, 

and others, 2005).

The estimates given here apply to total gallium content; 

only a small percentage of gallium metal in bauxite and zinc 

ores is recoverable using current methods because processing 

techniques to separate aluminum, zinc, and other byproduct 

metals (for example, germanium and indium) compete for 

gallium in source materials. Factors controlling gallium 

recovery are typically proprietary, and scant literature is available on the general operating practices of companies involved 

in gallium production. The potential remains for additional 

gallium production given more efficient and improved extraction technologies and market-driven economic incentives.

Undiscovered Resources

The USGS has conducted no quantitative mineral 

resource assessments focused on undiscovered domestic or 

global resources of bauxite since the work of Patterson and 

others (1986). Because the major bauxite provinces of the 

world are well established and have been intensely explored, 

most of the significant bauxite deposits globally are thought 

to be known (Freyssinet and others, 2005). The major 

bauxite provinces (fig. H4) occur in the following five major 

regions: (a) the South American region—the Guiana shield 

(French Guiana, Guyana, Suriname) and the central Brazilian 

shield; (b) the West African region—Burkina Faso, Ghana, 

Guinea, Ivory Coast, Mali, and Sierra Leone; (c) the Indian 

region—the western and eastern coasts; (d) Southeast Asia—

Cambodia, Laos, and Vietnam; and (e) the Australian region. 

Other bauxite provinces occur in the southern United States 

(Arkansas region), the Caribbean, the southern Ural Mountains 

(Russia and Kazakhstan), and China.

The most recent USGS National Mineral Resource 

Assessment identified areas of the United States that have 

significant potential for undiscovered resources of zinc 

(Schruben, 2002). The estimate of the amount of zinc 

in undiscovered U.S. deposits ranged from a 90-percent 

probability of at least 130 million metric tons to a 10-percent 

probability of at least 290 million metric tons. The mean estimate of zinc in undiscovered deposits was 210 million metric 

tons. Nearly 40 percent of the zinc was thought to be contained 

in undiscovered Mississippi Valley-type deposits. Other 

major deposit types were sedimentary exhalative (31 percent), 

volcanogenic massive sulfide (15 percent), and polymetallic 

replacement (8 percent) deposit types. At that time, the total 

identified zinc resources in the United States were estimated 

to be 99 million metric tons. No comparable estimates of 

undiscovered global resources of zinc have been made.

Production

In 2012, world primary gallium production was 

estimated to be 273 metric tons (fig. H4A). China, Germany, 

Kazakhstan, and Ukraine were the leading producers; countries with lesser output were Hungary, Japan, the Republic of 

Korea, and Russia. Refined gallium production was estimated 

to be about 354 metric tons; this estimate includes primary 

gallium production and some scrap refining. China, Japan, 

the United Kingdom, and the United States were the principal 

producers of refined gallium. Gallium was recycled from new 

scrap in Canada, Germany, Japan, the United Kingdom, and 
H22 Critical Mineral Resources of the United States—Gallium

the United States. World primary gallium production capacity 

in 2012 was estimated to be 474 metric tons; refinery capacity, 

270 metric tons; and recycling capacity, 198 metric tons 

(Jaskula, 2013a). Imports of gallium into Japan and the 

United States (the two leading consuming countries) were 

used as the basis for estimating world gallium production. 

In addition, Metal Bulletin (2013) provided an updated 

Chinese gallium production estimate, expanding the estimate 

of China’s production considerably owing to the substantial 

capacity increases of China’s gallium operations (fig. H4B).

No domestic production of primary gallium was 

reported in 2012. Molycorp Inc. recovered gallium from 

scrap materials, predominantly those generated during the 

production of GaAs. Molycorp’s facility in Blanding, Utah, 

has the capability to produce about 50 metric tons per year of 

high-purity gallium. The company recovered gallium from 

its customers’ scrap on a fee basis and purchased scrap and 

low-purity gallium for processing into high-purity material.

Exploration for New Deposits

Most domestic exploration for gallium is done as 

a second-order effort related to new discoveries of zinc 

resources because U.S. bauxite resources are currently 

subeconomic. Exploration is active in the United States and 

globally for all types of large-tonnage zinc deposits, which 

constitute a major source for gallium (Emsbo, 2009; Taylor 

and others, 2009; Leach, Bradley, and others, 2010). There is 

some evidence that partitioning of gallium may be influenced 

by the amount of sphalerite in bulk ore, leading to a dilution 

effect, such that the largest zinc deposits of a given type 

may be expected to have relatively lower gallium contents 

(Cook and others, 2009). This, however, may have little effect 

on resource estimation because economic factors that support 

gallium extraction are tilted in favor of large tonnage values 

for the principal commodity over small changes in grade for 

a byproduct element. The most gallium-enriched sedimenthosted ores are the Kipushi-type deposits, which are mined 

for copper, lead, silver, and zinc and byproduct antimony, 

arsenic, germanium, and gallium; some deposits of this group 

may contain gallium (and germanium) in amounts that justify 

extraction as a primary product. Active exploration efforts are 

underway in the major Kipushi-type provinces of the Central 

African copperbelt—Kipushi and Tsumeb. Important domestic 

examples of Kipushi-type deposits include the Ruby Creek 

deposit in Alaska and the Apex deposit in Utah. Additional 

domestic exploration efforts are focused on gallium resources 

associated with epithermal mineralized and hydrothermally 

altered volcanic rock suites.

Ruby Creek deposit, Alaska.—This deposit is located along the 

northern flank of the Cosmos Hills near the boundary between 

a carbonate platform and a shale basin within the Ambler 

district. The 1-kilometer-diameter body of hydrothermal 

dolostone is one of several similarly mineralized occurrences in the Cosmos Hills. The Ruby Creek deposit contains 

more than 90 million metric tons of 1.2 percent copper, plus 

elevated contents of silver, cobalt, germanium, gallium, and 

zinc (Hitzman, 1986). The metacarbonate sequence consists 

primarily of thin-bedded to massive metadolostone and 

marble that is locally graphitic or phyllitic and contains thin 

micaceous layers interpreted as airfall tuffs (Hitzman, 1986; 

Hitzman and others, 1986). The presence of coeval volcanogenic massive sulfide deposits just north of the Ruby Creek 

deposit is possible evidence for regionally extensive hydrothermal activity, perhaps related to igneous activity associated 

with continental extension (Hitzman and others, 1986; Selby 

and others, 2009).

Apex deposit, southwestern Utah.—This deposit has served 

as the only primary source in the world for the production of 

gallium and germanium. The Apex Mine operated intermittently 

from 1884 to 1962, and more than 7,000 metric tons of copper, 

5,600 kilograms of silver, and small amounts of gold and lead 

were produced. The mine was reopened in 1985 with plans to 

produce gallium and germanium (Washington County Historical 

Society, 2013). At that time, the mine was estimated to contain 

220,000 metric tons of ore grading 0.064 percent germanium, 

0.032 percent gallium, 1.63 percent copper, 0.77 percent 

lead, 1.58 percent zinc, 0.50 percent arsenic, 16.73 percent 

iron, and 38 grams per metric ton silver (Bernstein, 1986). 

Surface dumps from the old mine contained an additional 

45,000 metric tons of material grading 0.037 percent 

germanium, 0.019 percent gallium, and 1.55 percent copper. 

Similar deposits may occur in association with copper 

in fault-hosted breccias in Pennsylvanian limestone near 

Hansonburg, New Mexico (Eveleth and Lueth, 2009).

Cordero Gallium Project, Humboldt County, Nev.—This 

project is reportedly North America’s largest known primary 

gallium occurrence (Rêserva International LLC, 2008). 

Anomalous gallium mineralization occurs over a wide area 

in and around the historic Cordero and McDermitt mercury 

mines. There are indications that a number of high-grade 

“feeder zones” occur through the deposit, which also hosts 

significant occurrences of rare-earth elements. The resources 

for the project above a cutoff grade of 30 ppm gallium are estimated to exceed 13 million metric tons (Rêserva International 

LLC, 2008).

Environmental Considerations

Sources and Fate in the Environment

In weathering environments, gallium is generally 

immobile and is only slightly less reactive than aluminum 

(Shiller and Frilot, 1996). The tendency for gallium to be 

dissolved and transported depends upon its chemical form, 

which in turn depends largely upon the pH and temperature 

of weathering solutions. Gallium associated with bauxite at 

a pH near 4 and a weathering solution temperature of 25 oC 

would be expected to remain as a solid (for example, Ga(OH)3 
Environmental Considerations H23

or GaO(OH)), but can theoretically dissolve to form aqueous 

hydroxide complexes (Ga(OH)3 or Ga(OH)4 at pH values >4 

and Ga(OH) and Ga(OH)2 at pH values <4; Bénézeth and 

others, 1997; Diakonov and others, 1997). Gallium is expected 

to be transported in soils, streams, and rivers in the solid phase 

as small particulates because of the low solubility of gallium 

minerals in natural waters.

Gallium associated with sulfide minerals is more 

mobile based on the instability of sulfides at Earth’s surface. 

Dissolution of sulfide minerals releases metals and sulfuric 

acid, and the acidic pH values allow higher concentrations 

of metals to be dissolved in solution—potentially causing 

the environmental problem known as acid mine drainage 

(AMD). Under acidic conditions, gallium would most likely 

form dissolved complexes with sulfate and (or) chloride 

(for example, GaSO4

+

 or GaCl−) (Wood and Samson, 2006). 

Metals dissolved in AMD can be naturally attenuated through 

precipitation, sorption to oxyhydroxide minerals, or dilution 

by mixing with water at circumneutral pH; otherwise, metals 

are transported downstream.

Examples of natural concentrations of gallium in soil, 

water, and air are given in table H2. Gallium contents in 

soils range from 3 to 70 milligrams per kilogram (mg/kg) 

(or 3 to 70 ppm) (Kabata-Pendias and Pendias, 2001), and 

concentrations as high as 220 mg/kg have been observed 

in agricultural soil (Połedniok and others, 2012). Gallium 

generally behaves like aluminum in the environment, but, 

under certain pH and redox conditions, the two elements 

may be separated. As a result, values of the ratio of gallium 

to aluminum (Ga:Al ratio) can aid in detecting gallium 

enrichment in various materials. Shiller and Frilot (1996) 

showed that the Ga:Al ratio of parent rocks (and residual 

soils) and streambed solids were similar, whereas the Ga:Al 

ratio in nearby streams was higher than those in parent 

rocks. The concentration of gallium in these streams ranged 

from 0.0001 to 0.006 micrograms per liter (μg/L) (or 

0.0001 to 0.006 parts per billion). These concentrations are 

generally lower than the 0.001 to 0.12 μg/L range of dissolved 

gallium found in large rivers around the world and are comparable to gallium concentrations measured in the Atlantic Ocean 

(Shiller, 1998). Compared with dissolved species, gallium is 

highly concentrated in small particulates in river water, with 

a global average of 18.1 mg/kg in suspended sediment (Viers 

and others, 2009). Gallium occurs naturally in the atmosphere 

as part of mineral dust particles and has been measured to 

be less than 0.14 nanograms per cubic meter (ng/m3 ) at the 

South Pole (Kabata-Pendias and Pendias, 2001).

Mining and ore processing can hasten the weathering 

process and can lead to above-background concentrations of 

gallium in soil and water (table H2). For example, gallium 

concentrations ranging from 93 to 270 mg/kg have been 

observed in soils near bauxite and lead-zinc mine sites in 

Poland (Połedniok and others, 2012). In addition, anthropogenic release of gallium to the atmosphere occurs during 

coal combustion, ore processing, and semiconductor manufacturing, leading to concentrations as high as 1 to 12 ng/m3

 near 

industrial areas (Chen, 2007; Kabata-Pendias and Mukherjee, 

2007). Likewise, dissolved and colloidal concentrations of 

gallium in groundwater near semiconductor manufacturing 

areas in China have been measured to be greater than 10 μg/L 

(Chen, 2006). The recycled content, or proportion of scrap, 

used in gallium production is between 10 and 25 percent, but 

the fraction of gallium in discarded products that get recycled 

is less than 1 percent (Graedel and others, 2011). One of the 

main challenges of gallium recycling is that it is commonly 

comingled with other “specialty metals” in high performance 

alloys, making recovery technologically and economically 

unfeasible (Reck and Graedel, 2012). Given the importance of 

gallium in thin-film photovoltaic cells and integrated circuits, 

however, gallium recycling is likely to increase in the future 

(Graedel, 2011).

Mine Waste Characteristics

Mine waste is generally considered to be the material 

that originates and accumulates at a mine site that has no 

current economic value (Lottermoser, 2010), and it includes 

both solid and liquid waste. Gallium associated with bauxite 

is usually mined by open pit or underground methods and then 

refined through the Bayer process, which can lead to large 

volumes of processing waste. Based on its appearance, bauxite 

(or alumina) processing waste is commonly referred to as 

red mud. Red mud depositories are connected with alumina 

processing plants and are not necessarily located at the mine 

site. For example, four active plants in the United States 

predominantly process alumina ore from overseas (Bray, 

2012). On October 4, 2010, a breached red mud depository 

released between 600,000 to 700,000 cubic meters of material 

from the Ajka Timfoldgyar Zrt alumina plant in Hungary 

(Reeves and others, 2011). This spill was unprecedented 

both in scale (the spill covered an estimated 800 hectares of 

land [Ruyters and others, 2011]) and in its delivery of trace 

elements across such a vast area (Mayes and others, 2011).

Bauxite ore results from extreme chemical weathering, 

and many elements that were originally present in the parent 

rock have been leached away. The exceptions are minor 

elements that can sorb to or coprecipitate with aluminum and 

iron (oxyhydr)oxide minerals that make up the ore and waste. 

For example, in addition to Al and Fe, bauxite processing 

waste has been shown to contain elevated concentrations of 

As, Cr, Ga, Mo, Ni, and V (Mayes and others, 2011). Bauxite 

mine and processing waste generally lacks acid-generating 

minerals, so these elements are expected to remain immobile 

under oxidizing and near-neutral (5<pH<8) weathering 

conditions as long as abundant iron-rich particles are present 

(Smith and Huyck, 1999).

Waste products that result from mining sediment-hosted 

deposits usually consist of waste rock, tailings piles, and 

possibly pit lakes. For example, the cumulative volume of 

tailings generated at the Red Dog Mine, which is a clasticdominated lead-zinc deposit, totaled 27.4 million metric tons 

in 2006, and it is projected to total 88 million metric tons 
H24 Critical Mineral Resources of the United States—Gallium

by 2031 (SRK Consulting [Canada] Inc., 2007). The largest 

U.S. Environmental Protection Agency Superfund site occurs 

in the Tri-State zinc mining district and contains an estimated 

91 million metric tons of waste rock (U.S. Environmental 

Protection Agency, 2006). The district contains a large number 

of Mississippi Valley-type lead-zinc deposits and covers 

approximately 1,800 square kilometers of Kansas, Oklahoma, 

and Missouri (Leach, Taylor, and others, 2010). The 

mineralogy of the mine waste associated with these deposits 

reflects the mineralogy of the unmined deposit, except that the 

proportion of lead-zinc-copper sulfides, such as chalcopyrite, 

galena, and sphalerite, is reduced relative to the gangue 

minerals (barite, calcite, dolomite, pyrite, and quartz) (Kelley 

and others, 1995; Leach, Viets, and others, 1995). Because of 

their association with sulfides and secondary minerals, such 

elements as Ag, Al, As, Au, B, Ba, Cd, Co, Fe, Ga, Ge, Hg, 

Mn, Mo, Ni, and Sb can also be found in clastic-dominated 

deposits (Briskey, 1986; Kelley and others, 1995). Likewise, 

Mississippi Valley-type deposits can contain minor to trace 

amounts of Ag, As, Au, Ba, Bi, Cd, Co, Cu, F, Fe, Ga, Ge, In, 

Mo, Ni, Pb, Sb, Sn, and Tl (Leach, Viets, and others, 1995; 

Foley, 2002a). Under oxidizing and acidic (pH<3) weathering 

conditions, which would be expected in clastic-dominated 

mine waste having little to no acid-neutralizing capacity and 

significant pyrite content, many of these elements are mobile 

(Smith and Huyck, 1999). The acid-neutralizing capacity 

of carbonate-hosted mine waste is greater, however, and 

the oxidizing and circumneutral (5<pH<8) conditions that 

prevail during weathering would cause most of the elements 

listed above to be less mobile or immobile (Smith and Huyck, 

1999). An important exception is zinc, which can make up a 

large percentage of the total dissolved metals draining from 

Mississippi Valley-type deposits (Plumlee and others, 1999). 

A review of the environmental geochemistry of platform 

carbonate-hosted sulfide deposits identifies surface disposal 

of tailings as a potential environmental concern, especially 

if As, Cd, Tl, and Zn are present, because airborne dust may 

contain metal particulates (Foley, 2002a). A study of the 

Austinville lead-zinc mine in Wythe County, Virginia (Foley, 

2002b), shows that residual soils near a Mississippi Valleytype deposit can be acidic enough to influence the mobility of 

metals in surface environments and that this is especially true 

for Mississippi Valley-type deposits that have high jasperoid 

and pyrite content.

Human Health Concerns

Study of the health effects of gallium on humans largely 

focuses on exposure to synthetic gallium arsenide (GaAs) by 

workers in the semiconductor industry, where small GaAs 

particles can be inhaled or ingested (Flora, 2000). The toxicity 

of GaAs may result from the arsenic (because arsenic trioxide 

may be released from GaAs after absorption); gallium toxicity, 

however, though not well studied, has not been ruled out 

(Chitambar, 2010). Humans can tolerate intravenous infusions 

of gallium nitrate during medical treatments for diseases 

such as cancer (Chitambar, 2010), but the long-term effects 

of human exposure to gallium in drinking water and soil 

resources remain unknown. Gallium does not appear in the 

U.S. Clean Water Act, and the Occupational Safety and Health 

Administration (OSHA) does not consider gallium metal to be 

hazardous, although some State governments include gallium 

in their right-to-know lists (Ziegler and others, 2004).

Given their similar chemical characteristics, the health 

effects of gallium may be similar to those of aluminum. 

The Agency for Toxic Substances and Disease Registry 

(ATSDR) provides a useful summary of the human toxicology 

of aluminum (Agency for Toxic Substances and Disease 

Registry, 2008). Exposure to aluminum by the general public 

is most likely to occur through consumption of food, water, 

and aluminum-bearing medicines (such as antacids or antidiarrheal agents). Occupational exposure to aluminum is most 

likely to occur during ore processing and product fabrication. 

Most ingested or inhaled aluminum is not absorbed by the 

body, although absorption can vary depending upon the 

chemical form and particle size. Overexposure to aluminum 

can affect the nervous system, causing impaired performance 

in motor, sensory, and cognitive function. In addition, 

aluminum workers may suffer respiratory effects, such as 

impaired lung function and fibrosis. The United States national 

secondary drinking water regulation for aluminum, which 

is a nonenforceable guideline, ranges from 50 to 200 μg/L 

(U.S. Environmental Protection Agency, 2013b). Most surface 

waters contain aluminum concentrations below 100 μg/L 

(Agency for Toxic Substances and Disease Registry, 2008).

Mining of deposits where gallium is a byproduct can 

potentially mobilize elements that are known human toxins; 

these elements have the potential to affect human health 

when present above threshold concentrations in air, drinking 

water, and soils. These concentrations can be exceeded near 

mine sites. The best known examples are the neurological 

effects of lead on children (Holecy and Mousavi, 2012) and 

the carcinogenic effects of arsenic in drinking water (Gupta 

and others, 2012). The current U.S. National Ambient Air 

Quality Standard for lead is 0.15 micrograms per cubic meter 

(U.S. Environmental Protection Agency, 2013a), and current 

U.S. primary and secondary drinking water standards for lead, 

zinc, and copper are 0, 5, and 1 milligrams per liter (mg/L), 

respectively (U.S. Environmental Protection Agency, 2013b). 

Canadian agricultural soil quality guidelines for lead, zinc, 

and copper are 70, 200, and 63 mg/kg, respectively (Canadian 

Council of Ministers of the Environment, 2007).

Ecological Health Concerns

Relatively few studies focus on the ecological effects of 

gallium mobility in the environment. Fish tend to be sensitive to 

low concentrations of dissolved metals, and they are therefore 

useful indicators of contamination in aquatic systems. One 

of several useful endpoints used in toxicity tests is the lethal 

concentration that leads to 50 percent mortality (LC50 ) 

after exposure to a substance for a certain amount of time. 
Environmental Considerations H25

A study on the acute toxicity of dissolved gallium to carp 

(Cyprinus carpio Linnaeus) revealed a mean LC50 value of 

95.6 ± 14.3 mg/L after 96 hours of exposure (Betoulle and 

others, 2002). Chronic toxicity tests with dissolved gallium 

and developing rainbow trout (Oncorhynchus mykiss)

revealed a mean LC50 value of 3.5 mg/L after 28 days of 

exposure (Birge and others, 1980). In acute toxicity tests with 

saltwater organisms, LC50 values ranged from 12 to 55 mg/L 

of gallium after 24, 48, or 96 hours of exposure (Onikura and 

others, 2005). Gallium is typically not included in the suite 

of elemental concentrations reported in studies of mine waste 

water chemistry, so these LC50 values cannot be compared 

with likely gallium concentrations in mining environments.

The ecological effects of gallium may be similar to those 

of aluminum given their similar behavior in the environment. 

When brown trout (Salmo trutta Linnaeus) were exposed to 

stream water samples of various pH levels and aluminum 

concentrations, the healthiest populations were observed in 

waters with pH above 5.0 and aluminum concentrations below 

20 μg/L; significantly higher aluminum concentrations in gills 

were observed in moribund fish, suggesting that complications 

with respiration contributed to their decline (Andrén and 

Rydin, 2012). In toxicity tests with tropical freshwater algal 

and zooplankton species, LC50 values for aluminum increased 

with increasing dissolved organic carbon concentrations 

(Trenfield and others, 2012), suggesting that the affinity of 

aluminum for complexing with organic carbon renders it less 

toxic. The National Ambient Water Quality Criteria recommended for the protection of aquatic life and adopted by some 

regions in the United States are 750 micrograms of aluminum 

per liter (μg Al/L) of water (acute) and 87 μg Al/L of water 

(chronic) (Suter and Tsao, 1996). Because aluminum is more 

mobile under acidic conditions, plants growing in acidic soils 

may suffer from aluminum toxicity, which is known to inhibit 

root growth and crop yields (Kabata-Pendias and Mukherjee, 

2007). As a result, some regions of the United States have 

adopted a soil screening benchmark of 50 ppm of soil for 

terrestrial plants, as recommended by Oak Ridge National 

Laboratory (Efroymson and others, 1997).

The ecological effects of other elements that gallium is 

associated with are relatively well documented. For example, 

in toxicity tests with aquatic invertebrates incubated with 

mine tailings from Mississippi Valley-type deposits, Besser 

and Rabeni (1987) showed that increased concentrations of 

dissolved cadmium, lead, and zinc correlated with decreased 

survival and growth (Besser and Rabeni, 1987). Likewise, 

various fish species experienced rapid mortality when exposed 

to increasing dissolved zinc concentrations, but this effect 

decreased with increasing calcium, magnesium, and sodium 

concentrations (De Schamphelaere and Janssen, 2004). High 

metal concentrations in soils can also cause phytotoxicity. 

Zinc is a micronutrient for many plant species, but when 

highly concentrated in soils, it can inhibit a plant’s metabolic 

functions and cause deficiencies in other essential nutrients, 

such as copper, iron, manganese, and phosphorus (Nagajyoti 

and others, 2010, and references therein).

Carbon Footprint

Gallium has potential application in the effort to reduce 

the carbon footprint from energy production because of its use 

in efficient thin-film photovoltaic cells (solar panels) (Miles 

and others, 2007). The copper-indium-gallium-(di)selenide 

(CIGS) alloy, among others, is an attractive thin-film product 

because of its high efficiency and relatively low material 

and manufacturing costs (Kaneshiro and others, 2010). In 

a comparative study of lifecycle greenhouse gas emissions 

for various technologies, Kim and others (2012) found that 

estimates for the carbon footprint of CIGS technologies under 

test conditions are about 26 grams of CO2 equivalent per 

kilowatthour (g CO2-equiv/kWh), which compares well with 

older technologies that have greenhouse gas emissions in the 

range of 39 to 110 g CO2-equiv/kWh (Fthenakis and others, 

2008). The CIGS photovoltaic cells are in the initial stages 

of commercialization, but they are expected eventually to 

compete with other forms of energy production as economies 

of scale permit future cost reductions (Miles and others, 2007; 

Rockett, 2010). With the rapid evolution of the thin-film 

photovoltaic cell industry, the recyclability of generated 

products and the potential for leaching of toxic elements from 

improperly disposed products would need to be addressed.

Mine Closure

Most recent and new mining operations include closure 

plans that address issues related to the mine footprint, which 

includes waste left onsite and locally affected soil and water, 

as well as ecological effects, such as habitat destruction and 

loss of biodiversity. Bauxite mines can be particularly prone 

to soil and habitat loss given that many are located in tropical, 

biodiverse locations. A study of a 10-year-old reforested area 

at a closed bauxite mine in Brazil showed that, although the 

reforestation program created adequate recovery conditions, 

recovery of native flora was influenced by the uneven return of 

important seed-dispersing animals (Parrotta and others, 1997). 

At one of the largest bauxite mine operations in the State 

of Western Australia, Australia, the Alcoa World Alumina 

Australia company clears the Jarrah (Eucalyptus marginata) 

forest before developing an open pit mine; after extracting 

bauxite, the company returns the topsoil and spreads Jarrah 

seeds at a rate of approximately 550 hectares per year 

(Gardner and Bell, 2007).

With respect to mining in zinc deposits, mine closure 

issues often relate to the stability of large waste piles and 

tailings dams, and the potential for acid mine drainage from 

the site. Mine waste pile stability is commonly addressed 

through grading and covering dewatered piles. Acid mine 

drainage can be addressed with active water treatment 

facilities, passive limestone-lined channels, or constructed 

wetlands (Plumlee and Logsdon, 1999). The end result of both 

active and passive approaches is precipitation of dissolved 

metals. Precipitated metals in reconstructed wetland systems 

tend to be more stable under the prevailing anoxic conditions, 
H26 Critical Mineral Resources of the United States—Gallium

whereas the metal-rich precipitates that result from active 

treatment facilities form a sludge that can be similar to the 

metal-rich bauxite processing residues. At large mines, mine 

waste is typically consolidated into pits and submerged under 

water, forming a tailings pond. Acid-generating minerals are 

less reactive under water, but seepage is typically monitored 

and, if necessary, treated. At such sites as the Red Dog Mine 

in Alaska, seepage from the tailings ponds may need to be 

treated in perpetuity (U.S. Environmental Protection Agency, 

2010). Failure of tailings pond impoundments can cause 

catastrophic flooding (McDermott and Sibley, 2000). Thus, 

plans for securing waste piles and prevention and treatment 

of acid mine drainage are important aspects of long-term 

active and proposed metal mining projects.

Problems and Future Research

Future research efforts related to gallium resources are 

likely to be directed at understanding the factors that may 

affect future supplies. Such research could include developing 

improved assessment models, conducting experimental 

studies, and seeking new and more efficient extraction and 

recycling technologies.

Assessment models.—Improved models could help in locating 

new and unconventional resources. Efforts focused on 

incorporating additional geochemical and mineralogical data 

into existing mineral deposit and grade tonnage models (for 

zinc and bauxite deposits) can contribute to the development 

of quantitative grade-tonnage models and new approaches to 

evaluating the geologic availability of gallium as a byproduct. 

New models to estimate the occurrence of gallium in polymetallic resources of the Kipushi type could also contribute to 

more accurate quantitative assessments of the geologic availability of gallium. Also, models can assist in the evaluation 

of resources of gallium currently locked up in unprocessed 

mine waste, such as the extensive waste piles left over from 

the mining of Mississippi Valley-type deposits in the midcontinent region of the United States.

Experimental studies.—Studies of gallium at conditions 

applicable to biochemical processes occurring in surface 

environments would provide for a better understanding of the 

natural gallium lifecycle and the potential issues related to 

gallium waste disposal.

New technologies.—Devising new and more efficient extraction technologies for both raw ore and recycled products 

would help prevent or alleviate future bottlenecks in the 

supply chain that may result from increased demand.

Recycling.—Because so little gallium in used products gets 

recycled, greater gallium recycling efficiency in the future 

can reduce reliance on mining of primary gallium resources. 

Future research could be focused on improving gallium 

recycling from current products, as well as future product 

design that allows greater gallium recycling efficiency.

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s_pb_history.html.

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Germanium and Indium

By W.C. Pat Shanks III, Bryn E. Kimball, Amy C. Tolcin, and David E. Guberman

Chapter I of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–I

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter I

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract......................................................................................................................................................... I1

Introduction................................................................................................................................................... I1

Background.......................................................................................................................................... I1

Germanium .................................................................................................................................. I2

Indium .......................................................................................................................................... I2

Uses and Applications ....................................................................................................................... I2

Germanium .................................................................................................................................. I2

Indium .......................................................................................................................................... I4

Demand and Availability of Supply .................................................................................................. I4

Substitutes for Germanium and Indium ................................................................................. I4

Strategic and Critical Resource Issues ................................................................................. I5

Geology.......................................................................................................................................................... I5

Geochemistry....................................................................................................................................... I5

Mineralogy........................................................................................................................................... I5

Deposit Types ...................................................................................................................................... I6

Germanium .................................................................................................................................. I6

Indium ........................................................................................................................................ I11

Mining and Beneficiation Methods ............................................................................................... I11

Resources and Production ....................................................................................................................... I13

Reserves, Other Identified Resources, and Undiscovered Resources ................................... I13

Production.......................................................................................................................................... I13

Prices and Pricing............................................................................................................................. I14

Environmental Considerations ................................................................................................................. I15

Sources and Fate in the Environment ........................................................................................... I15

Mine Waste Characteristics ........................................................................................................... I18

Human Health Concerns .................................................................................................................. I18

Ecological Health Concerns ............................................................................................................ I19

Carbon Footprint................................................................................................................................ I19

Mine Closure...................................................................................................................................... I20

Problems and Future Research ............................................................................................................... I20

References Cited........................................................................................................................................ I20

Suggested citation:

Shanks, W.C.P., III, Kimball, B.E., Tolcin, A.C., and Guberman, D.E., 2017, Germanium and indium, chap. I of Schulz, K.J., 

DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic 

and environmental geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. I1– I26, 

http://dx.doi.org/10.3133/pp1802I.
Germanium and Indium

By W.C. Pat Shanks III, Bryn E. Kimball, Amy C. Tolcin, and David E. Guberman

Abstract

Germanium and indium are two important elements used 

in electronics devices, flat-panel display screens, light-emitting 

diodes, night vision devices, optical fiber, optical lens systems, 

and solar power arrays. Germanium and indium are treated 

together in this chapter because they have similar technological uses and because both are recovered as byproducts, 

mainly from copper and zinc sulfides.

The world’s total production of germanium in 2011 

was estimated to be 118 metric tons. This total comprised 

germanium recovered from zinc concentrates, from fly ash 

residues from coal burning, and from recycled material. 

Worldwide, primary germanium was recovered in Canada 

from zinc concentrates shipped from the United States; in 

China from zinc residues and coal from multiple sources 

in China and elsewhere; in Finland from zinc concentrates 

from the Democratic Republic of the Congo; and in Russia 

from coal.

World production of indium metal was estimated to be 

about 723 metric tons in 2011; more than one-half of the total 

was produced in China. Other leading producers included 

Belgium, Canada, Japan, and the Republic of Korea. These 

five countries accounted for nearly 95 percent of primary 

indium production.

Deposit types that contain significant amounts of 

germanium include volcanogenic massive sulfide (VMS) 

deposits, sedimentary exhalative (SEDEX) deposits, Mississippi Valley-type (MVT) lead-zinc deposits (including 

Irish-type zinc-lead deposits), Kipushi-type zinc-lead-copper 

replacement bodies in carbonate rocks, and coal deposits.

More than one-half of the byproduct indium in the 

world is produced in southern China from VMS and 

SEDEX deposits, and much of the remainder is produced 

from zinc concentrates from MVT deposits. The Laochang 

deposit in Yunnan Province, China, and the VMS deposits 

of the Murchison greenstone belt in Limpopo Province, 

South Africa, provide excellent examples of indium-enriched 

deposits. The SEDEX deposits at Bainiuchang, China (located 

in southeastern Yunnan Province), and the Dabaoshan 

SEDEX deposit (located in the Nanling region of China) 

contain indium-enriched sphalerite. Another major potential 

source of indium occurs in the polymetallic tin-tungsten belt 

in the Eastern Cordillera of the Andes Mountains of Bolivia. 

Deposits there occur as dense arrays of narrow, elongate, 

indium-enriched tin oxide-polymetallic sulfide veins in 

volcanic rocks and porphyry stocks.

Information about the behavior of germanium and 

indium in the environment is limited. In surface weathering 

environments, germanium and indium may dissolve from 

host minerals and form complexes with chloride, fluoride, 

hydroxide, organic matter, phosphate, or sulfate compounds. 

The tendency for germanium and indium to be dissolved and 

transported largely depends upon the pH and temperature 

of the weathering solutions. Because both elements are 

commonly concentrated in sulfide minerals, they can be 

expected to be relatively mobile in acid mine drainage where 

oxidative dissolution of sulfide minerals releases metals and 

sulfuric acid, resulting in acidic pH values that allow higher 

concentrations of metals to be dissolved into solution.

Introduction

Background

This chapter provides an update of the U.S. Geological 

Survey (USGS) assessment of germanium and indium 

resources published in the early 1970s (Weeks, 1973); 

developments in shorter-term resource availability are 

addressed in the periodical USGS minerals information reports 

(part of the U.S. Bureau of Mines from 1925 through 1995). 

Germanium and indium are in significant demand for hightechnology applications, and this has led to many studies of 

concentrations and distributions in primary mineral deposits, 

advances in extraction technology for these byproduct 

elements, increased attention on recycling, a flourishing field 

developing new applications for these elements and their 

compounds, and consideration of the impacts that increased 

consumption could have on the environment.
I2 Critical Mineral Resources of the United States—Germanium and Indium

Germanium

In 1864, English chemist John Newlands postulated 

the existence of an element that would be intermediate 

between silicon and tin in his scheme of elemental octaves 

(Newlands, 1864). In 1871, the salient properties of the 

undiscovered element were predicted by the Russian chemist 

Dmitri Mendeleev (Weeks, 1932), who called it ekasilicon 

(beyond silicon). In 1886, the element was isolated from the 

silver sulfide mineral argyrodite (Ag8GeS6) and described by 

the German chemist Clemens Winkler (1886), who named 

it germanium, after his native country. More than one-half 

century elapsed after the isolation of germanium before its first 

commercial use in diodes and transistors was developed.

Germanium is a hard, brittle, grayish-white, semiconducting element that has electrical properties between 

those of a metal and an insulator. Germanium is metallic in 

appearance and has unique properties that make it critical to 

the function of numerous commercial, industrial, and military 

applications. It is consumed as a pure element or in compound 

form, depending on the application.

Indium

Indium was discovered in 1863 by German chemists 

Ferdinand Reich and H.T. Richter during a spectrographic 

analysis of sphalerite ore samples from Freiberg, Germany 

(Reich and Richter, 1863). They named the element indium 

after the distinctive indigo-blue line in its emission spectrum. 

Indium remained only a scientific curiosity for years following 

its discovery; little was known about its occurrence apart 

from the Freiberg ore.

Indium is a soft, lustrous, silvery-white metal with 

a low melting point relative to other metals. It is ductile 

and malleable, even at cryogenic temperatures; it has high 

wettability when melted and is resistant to thermal fatigue.

Uses and Applications

Germanium

Multiple stages of the germanium production process 

yield germanium compounds and metals that are designed for 

specific applications. Concentrated germanium is chlorinated 

and distilled to form the first usable product in the refining 

process, germanium tetrachloride (GeCl4 ), which is a colorless 

liquid that is used primarily as a reagent in the production of 

fiber-optic cable. Germanium tetrachloride can be hydrolyzed 

and dried to produce germanium dioxide (GeO2 ), which is 

another commonly used compound. Germanium dioxide is 

a white powder that is used to manufacture certain types of 

optical lenses and as a catalyst in the production of plastic 

polyethylene terephthalate (PET) resin.

Germanium dioxide can be reduced with hydrogen to 

produce a germanium metal powder, which is subsequently 

melted and cast into first-reduction bars. The germanium 

bars are then zone-refined (a process that involves melting 

and cooling germanium bars to isolate and remove impurities 

and ultimately yield extremely pure germanium) to produce 

electronic-grade germanium metal. Zone-refined germanium 

metal can then be grown into crystals and sliced for use as 

semiconductors or recast into forms suitable for lenses or 

window blanks in infrared optical devices (Bleiwas, 2010).

Germanium owes its usefulness to at least five salient 

properties. First, it is an intrinsic semiconductor, which in the 

pure state—that is, in the absence of contaminant elements in 

its crystal structure—will conduct electricity, albeit poorly. 

It is particularly effective as a conductor at high frequencies 

and low operating voltages. Second, germanium is transparent 

to part of the infrared electromagnetic spectrum, whether in 

the crystalline or glassy states. Third, like silicon, germanium 

is a glass-former, able to form extended three-dimensional 

networks of more or less randomly ordered germaniumoxygen tetrahedra. Fourth, it has an exceptionally high 

refractive index. Fifth, it exhibits low chromatic dispersion. 

These five properties, singly and sometimes in combination, 

determine the usefulness of germanium in electronics devices, 

night vision devices, fiber-optic cable, optical lens systems, 

and solar power arrays. A sixth property, of specific value 

to plastic bottle and container manufacturing—a single but 

commercially important use—is the ability of germanium to 

catalyze the polymerization of the plastic PET resin without 

undesirable coloring of the plastic product.

Germanium is relatively strong and is easily machined 

into infrared lenses and windows for infrared optical 

devices. Infrared imaging devices are used extensively 

by the military and law-enforcement agencies for surveillance, reconnaissance, and target-acquisition applications. 

Infrared optical devices improve a soldier’s ability to operate 

weapon systems in harsh conditions effectively, and they are 

increasingly used in remotely operated unmanned weapons 

and aircraft. Infrared optical devices are also used for border 

patrol and by emergency response teams for conducting 

search-and-rescue operations.

Germanium substrates are used to form the base layer in 

multijunction solar cells, which are the highest efficiency solar 

cells currently available. The substrates are produced from 

high-purity, single-crystal germanium ingots that are sliced 

into wafers and polished. Ultrathin layer combinations of 

materials are “grown” on top of the germanium substrate, and 

each layer captures a specific part of the solar spectrum and 

converts it into electricity. These multijunction cells are the 

preferred type for use in space-based solar power applications 

because of their high energy-conversion efficiency and 

strength at minimal size (fig. I1).

Terrestrial-based photovoltaic installations are a potential 

growth area for germanium use (Bleiwas, 2010). Solar 
Introduction I3

powerplants that use concentrator technology composed of 

lenses or mirrors that focus high concentrations of direct 

sunlight onto germanium-based multijunction solar cells 

have emerged as viable sources for large-scale renewable 

energy generation. Germanium substrates are also used in 

high-brightness light-emitting diodes (LEDs) for backlighting 

liquid crystal display televisions and in vehicle headlights 

and tail lights.

Germanium tetrachloride is used to manufacture fiberoptic cables. Germanium is added to the pure silica glass core 

of the fiber-optic cable to increase its refractive index and 

minimize signal loss over long distances. Global demand for 

fiber-optic cable has increased significantly in recent years 

as technological advances have required ever-increasing 

bandwidth to transmit and receive data.

The major end uses for germanium worldwide are estimated 

to be fiber-optic systems (30 percent), infrared optics (25 percent), 

polymerization catalysts (25 percent), electronics and solar 

electric applications (15 percent), and other uses, including 

for chemotherapy, metallurgy, and phosphors (5 percent) 

(fig. I2A). The domestic end-use distribution is different and 

is estimated to be infrared optics (50 percent), fiber-optic 

systems (30 percent), electronics and solar electric applications 

(15 percent), and other uses, including for chemotherapy, metallurgy, and phosphors (5 percent) (Guberman, 2013a). Germanium 

is not used in polymerization catalysts in the United States.

Figure I1. Photograph of a concentrator photovoltaic 

(CPV) solar power system which, in 2013, earned a National 

Renewable Energy Laboratory (NREL) outdoor efficiency 

rating of 34.9 percent under international standard operating 

test conditions for concentrator photovoltaics—a new world 

record. Solar powerplants such as this that use lenses or 

mirrors to focus high concentrations of direct sunlight onto 

germanium-based multijunction solar cells have emerged as 

viable sources for large-scale renewable energy generation 

(Bleiwas, 2010; BusinessWire, 2013). Photograph courtesy of 

Arzon Solar, LLC.

A. Germanium 

Figure Germanium and Indium I2

B. Indium 

Other

5%

Other uses

10%

Fiber-optic

systems

30%

Infrared

optics

25%

Polymerization

catalysts

25%

Electronics and

solar electric

applications

15%

Flat-panel

display screens

and touchscreens 

65% Low-melting alloys

and solders

9%

Semiconductors

9%

Transparent

heat reflectors

7%

Figure I2. Pie charts showing major end uses of A, germanium and B, indium as a percentage of world consumption in 2012. 

Data are from Guberman (2013a) and Tolcin (2013a).
I4 Critical Mineral Resources of the United States—Germanium and Indium

Indium

Indium metal with purities of up to 99.99995 percent is 

converted to different shapes and forms that have different 

uses. High-purity indium metal ingots are commonly used 

to produce indium compounds; indium shot is used in vapor 

deposition and plating applications; and indium wire, ribbon, 

and foil are typically used for soldering.

Indium is consumed mostly (65 percent) for the production of tin-doped indium oxide (or indium-tin oxide [ITO]), 

which is a transparent conducting oxide used in virtually every 

flat-panel display screen and touchscreen (figs. I2B and I3). 

ITO is typically deposited as a thin-film coating on the display 

surface, where it transforms incoming electrical data into an 

optical form. Most ITO production is concentrated in Japan, 

although significant quantities are also produced in China, 

the Republic of Korea, and Taiwan.

Alloys and solders are the second-ranked end use of 

indium globally (9 percent). Indium-containing solders have 

lower crack propagation and improved resistance to thermal 

fatigue compared with tin-lead solders. Indium-containing 

solders also inhibit the leaching of gold components in 

electronic apparatus. Certain types of indium alloys can be 

used as bonding agents between nonmetallic materials, such 

as glass, glazed ceramics, and quartz. Indium also is used in 

dental alloys and in white-gold alloys. Other indium alloys 

have been used as a substitute for mercury and in nuclear 

control rods.

Another important use of indium is in semiconductor 

materials for LEDs and laser diodes (9 percent). Indium 

antimonide, indium arsenide, or indium phosphide can be used 

as the substrate for indium-based semiconductors, and several 

indium-containing compounds can be used as the epitaxial 

layer (or substrate coating), such as indium-gallium-arsenide. 

Indium-based LEDs are used predominantly to transmit data 

optically and, to a lesser extent, in LED displays. Indiumbased laser diodes are used in fiber-optic communications.

Demand and Availability of Supply

Substitutes for Germanium and Indium

Germanium.—Silicon can be a less-expensive substitute 

for germanium in certain electronic applications. Some 

metallic compounds can be substituted in high-frequency 

electronics applications and in some LED applications. Zinc 

selenide and germanium glass substitute for germanium metal 

in infrared applications systems but commonly at the expense 

of performance. Titanium has the potential to be a substitute as 

a polymerization catalyst.

Indium.—The indium-containing compound ITO, which is 

used as the transparent conducting film in many flat-panel display 

screens and touchscreens, has few substitutes. Various groups, 

however, have recently explored using antimony tin oxide, carbon 

nanotube coatings, and a number of other compounds to substitute 

for ITO in flat-panel and flexible display screens (Tolcin, 2013a).

Figure I3. Photographs showing A, indium-tin oxide (ITO), which is a transparent conducting oxide, and B, 

examples of flat-panel display screens and touchscreens. ITO is typically deposited as a thin-film coating on the 

display surface where it transforms incoming electrical data into an optical form. Photograph A courtesy of Indium 

Corp.; photograph B courtesy of CI Systems, Inc.
Geology I5

Strategic and Critical Resource Issues

Germanium.—The extensive use of germanium for 

military and commercial applications has made it a critical 

material in the United States and the rest of the world. 

Germanium was included in the National Defense Stockpile 

in 1984 and was more recently added to a stockpile program 

in China. In 2010, the European Union included germanium 

in a list of raw materials of critical concern for its member 

countries owing to its expected economic importance and 

relative supply risk (European Commission, 2010).

Few adequate substitutes exist for germanium for 

applications that are important for defense and law enforcement, such as infrared optics and substrates for solar cells 

in satellites. Most germanium production is concentrated in 

a few countries—Canada, China, Finland, and Russia. The 

United States is dependent on imports for its germanium 

consumption. As a byproduct metal, the supply of germanium 

is heavily reliant on zinc production. It has been estimated 

that less than 5 percent of the germanium contained in zinc 

concentrates reaches refineries that are capable of extracting 

and producing germanium.

Indium.—Indium can be considered a critical material for 

display technology because there are few substitutes. Because 

indium is recovered as a byproduct of zinc production, the 

supply of primary indium is determined by the supply of zinc, 

regardless of the market demand for indium. Additionally, a 

large portion of the indium contained in zinc ores and concentrates is not recovered—most zinc smelters are not equipped 

to extract indium. At the few smelters that do include indiumprocessing circuits, the average indium recovery rate is only 

about 50 percent (ranging from 30 to 80 percent) (Jorgenson 

and George, 2004). Increased consumption of indium is 

expected to be satisfied by increased recycling and additional 

primary supply through improved recovery rates, the construction of new plants, and expansions at existing recovery circuits 

(Alfantazi and Moskalyk, 2003).

Geology

Geochemistry

Germanium.—Germanium has a crustal abundance of 

1.6 parts per million (ppm), making it about the 50th most 

abundant element. In elemental form, germanium is a hard, 

gray-white metalloid with an atomic number of 32. It is a 

Group 4A element in the periodic table and has properties 

similar to its neighbors silicon and tin. Germanium most 

commonly exists in the +4 and +2 valence states and 

commonly substitutes for silicon or tin.

Aqueous germanium (as Ge4+ ) forms complexes with 

fluoride, hydroxide, phosphate, and sulfate in low-temperature environments. At hydrothermal temperatures (up to 

300 degrees Celsius [°C]), the principal complex is germanic 

acid (H4GeO4 ), but germanium-fluoride complexes may be 

important at high-fluorine concentrations (Wood and Samson, 

2006). Germanium concentrations of up to about 300 parts per 

billion (ppb) have been reported for some geothermal waters 

(Höll and others, 2007). Further evidence for hydrothermal 

transport comes from germanium concentrations of up to 

20 ppm in hydrothermal sinter deposits in New Zealand 

geothermal systems (Krupp and Seward, 1987) and of up 

to about 270 ppm in modern sea-floor hydrothermal sulfide 

deposits (Bischoff and others, 1983).

Indium.—Indium has an average crustal abundance 

of about 49 ppb and is the 61st most abundant element. 

Elemental indium is very soft, lustrous, metallic, and 

malleable, and it has a very low melting point of 156.6 °C. It 

is a Group 3A element in the periodic table and has chemical 

properties similar to its group neighbors gallium and thallium. 

Indium most commonly occurs in the +3 valence state.

Schwarz-Schampera and Herzig (1999) reported indium 

concentrations of up to 590 ppm in modern sea-floor hydrothermal sulfide deposits of the Valu Fa Ridge in the Lau Basin 

of the western Pacific Ocean, so indium clearly is transported 

and deposited in hydrothermal systems. Seward and others 

(2000) experimentally determined that indium is complexed by 

chloride in aqueous solutions at temperatures of up to 350 °C.

Chaplygin and others (2007) have shown that abundant 

indium— of up to 4.75 weight percent in sphalerite—occurs 

at fumaroles of the Kudriavy volcano on Iturup Island in the

Kuril Islands, Russia. The fumarole deposits form at 400 to 750°C,

showing that vapor transport in high-temperature systems is 

also possible.

Mineralogy

Germanium.—Germanium can occur as the 

following rare minerals (with approximate compositions): 

argyrodite (Ag8GeS6 ), germanite (Cu13Fe2Ge2S16 ), renierite 

((Cu,Zn)11(Ge,As)2Fe4S16 ), or briartite (Cu2(Fe,Zn)GeS4 ). 

Germanium is primarily recovered as a byproduct from zinc, 

silver, lead, and copper ores, however. Laser ablation and 

inductively coupled plasma-mass spectrometry (ICP–MS) 

traverses across minerals by Ye and others (2011) have 

shown that germanium occurs in true solid solution in sulfide 

minerals and is most strongly enriched in sphalerite from 

Mississippi Valley-type (MVT) deposits, which are stratabound, carbonate-hosted lead-zinc deposits that form epigenetically from basinal brines at low to moderate temperatures 

(typically 150 to 225 °C).

Indium.—Indium forms a few minerals, such as 

dzhalindite (In(OH)3 ) and indite (Fe+2In2

S4 ), but is not found 

concentrated into significant deposits. Indium is primarily 

produced as a byproduct from zinc ores. Ye and others (2011) 

provide evidence that indium, like germanium, is in true solid 

solution in sphalerite.
I6 Critical Mineral Resources of the United States—Germanium and Indium

Deposit Types

Germanium

Germanium does not form specific deposits, but rather 

occurs as a byproduct in a variety of deposit types that contain 

copper, gold, lead, silver, and zinc. Germanium concentrations 

in sphalerite from these deposits are typically a few hundred 

parts per million. Because byproduct production comes from 

metallurgical operations that are commonly fed by concentrates from any number of different deposits and locations, it 

is difficult to track germanium production back to a specific 

deposit. The example deposits discussed below, however, are 

known to be significant contributors to major germaniumproducing facilities.

Types of deposits that contain significant germanium 

include volcanogenic massive sulfide (VMS) deposits, 

sedimentary exhalative (SEDEX) deposits, Mississippi Valleytype (MVT) lead-zinc deposits (including Irish-type lead-zinc 

deposits), and Kipushi-type zinc-lead-copper replacement 

bodies in carbonate rocks (table I1; Höll and others, 2007). 

Germanium is most enriched in the Kipushi-type deposits, 

but worldwide production is mostly from low-temperature 

stratiform sphalerite deposits (where mineralization follows 

stratigraphic layering) and strata-bound sphalerite deposits 

(where mineralization may cross-cut strata but is restricted to 

a particular stratigraphic unit).

Sedimentary exhalative deposits.—The Red Dog zinclead-silver mining district (fig. I4) in the western Brooks 

Range of northern Alaska contains several zinc-lead-silver 

sulfide bodies and prospects separated into thrust fault slices 

(Johnson and others, 2015). The four primary deposits of the 

Red Dog Mine have reserves and resources totaling 141 million 

metric tons at grades of 16.6 weight percent zinc and 4.6 weight 

percent lead (Jennings and King, 2002). The deposits are 

hosted in the Mississippian Kuna Formation black shales and 

mudstones that were deposited in a closed extensional basin that 

experienced euxinic conditions (Johnson and others, 2015).

Kelley and others (2004) showed that the deposits in the 

Red Dog district have a complex paragenetic history that they 

divide into four stages: (1) shallow subsurface impregnation of 

unconsolidated muds with abundant barite (BaSO4 ) and early 

brown sphalerite, (2) subsea-floor deposition of yellow-brown 

sphalerite with barite, (3) deposition of red-brown sphalerite, 

and (4) faulting and tectonic brecciation accompanied by 

deposition of late tan sphalerite. Kelley and others (2004) 

Table I1. Classification of deposits that host germanium and indium resources.

[Elements: Ag, silver; As, arsenic; Au, gold; Ba, barium; Be, beryllium; Bi, bismuth; Cd, cadmium; Co, cobalt; Cr, chromium; Cu, copper; Fe, iron; Ga, gallium;

Ge, germanium; Hg, mercury; In, indium; Mn, manganese; Mo, molybdenum; Ni, nickel; Pb, lead; PGE, platinum-group elements; Re, rhenium; Sb, antimony;

Se, selenium; Sn, tin; Te, tellurium; Tl, thallium; U, uranium; W, tungsten; Zn, zinc]

Deposit type Short name Characteristics Metals present

(and source of information)

Volcanogenic massive 

sulfide deposits

VMS Stratiform Cu-Pb-Zn sulfide deposits in submarine 

mafic to felsic volcanic terranes that form at the 

sea floor by venting of hydrothermal fluids driven 

by magmatic heat and volatiles

Cu-Zn-Pb-Au-Ag with Be, Bi, Cd, 

Co, Cr, Ga, Ge, Hg, In, Mn, Mo, 

Ni, Se, Sn, Te, and PGE (Shanks 

and Thurston, 2010)

Sedimentary exhalative 

deposits

SEDEX Stratiform, often finely laminated, Zn-Pb-Ag sulfide 

deposits in carbonaceous and pyrite shales and 

siltstones that form by exhalation of hydrothermal 

fluid on the sea floor without any direct igneous 

association

Zn-Pb-Ag with As, Bi, Cd, Co, 

Cu, Ga, Ge, Hg, In, Mn, Ni, Sb, 

Se, Sn, and Tl (Goodfellow and 

Lydon, 2007; Kelley and others, 

2004; Slack and others, 2004)

Mississippi Valley-type 

Zn-Pb deposits

MVT Strata-bound epigenetic Zn-Pb sulfide replacement 

deposits along faults and permeable zones and openspace fill in solution collapse breccias in dolostone 

and limestone. Not associated with igneous activity

Zn-Pb with Cu, Ni, and Co (Leach 

and others, 2005)

Kipushi-type Zn-Pb-Cu 

replacement bodies in 

carbonate rocks

Carbonate 

replacement 

deposits

(CRD)

Strata-bound epigenetic Zn-Pb-Cu sulfide ore 

deposits in irregular pipe-like bodies associated 

with collapse breccias and faults as well as lenticular 

bodies subparallel to the bedding. Association with 

igneous activity is unclear

Zn-Pb-Cu with Ag, As, Bi, Cd, Co, 

Ga, Ge, Mo, Re, Sb, Sn, and W 

(Kampunzu and others, 2009)

Polymetallic Zn-Sn 

vein and fissurefilling deposits 

Polymetallic 

vein

deposits

Polymetallic sulfide veins forming dense arrays in 

volcanic rocks, eroded volcanic-intrusive complexes, 

porphyry stocks, and associated breccia

Zn-Cu-Pb-Sn-In with Ag, As, Bi, 

Co, Ga, Mn, Ni, Sb, Se, and W 

(Ishihara and others, 2011)

Coal and lignite 

deposits

None Ge concentrated in organic matter of coal seams that 

have been affected by hydrothermal activity and 

are generally associated with siliceous layers

Ge in coal with As, Ba, Sb, U, 

and W
Geology I7

Anarraaq

deposit

Wulik

prospect

Oak showing

Su-Lik deposit

Red Dog

deposits

A

A'

Brooks Range

EXPLANATION

Kelly River allochthon

Line of section shown in B

Zn-Pb-Ag±barite deposit

Sulfide showing or prospect

Endicott Mountains allochthon

Thrust fault—D, downthrown

 side; U, upthrown side

Key Creek plate

Red Dog plate

Wolverine Creek plate

A A'

163°10'

68°15'

68°00'

163°00' 162°50'

Map area

ALASKA

Black

mudstone

WRENCH CREEK FAULT

Wulik

Peaks

Deadlock

Mountain

Rok

0 Hills

0 4 8 KILOMETERS

4 8 MILES

U

D

Carbonate turbidites

Varicolored shale and chert

Siltstone, sandstone, conglomerate

Black mudstone Barite

A

A

B

A'

NORTHWEST SOUTHEAST

Devonian

Mississippian

PennsylvanianPermianEtivluk

Group

Siksikpuk

Formation

Kuna

Formation

Kayak

Shale

Kanayut

Conglomerate

Lisburne

Group

Endicott

Group

Su-Lik Anarraaq

Red Dog

Figure I4

Alkaline 

intrusives

VERTICAL EXAGGERATION ×20

Figure I4. Map and schematic cross section showing the geology of the Red Dog mining 

district in Alaska and the stratigraphy of selected deposits in the district. A, Simplified 

geology of the Red Dog district in the western Brooks Range, Alaska, and the locations 

of major zinc-lead-silver (Zn-Pb-Ag) ± barite deposits. The exposure is distal to known 

sulfide deposits, but sulfide showings have been identified nearby. Modified from Johnson 

and others (2015), De Vera and others (2004), and Young (2004). B, Stratigraphic positions 

of Zn-Pb-Ag sulfide deposits and barite deposits in the Red Dog district. This geologic 

section has Jurassic-Cretaceous thrust faulting removed, so the structural plates above 

the Red Dog plate are not shown and originally underlying strata have been restored. 

Modified from Johnson and others (2015) and Kelley and Jennings (2004).
I8 Critical Mineral Resources of the United States—Germanium and Indium

further show that sphalerite from all four stages is enriched 

in germanium and that the average concentrations range 

from 104 to 249 ppm. The highest germanium values occur 

in low-iron, lower temperature, late tan sphalerite. Most of 

the germanium produced at metallurgical facilities at Trail, 

British Columbia, Canada, likely comes from sphalerite 

concentrates from the Red Dog district.

The Fankou lead-zinc deposit in Guandong Province, 

China, is a carbonate-hosted strata-bound deposit hosted 

in Devonian to Carboniferous carbonates and shales (Höll 

and others, 2007). Gu and others (2007) cite evidence for 

the classification of the Fankou deposit as SEDEX with 

both stratiform and cross-cutting mineralization. Since 

mining began in the 1960s, the Fankou Mine has produced 

150,000 metric tons of lead-zinc sulfide containing about 

15 weight percent lead and zinc combined; the sulfide is 

also enriched in sphalerite that contains from 30 to 170 ppm 

germanium (Xuexin, 1984; Höll and others, 2007).

Irish-type carbonate-hosted lead-zinc deposits have 

characteristics of both MVT and SEDEX deposits. The 

Silvermines–Lisheen group of deposits in the Midlands basin 

of County Tipperary, Ireland, formed by replacement of 

lower Carboniferous carbonate sediments near the sea floor 

and represent potentially important deposits of germanium. 

Wilkinson and others (2005) reported that microprobe 

analysis of minerals from drill core samples of the Lisheen 

deposit determined the germanium concentrations to be 

400 to 900 ppm in sphalerite, 200 to 1,300 ppm in galena, 

and 200 to 1,000 ppm in tennantite ((Cu,Fe)12As4S13 ). 

Despite these favorable reports, no production of byproduct 

germanium was reported by Vedanta Resources plc of India, 

which owned the Lisheen deposit.

Mississippi Valley-type deposits.—The GordonsvilleElmwood zinc-lead district in Tennessee hosts examples 

of zinc-rich MVT deposits that formed in collapse breccias 

related to large cavernous underground areas; these deposits, 

on average, have grades of about 400 ppm germanium 

in zinc ore concentrate (Misra and others, 1996). Other 

MVT deposits in the United States average about 50 ppm 

germanium in sphalerite. The Gordonsville Mine near 

Nashville, Tennessee, has produced 45,000 metric tons per 

year of zinc, and byproduct germanium production has been 

significant. In 2010, Nyrstar N.V. of Switzerland reopened 

the zinc mines in the Carthage mineral district in Tennessee, 

which includes the Elmwood, Gordonsville, and Cumberland 

Mines (Nyrstar N.V., 2016). The company reported production 

of 109,000 metric tons of zinc concentrate from these mines 

in 2012; the concentrates were processed at Nyrstar’s Clarksville, Tenn., roast leach electrowin smelter complex, which 

produced a germanium concentrate as well as cadmium metal, 

copper sulfate, sulfuric acid, synthetic gypsum, and zinc metal 

(Nyrstar N.V., 2013, p. 6, 23).

The Huize MVT deposit, which is located in Carboniferous dolomites and limestones of China’s eastern Yunnan 

Province (fig. I5), is one of the largest MVT deposits in 

China. The Huize Mine produces, in order of abundance, 

zinc-lead and byproduct silver, germanium, and cadmium 

(fig. I5; Ye and others, 2011).

Kipushi-type Zn-Pb-Cu replacement bodies in carbonate 

rock.—The most significant carbonate-hosted zinc-leadcopper deposits that contain notable amounts of germanium 

are the Kipushi deposit in the Democratic Republic of the 

Congo (Congo [Kinshasa]) and the Kabwe deposits in 

Zambia (fig. I6; Kampunzu and others, 2009). Kipushi-type 

deposits are strata-bound irregular pipelike sulfide bodies in 

platform carbonate rocks. Collapse breccias, faulting, and 

stratiform lensoidal sulfide bodies are commonly associated 

with Kipushi-type deposits. The Kipushi deposit is hosted in 

the upper Katangan Kundelungu Group (760 to 565 megaannum [Ma]) that consists of alternating carbonate rocks, 

shales, and sandstones. Schneider and others (2007) report 

concordant ages of 451.1± 6.0 Ma and 450.5 ± 3.4 Ma from 

direct rubidium-strontium and rhenium-osmium isochron 

dating, respectively, of ore-stage zinc-copper-germanium 

sulfides. Further, this age of ore formation indicates 

that mineralization is related to crustal fluids related to 

Ordovician extension.

Germanium averages 68 ppm in bulk samples in the 

Kipushi deposit and occurs substituted in sulfide minerals, 

although it sometimes occurs in separate copper-irongermanium sulfide minerals. The actual amount of byproduct 

germanium production from mining is unknown.

Coal and lignite deposits.—China is a major producer 

of germanium from coal. The Lincang lignite mine, which 

is located close to Lincang City in Yunnan Province, is 

a significant source of germanium; the mine produces 

16 metric tons of high-grade germanium dioxide (GeO2 ) 

annually, of which 90 percent is exported (fig. I5; Höll and 

others, 2007). Germanium is produced at the Lincang Mine 

from a lower lignite unit that is close to the contact with 

germanium-rich granitic rocks below. Hu and others (2009) 

report that the Lincang deposit contains 1,000 metric tons or 

more of germanium at a grade of 850 ppm. Germanium-rich 

coal seams are interbedded with siliceous rocks, including 

siliceous limestones, whereas other nearby seams that do 

not have the association with siliceous rocks are barren 

of germanium. The siliceous rocks have oxygen and 

carbon isotope characteristics that suggest a hydrothermal 

origin. Hu and others (2009) propose that hydrothermal 

fluids first leached germanium from granitic rocks. The 

fluids were then discharged as hot springs along fault 

zones into Miocene basins where the germanium was 

concentrated in lignite seams within stratiform siliceous 

and siliceous-limestone deposits.
Geology I9

West R

Beijing

0 100 200 300 400 KILOMETERS

0 100 200 MILES

Yellow

Yangtze

River

River

Yangtze River

Niujiaotang

Bainiuchang

Huize

Kunming

Guiyang

Dabaoshan

Fankou

Jiangshao paleo-suture

Guangzhou

South China craton

Yuebei

 basin

Sanjiang

fold belt

system

Cathaysian terrane

Yunnan

Sn-polymetallic

metallogenic belt

Xiangxi-Qiandong Zn-Pb

metallogenic belt

Jinshajiang-Ailaoshan fault

Jinding

Hetaoping

Figure Germanium and Indium I5. 

CHINA

Sichuan-Yunnan-Guizhou

metallogenic region

Map

area

Skarn deposit

VMS

SEDEX

MVT

Sandstone-hosted deposit

Coal (lignite)

<1

Ore tonnage, in 

 million metric tons

Deposit type

>200

1 to 5

EXPLANATION

Mengxing

Luziyuan

Lincang

Laochang

Simao

30° N

100° E 105° E 110° E 115° E

25° N

20° N

Figure I5. Map showing the locations and geologic 

settings of selected volcanogenic massive sulfide (VMS), 

sedimentary exhalative (SEDEX), Mississippi Valley-type 

(MVT), and coal (lignite) deposits and other types of deposits 

in southern China. Modified from Ye and others (2011). 

Pb, lead; Sn, tin; Zn, zinc
I10 Critical Mineral Resources of the United States—Germanium and Indium

22

21

20

2

3 1

6

10

11

4 12

7 8

9

14

13

18 19 5

16 15

17

0 100 200 300 KILOMETERS

0 50 100 MILES

Bangweulu

block

Congo craton Irumide belt

Irumide belt

Ubendian belt

Zambezi belt

Zimbabwe

craton

DEMOCRATIC REPUBLIC

 OF THE CONGO

ZAMBIA

TANZANIA

MOZAMBIQUE

ZIMBABWE

MALAWI

Lufilian arc

2,050–1,850 Ma

Figure Germanium and Indium I6. 

Zn-Pb

Zn

Pb

Deposit type

EXPLANATION

Age and type of rock

Paleozoic to recent

Neoproterozoic to Paleozoic granitoids

Neoproterozoic metasediments

Mesoproterozoic quartzite and schist

Paleoproterozoic granite gneiss

Archean craton

AFRICA

Map area

Kipushi

deposit

Kabwe

deposits

10° S

15° S

25° E 30° E

Figure I6. Map showing the location of Kipushi-type deposits (including the Kabwe deposits) and major Neoproterozoic 

orogenic belts and basins in the Precambrian tectonic framework of southern Africa. Modified from Kampunzu and others 

(2009). Pb, lead; Zn, zinc
Geology I11

Indium

Approximately one-half of the byproduct indium in the 

world is produced at smelters located in southern China, but 

the Republic of Korea, Japan, Canada, Belgium, and Peru 

(in order of output) also produce significant byproduct indium. 

More than one-half of the material for the smelters in China 

comes from VMS and SEDEX deposits, and much of the 

remaining production comes from MVT deposits. It is difficult 

to decipher which VMS, SEDEX, and MVT deposits in the 

world provide the zinc feedstock processed at the Chinese 

smelters, however.

Volcanogenic massive sulfide deposits.—The Laochang 

deposit in Yunnan Province, China (fig. I5), and the VMS 

deposits of the Murchison greenstone belt in Limpopo Province, South Africa, provide excellent examples of indiumenriched VMS deposits. The Laochang deposit, which is 

hosted by a lower Carboniferous volcano-sedimentary 

sequence, occurs at the junction between southeast- and 

northwest-trending basement faults. Sphalerite samples 

analyzed by Ye and others (2011) have indium contents of 

up to 544 ppm but average about 200 ppm.

The dozen or so VMS deposits of the Archean Murchison 

greenstone belt in South Africa occur in a felsic volcanic 

back-arc basin that was accreted to a continental margin 

(Schwarz-Schampera and others, 2010; Zeh and others, 2013). 

The VMS deposits typically consist of lower temperature 

zinc-rich sea-floor sulfide mineralization that, in some cases, is 

overprinted by higher temperature copper-rich mineralization. 

Indium is closely associated with copper-rich ores, and indium 

concentrations range from 24 to 641 ppm in the Maranda J, 

Romotshidi, and Solomons deposits of the Murchison greenstone belt (Schwarz-Schampera and others, 2010).

Sedimentary exhalative deposits.—The SEDEX deposit 

at Bainiuchang, China, is located in southeastern Yunnan 

Province (fig. I5). Mineralization occurs in stratiform massive 

sulfide lenses in mid-Cambrian dolomite, limestone, and 

sandstone. Indium in sphalerite at Bainiuchang averages 

71 ppm but shows large fluctuations, from 4 to 262 ppm 

(Ye and others, 2011).

The Dabaoshan SEDEX deposit is located in the Nanling 

region of China and occurs at the contact between Devonian 

argillaceous limestones and underlying sandstones and shales 

(fig. I5). The deposit contains about 40 lenses of zinc-lead 

ore, and pyritic copper zones occur on the margins of some 

lenses. Indium in sphalerite of the Dabaoshan deposit averages 

300 ppm and shows little variation in the samples studied by 

Ye and others (2011).

Polymetallic tin vein deposits.—Polymetallic tin deposits 

of Bolivia (including the Bolívar, Hauri Hauri, Porco, Potosí, 

and San Lorenzo deposits, and others) constitute a major 

potential source of indium. Total indium metal resources in 

the Bolivian polymetallic tin deposits are estimated to be 

12,000 metric tons (Ishihara and others, 2011). The polymetallic tin belt that occurs in the Eastern Cordillera of the 

Andes Mountains of Bolivia (Ishihara and others, 2011) is 

underlain by Paleozoic marine sedimentary rocks overlain 

and intruded by Cenozoic felsic volcanic and intrusive rocks 

that are associated with the mineralization (fig. I7). Deposits 

occur as dense arrays of narrow, elongate zinc sulfide and 

cassiterite (SnO2) veins that occur in volcanic rocks, eroded 

volcanic-intrusive complexes, porphyry stocks and associated 

breccia, and Paleozoic sedimentary rocks. Individual Bolivian 

polymetallic tin deposits have indium contents in composite 

ore samples or zinc concentrates of 584 ppm at Bolívar, 

3,080 ppm at Hauri Hauri, 499 ppm at Porco, 292 ppm at 

Potosí, and 1,080 ppm at San Lorenzo (Ishihara and others, 

2011; Murakami and Ishihara, 2013).

Mining and Beneficiation Methods

Germanium.—Germanium is initially recovered 

from the leaching of zinc residues or coal ash followed by 

precipitation of a germanium concentrate. The extraction of 

germanium from its ores includes two stages—the production 

of a germanium concentrate by retorting, roasting, or pyrometallurgy and deposition of germanium sulfide or oxide. 

Concentrates are chlorinated to germanium tetrachloride 

(GeCl4 ) and subsequently purified by hydrolysis to germanium 

dioxide (GeO2), reduced pyrolytically with hydrogen gas (H2 ) 

to germanium metal powder, and melted into bars (Butterman 

and Jorgenson, 2005).

Indium.—Indium is recovered as a byproduct during the 

refining process of other base-metal ores and concentrates, 

most commonly the zinc ore mineral sphalerite. Several 

complex and proprietary methods have been developed to 

extract indium from different source materials; however, a 

generalized recovery process is described, as follows, by 

Alfantazi and Moskalyk (2003). Waste products generated 

during the zinc refining process, such as dusts, fumes, 

residues, and slag, are collected and treated for the recovery of 

indium. These materials are first leached with hydrochloric or 

sulfuric acid to dissolve the indium into an aqueous solution. 

The solution then undergoes a solvent extraction process to 

increase the concentration of indium in the solution. Next, 

indium is removed from the solution by means of cementation, and the resulting indium sponge is cast into anodes for 

electrolytic refining to produce indium metal of standard-grade 

purity (99.97 or 99.99 percent).
I12 Critical Mineral Resources of the United States—Germanium and Indium

5,035

5,950

5,023

5,025 5,545

5,655

5,921

5,833

5,680

6,045

5,118

5,320

5,218

5,995

6,060

6,542

5,213

6,402

3,812

5,833

Elevation, in meters

0 100 200 KILOMETERS

0 50 100 MILES

Sucre

PERU

CHILE

ARGENTINA

BOLIVIA

Tupiza

Uyuni

Potosi

Oruro

Cochabamba

La Paz

Lake Poopo

Lake Titicaca

Salar

de Coipasa

Salar de Uyuni

Pirquitas

Tatasi

San Vicente

Chorolque

Colquechaca

Avicaya

Pulacayo

San Lorenzo

Reserva

Tuntoko

Carguaicollo

Bolivar

San Cristobal

Porco

Potosi

Colquiri

Comanche

Huari Huari

Siete Suyos-Animas

Figure Germanium and Indium I7

Sn-rich, Pb-Zn zone

Cu-rich, Pb-Zn zone

Quaternary volcanics

Salt flat

In-rich vein

Vein

EXPLANATION

Map

area

SOUTH AMERICA

20° S

22° S

18° S

70° W 68° W 66° W

Figure I7. Map showing indium-bearing tin-polymetallic ore deposits in Bolivia. Modified from 

Ishihara and others (2011); copper-rich zone is from Schwarz-Schampera and Herzig (2002).

Cu, copper; In, indium; Pb, lead; Sn, tin; Zn, zinc
Resources and Production I13

Resources and Production

Reserves, Other Identified Resources, and 

Undiscovered Resources

Estimating reserves of either germanium or indium is 

difficult because they are byproduct commodities that come 

from a wide variety of ore deposit types. For example, 

both germanium and indium come largely from the zinc ore 

mineral sphalerite. Global zinc reserves are estimated to be 

250 million metric tons, and domestic U.S. reserves are 

estimated to be 11 million metric tons. Accurately converting 

these values to reserves for germanium or indium would 

require an intensive program of chemical analyses for these 

elements in sphalerite, and there are insufficient data available to calculate average concentrations in all the relevant 

deposits, which is a necessary step in determining reserves. 

Guberman (2013a) estimated U.S. reserves for germanium to 

be 450 metric tons.

Germanium and indium are closely associated with 

currently produced zinc, copper, and tin-polymetallic ores, 

and coal deposits (Weeks, 1973; Schwarz-Schampera and 

Herzig, 2002; Höll and others, 2007; Bleiwas, 2010). Much of 

the ore that might provide byproduct germanium and indium 

currently is processed without separation of these constituents. 

The extent of recovery in the future will depend on (a) a 

continued increase in demand for indium and germanium, 

(b) better metallurgical recovery technology from sulfides 

and other materials, and (c) careful implementation of 

recycling of electronic products and production waste. Ores 

of metals other than zinc (especially some copper-tin ores), 

coal deposits, and fly ash from coal burning, have potential 

significant concentrations of germanium and indium and might 

become important sources. It is unlikely that exploration for 

deposits containing germanium and indium will be conducted 

solely because the deposits contain these byproduct elements. 

Two deposits have been developed primarily for production 

of germanium or indium. The indium-producing Toyoha 

Mine in Hokkaido Prefecture, Japan, closed in 2006 and the 

germanium-producing Apex Mine in Utah closed in 1992.

The Toyoha Mine, which is a lead-zinc-silver-indium 

vein deposit located near Sapporo, Japan, was until recently 

the world’s leading indium producer (30 metric tons per year), 

but the mine was closed when ore reserves were exhausted. 

The deposit consists of a series of steeply dipping zinc-leadsilver-copper-tin-indium veins that occur across an area 

500 meters (m) long and 2,500 m wide and has vein widths of 

up to 30 m. Total indium production from the Toyoha Mine 

was estimated to be 5,000 metric tons, based on an average 

indium grade of 138 ppm (Shimizu and Morishita, 2012).

The Apex Mine near Saint George, Utah, was the first 

mine in the world devoted primarily to production of gallium 

and germanium when it was reopened in 1986 (Bernstein, 

1986). The Apex deposit was mined from 1884 to 1962 for 

copper, lead, and silver, which removed the copper-rich ores. 

Gallium and germanium mining after 1986 focused mainly on 

azurite, goethite, hematite, jarosite, limonite, malachite, and 

minor amounts of other metal arsenates, carbonates, oxides, and 

sulfates. These minerals occur in dolomitized and silicified 

breccia, gouge, and fissures in steeply dipping fault zones 

within the gently dipping Pennsylvanian Callville Limestone. 

Germanium is concentrated (up to 7,000 ppm) mostly in the 

goethite, hematite, and limonite (Bernstein, 1986). Reserves 

of 247,000 metric tons at an average grade of 1,000 ppm 

germanium were estimated in 1986 (Ludington and others, 

2006). Production was reported as 2,555 kilograms of germanium 

in 1986 (Burgin, 1988, p. 485). Cox and Singer (1986) and 

Ludington and others (2006) consider the Apex deposit to be a 

Kipushi-type deposit.

Although deposit and mine development exclusively for 

germanium and indium is unlikely, two promising deposits 

that could be sources of production primarily for these minerals 

have recently been described. Ludington and others (2006) 

report small base-metal deposits in Paleozoic limestone in the 

Gold Butte area and the Tramp Ridge area in Clark County, 

Nevada. These are unlikely ever to be significant sources of 

copper, lead, silver, or zinc; however, they may contain potentially significant concentrations of gallium and germanium, as 

well as cobalt. The characteristics of these deposits indicate 

clearly that they can be considered to be Kipushi-type deposits 

(Ludington and others, 2006).

Another prospect of interest is the Crypto zinc-copperindium project in Juab County, Utah, which contains copper 

and zinc skarn deposits in Paleozoic limestones (Nilsson and 

others, 2010). Indium concentrations average about 31 ppm, 

and indium resources are estimated to be about 475 metric 

tons of contained indium.

Production

Worldwide production of both germanium and indium 

has increased dramatically over the past several decades. 

Figure I8 shows the increases in global production since 

1999 (Guberman, 2013a; Tolcin, 2013a). Production of each 

commodity in individual countries in 2011 and 2012 are 

discussed separately below.

Germanium.—In 2011, the world’s total production of 

germanium was estimated to be 118 metric tons (table I2).

This total includes germanium recovered from zinc concentrates,

fly ash from coal burning, and recycled material. Worldwide, 

primary germanium was recovered in Canada from zinc 

residues (concentrates) shipped from the United States 

(specifically, the Red Dog zinc-lead mine in Alaska and 

a previously idled zinc mine complex in Tennessee); in 

China, from zinc residues and coal from multiple sources; in 

Finland, from zinc residues (concentrates) originating in Congo 

(Kinshasa); and in Russia, from coal from domestic sources 

on Sakhalin Island in Sakhalinskaya Oblast’ (Höll and others, 

2007). The vast majority of germanium production was 

concentrated in Canada and China (Bleiwas, 2010).
I14 Critical Mineral Resources of the United States—Germanium and Indium

Figure Germanium I8. 

1995

900

800

700

600

500

400

300

200

100

0

2000

Year

Indium

Germanium

Production, in metric tons of element

2005 2010 2015

Figure I8. Graph showing worldwide production of 

germanium and indium from 1995 to 2012. Production 

of both minerals increased significantly. Graph was 

compiled using data from Jorgenson and George (2004), 

Guberman (2013a, b) and Tolcin (2013a, b).

Table I2. Average estimated annual refinery production of 

germanium and indium, by area, for 2011 and 2012.

[Sources: Guberman, 2013a, 2014; Tolcin, 2013a, 2014. W, withheld to 

avoid disclosing proprietary data; —, zero production; n.d., no data]

Country

Refinery production

(metric tons)

2011 2012

Germanium

United States 3 W

China 80 105

Russia 5 5

Other countries1 30 40

World total 118 150

Indium

United States — —

Belgium 30 30

Brazil 5 n.d.

Canada 75 62

China 380 405

Japan 70 71

Korea, Republic of 70 165

Peru n.d. 11

Russia 5 13

Other countries 27 25

World total 662 782

1

Germanium production in Canada is substantial, but exact figures are 

proprietary, so Canadian production is included with “Other countries.”

Owing to the value of refined germanium, new scrap 

generated during the manufacture of fiber-optic cables, infrared 

optical fiber, and substrates is typically reclaimed and fed back 

into the production process. Recycling of germanium recovered 

from used materials, such as fiber-optic window blanks in 

decommissioned military vehicles or fiber-optic cables, has 

increased during the past decade. A germanium refinery in Utica, 

New York, produces germanium tetrachloride for optical fiber 

production. Another refinery in Quapaw, Oklahoma, produces 

refined germanium compounds for the production of fiber optics, 

infrared devices, and substrates for electronic devices.

Indium.—Globally, primary production of indium metal 

in 2011 was estimated to be about 662 metric tons, of which 

about one-half was produced in China (table I2). Other leading 

producers were Belgium, Canada, Japan, and the Republic of 

Korea. These five countries accounted for almost 95 percent of 

primary indium production. Major production facilities outside of 

China included Teck Resources Ltd.’s Trail lead-zinc metallurgical 

complex in Canada, Korea Zinc Ltd.’s Onsan lead-zinc smelter in 

the Republic of Korea, Dowa Metals and Mining Co. Ltd.’s Akita 

zinc smelter in Japan, and Umicore NV’s Hoboken preciousmetals refinery in Belgium (Guberman, 2013a; Tolcin, 2013a).

In the United States, indium metal was not produced as a 

byproduct at any smelters or refineries. Domestic production 

of indium consisted of upgrading imported indium metal and 

powder. Indium Corp. of America and Umicore Thin Film 

Products accounted for the majority of U.S. production of 

indium metal and products.

Prices and Pricing

Germanium.—Germanium is traded through longterm supply contracts and individual trades between large 

consumers and suppliers as well as private trading houses. 

The terms of such trades are generally unavailable publicly, 

and a market price, in the conventional sense, does not exist. 

Publicly available price quotes actually represent estimates of 

representative prices in trades being executed on a particular 

day, which are compiled through recurring interviews with 

individual traders. Germanium prices have fluctuated on 

a regular basis owing to supply disruptions and increases 

in consumption of germanium for emerging applications. 

Germanium dioxide prices in 2012, for example, increased 

by 49 percent to $1,375 per kilogram in late September, up 

from $925 per kilogram in mid-March (Guberman, 2013a, b). 

Because germanium commonly is produced as a byproduct 

of zinc, the supply and price of germanium is often affected 

by the zinc market. Germanium metal trades at a significantly 

higher price than germanium dioxide, reflecting the value 

added and cost of this further refining.

Indium.—The average indium price in 2012 was 

$530 per kilogram. In 2011, indium prices fluctuated between 

a high of $785 at the beginning of the year and a low of

 $485 in July, ending the year at approximately $540 per 

kilogram (Tolcin, 2013a, b).
Environmental Considerations I15

Environmental Considerations

Germanium and indium are becoming increasingly more 

important to society, given their increasing use in microelectronic applications. Compared with more-abundant industrial 

metals, such as lead and zinc, information about the behavior 

of germanium and indium in the environment is limited.

Sources and Fate in the Environment

In surface weathering environments, germanium and 

indium may dissolve from host minerals and form complexes 

with chloride, fluoride, hydroxide, organic compounds, phosphate, or sulfate. In dissolved form, germanium and indium 

most often occur in the +4 and +3 oxidation states, respectively 

(Wood and Samson, 2006). The behavior of germanium in 

weathering environments mirrors that of silicon, so measurement of germanium-to-silicon ratios in diatoms (single-celled 

alga with a cell wall of silica) in rivers and oceans has been 

used as a tool to document weathering of silicate rocks through 

geologic time (for example, Froelich and others, 1992). Indium 

has no equivalent relationship with another element in weathering environments, but it does tend to co-precipitate with iron 

and manganese oxyhydroxide minerals in weathering environments (Kabata-Pendias and Mukherjee, 2007). Germanium has 

a particular affinity for organic material, which may be one of 

the mechanisms of germanium enrichment in coal.

The tendency for germanium and indium to be dissolved 

and transported largely depends upon the pH and temperature 

of weathering solutions. Both elements are commonly concentrated in sulfide minerals and are therefore relatively mobile 

under most weathering conditions because of the instability of 

sulfide minerals subjected to near-surface oxidizing conditions. Oxidative dissolution of sulfide minerals releases metals 

and sulfuric acid, resulting in the environmental problem 

known as acid mine drainage (AMD). Metals, including 

germanium and indium, dissolved in acidic drainage can be 

attenuated through precipitation, sorption to (oxyhydr)oxide 

minerals, or dilution by mixing with water at circum-neutral 

pH; otherwise, metals are transported downstream.

Examples of natural concentrations of germanium in 

soil and water are given in table I3. Germanium contents in 

soils generally range from 0.5 to 2.3 ppm (Shacklette and 

Boerngen, 1984; Eriksson, 2001), and concentrations as high 

as 173 ppm have been observed in soil in Sweden (Eriksson, 

2001). Germanium concentrations in seawater range from 

nearly 0.05 to 13 parts per trillion (ppt) (Froelich and others, 

1985; Santosa and others, 1997; Ellwood and Maher, 2003); 

seawater germanium concentrations follow a typical nutrient 

profile, as the surface water is more depleted of germanium 

than the bottom water (Froelich and others, 1985). Seawater 

germanium concentrations are generally lower than the 2 to 

80 ppt range (mean is 7 ppt) observed for uncontaminated 

rivers around the world (Froelich and others, 1985; Mortlock 

and Froelich, 1987; Gaillardet and others, 2003). Geothermal 

waters contain higher germanium contents, as evidenced by 

the range observed in hot springs (1,500 to 6,800 ppt) in the 

Cascade Range, United States (Siebert and others, 2006). 

Compared to dissolved forms, germanium is highly concentrated in small particulates in river water; the global average 

in suspended sediment is 1,230,000 ppt (1.23 ppm) (Viers and 

others, 2009).

Examples of natural concentrations of indium in soil, 

water, and air are given in table I4. The general worldwide 

indium content in soils ranges from 0.01 to 0.5 ppm (KabataPendias and Pendias, 2001). Indium concentrations in soils 

in the United States range from less than 0.02 to 0.12 ppm 

(Smith and others, 2005), and in soils in Sweden, from less 

than 0.04 to 0.07 ppm (Eriksson, 2001). Indium concentrations in seawater generally range from 0.003 to 1.23 ppt 

(Amakawa and others, 1996; Alibo and others, 1999; Nozaki, 

Lerche, Alibo, and Snidvongs, 2000; Obata and others, 2007), 

although dissolved and particulate indium in seawater near 

Japan has been measured to be as high as 4.7 and 35 ppt, 

respectively (Miyazaki and others, 2012). Rivers and lakes in 

Japan contain dissolved indium in concentrations ranging from 

0.16 to 3.2 ppt (Nozaki, Lerche, Alibo, and Tsutsumi, 2000; 

Miyazaki and others, 2012), whereas indium in particulate 

matter in rivers in Japan can be as high as 12.8 ppt (Miyazaki 

and others, 2012). Indium occurs naturally in the atmosphere 

as part of mineral dust particles; indium concentrations of 

0.05 and 20 picograms per cubic meter (pg/m3) of air have 

been measured in remote regions (Kabata-Pendias and 

Pendias, 2001, and references therein).

Above-background concentrations of germanium and 

indium in the environment can result from mining and ore 

processing, particularly of zinc, lead, and sometimes copper 

sulfide ores. Total emissions of indium to the atmosphere 

from copper and zinc production were estimated to be 

between 11,000 and 39,000 kilograms per year during 1983 

(Nriagu and Pacyna, 1988). Air indium concentrations over 

more-industrialized regions may range from 20 to 1,200 pg/m3

(Kabata-Pendias and Pendias, 2001, and references therein). 

Likewise, the concentration of indium in air near a leadsmelting complex was measured as 5,800 pg/m3

 (Ragaini 

and others, 1977). In addition to mining and ore processing 

inputs, anthropogenic release of germanium to the atmosphere 

occurs during coal combustion. For example, dissolved 

median germanium concentrations in world rivers are 7 ppt in 

locations distant from and 24 ppt in locations proximal to coal 

combustion centers, respectively (Froelich and others, 1985).

The recycled content, or proportion of scrap, used in 

germanium and indium production is between 25 and 50 percent, 

and the fraction of germanium and indium in discarded products 

that get recycled is less than 1 percent (Graedel and others, 

2011). One of the main challenges of germanium and indium 

recycling is that these metals are often comingled with 

other “specialty metals” in high performance alloys, making 

recovery technologically and economically unfeasible (Reck 

and Graedel, 2012). Given the importance of germanium and 

indium in emerging technologies, however, germanium and 

indium recycling is likely to increase in the future.
I16 Critical Mineral Resources of the United States—Germanium and Indium

Table I3. Germanium concentrations in rocks, soils, and waters.

[cm, centimeter; ppm, part per million; ppt, part per trillion; μm, micrometer]

Environment

and (or) location

Germanium 

concentration Unit Notes Reference(s)

Rocks

Upper continental crust 1.6 ppm Average Taylor and McLennan (1995)

Bulk continental crust 1.6 ppm Average Taylor and McLennan (1995)

Lower continental crust 1.6 ppm Average Taylor and McLennan (1995)

Soils

Western United States 1.2 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Eastern United States 1.1 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Topsoil, Sweden < 1.0 to 95 ppm Range; median is 7 ppm Eriksson (2001)

Subsoil, Sweden < 1.0 to 173 ppm Range; median is 21 ppm Eriksson (2001)

Waters

Antarctic Ocean 0.4 to 7.3 ppt Unfiltered; profile with enriched 

bottom water

Froelich and others (1985)

North Atlantic Ocean < 0.4 to 2.9 ppt Unfiltered; profile with enriched 

bottom water

Froelich and others (1985)

Pacific Ocean < 0.07 to 2.5 ppt Range of averages for surface water; 

dissolved and colloidal (< 0.45 μm)

Santosa and others (1997)

Pacific Ocean

(northwestern)

0.4 to 8.7 ppt Unfiltered; profile with enriched 

bottom water

Froelich and others (1985)

South Pacific Ocean

(New Zealand)

0.05 to 0.28 ppt Dissolved (< 0.2 μm); surface water Ellwood and Maher (2003)

South Pacific Ocean

 (New Zealand)

0.09 to 1.9 ppt Dissolved (< 0.2 μm); profile with 

enriched bottom water

Ellwood and Maher (2003)

Bering Sea 1.8 to 13 ppt Unfiltered; profile with enriched 

bottom water

Froelich and others (1985)

World rivers 2 to 18 ppt Dissolved and colloidal (< 0.4 μm); 

median is 7 ppt distant from coal 

combustion areas

Froelich and others (1985)

World rivers—Contaminated 

(near coal combustion or

Ge refinery areas)

0.7 to 3,050 ppt Dissolved and colloidal (< 0.4 μm); 

median is 24 ppt near heavy coal 

combustion areas

Froelich and others (1985)

Small rivers worldwide 2 to 20 ppt Dissolved and colloidal (< 0.4 μm) Mortlock and Froelich (1987)

African rivers 6.5 to 80 ppt Dissolved load (< 0.2 μm) Gaillardet and others (2003)

North American rivers 1.4 to 22 ppt Dissolved load (< 0.2 μm) Gaillardet and others (2003)

South American rivers 4 to 8 ppt Dissolved load (< 0.2 μm) References within Gaillardet 

and others (2003)

Amazon River, 

South America

5 to 11 ppt Dissolved and colloidal (< 0.4 μm) Mortlock and Froelich (1987)

Changjiang River, China 12.2 ppt Dissolved load (< 0.2 μm) Gaillardet and others (2003)

Columbia River, 

United States

12 to 20 ppt Dissolved and colloidal (< 0.4 μm) Mortlock and Froelich (1987)

Congo River, Africa 6 to 9 ppt Dissolved and colloidal (< 0.4 μm) Mortlock and Froelich (1987)

Idle River, England 8.2 ppt Dissolved load (< 0.2 μm) References within Gaillardet 

and others (2003)

Yukon River, North 

America

4 to 5 ppt Dissolved and colloidal (< 0.4 μm) Mortlock and Froelich (1987)

World rivers—Sediment 1,230,000 ppt 1.23 ppm; average suspended

sediment

Viers and others (2009)

Hotsprings, Cascades, 

United States

1,500 to 6,800 ppt Samples described in references Siebert and others (2006) and 

references therein

Geothermal waters <24 to 293 ppb None Höll and others (2007)
Environmental Considerations I17

Table I4. Indium concentrations in rocks, soils, waters, and air.

[b.d.l., below detection limit; cm, centimeter; μm, micrometer: ppm, part per million; ppt, part per trillion; pg/m3, picogram per cubic meter]

Environment

and (or) location

Indium 

concentration Unit Notes Reference(s)

Rocks

Upper continental crust 0.05 ppm Average Taylor and McLennan (1995)

Bulk continental crust 0.05 ppm Average Taylor and McLennan (1995)

Lower continental crust 0.05 ppm Average Taylor and McLennan (1995)

Soils

Soils, worldwide 0.01 to 0.5 ppm None Kabata-Pendias and Pendias (2001)

Conterminous

United States

< 0.02 to 0.12 ppm 0 to 5 cm depth; median is 0.04 ppm Smith and others (2005)

Conterminous

United States

< 0.02 to 0.04 ppm O horizon, if present; median

is 0.03 ppm

Smith and others (2005)

Conterminous

United States

< 0.02 to 0.12 ppm A horizon; median is 0.04 ppm Smith and others (2005)

Conterminous

United States

< 0.02 to 0.10 ppm C horizon; median is 0.04 ppm Smith and others (2005)

Topsoil, Sweden < 0.04 to 0.064 ppm Median is < 0.04 ppm Eriksson (2001)

Subsoil, Sweden < 0.04 to 0.073 ppm Median is 0.02 ppm Eriksson (2001)

Waters

Seawater, North 

Atlantic Ocean

0.07 to 0.19 ppt Dissolved and colloidal (< 0.45 μm) Alibo and others (1999)

Seawater, northwest 

Pacific Ocean

0.005 to 0.018 ppt Unfiltered Amakawa and others (1996)

Seawater, Mediterranean 0.72 to 1.23 ppt Dissolved and colloidal (< 0.6 μm) Alibo and others (1999)

Seawater, Japan b.d.l. to 4.7 ppt Particulate; median is 3.1 ppt Miyazaki and others (2012)

Seawater, Japan 0.2 to 34.8 ppt Dissolved and colloidal (< 0.45 μm); 

median is 2.6 ppt

Miyazaki and others (2012)

Sea of Japan 0.005 to 0.067 ppt Dissolved (< 0.04 μm) Obata and others (2007)

River water, Japan 0.16 to 1.7 ppt Dissolved (< 0.04 μm) Nozaki, Lerche, Alibo, and 

Tsutsumi (2000)

River water, Japan b.d.l. to 3.2 ppt Dissolved and colloidal (< 0.45 μm) Miyazaki and others (2012)

River water, Japan b.d.l. to 12.8 ppt Particulate Miyazaki and others (2012)

Lake water, Japan 1.1 to 3.2 ppt Dissolved and colloidal (< 0.45 μm) Miyazaki and others (2012)

Lake water, Japan 1.5 to 3.2 ppt Particulate Miyazaki and others (2012)

Estuary, Thailand 0.003 to 0.049 ppt Dissolved (< 0.04 μm) Nozaki, Lerche, Alibo, and 

Snidvongs (2000)

Air

North America 20 to 140 pg/m3 None Kabata-Pendias and Pendias 

(2001) and references therein

South Pole 0.05 pg/m3 None Kabata-Pendias and Pendias 

(2001) and references therein

Idaho, United States 5,800 pg/m3 Near a lead smelter complex Ragaini and others (1977)

Japan 1,200 pg/m3 None Kabata-Pendias and Pendias 

(2001) and references therein

Shetland Islands 20 pg/m3 None Kabata-Pendias and Pendias 

(2001) and references therein

West Germany 30 to 360 pg/m3 None Kabata-Pendias and Pendias 

(2001) and references therein
I18 Critical Mineral Resources of the United States—Germanium and Indium

Mine Waste Characteristics

Mine waste is generally considered to be the material that 

originates and accumulates at a mine site that has no current 

economic value (Lottermoser, 2010); it includes both solid 

and liquid waste. Mine waste that results from extraction of 

germanium- and indium-bearing copper-lead-zinc ore usually 

consists of waste rock and tailings piles, and possibly pit 

lakes. Tailings are the residual silt- to fine-sand-sized grains 

that result from ore grinding and processing, and they are 

usually disposed of in waste dumps. Germanium is often 

recovered from ores in SEDEX zinc-lead-silver deposits, such 

as at the Red Dog Mine in Alaska, and from MVT lead-zinc 

deposits, such as those in the Tri-State mining district, which 

covers approximately 1,800 square kilometers of Kansas, 

Oklahoma, and Missouri (Butterman and Jorgenson, 2005). 

Indium can be concentrated in ores in vein stockwork tin 

and tungsten deposits, such as the Mount Pleasant Mine in 

New Brunswick, Canada, and in VMS deposits, such as the 

Kidd Creek Mine in Ontario, Canada (Jorgenson and George, 

2004). The tonnage of tailings generated at the Red Dog 

Mine totaled 27.4 million metric tons in 2006 and is projected 

to total 88 million metric tons by 2031 (SRK Consulting 

(Canada) Inc., 2007). The Tri-State mining district is the 

largest U.S. Environmental Protection Agency Superfund site 

in the United States (Leach and others, 2010); the estimated 

total tonnage of waste rock at this site is 91 million metric 

tons (U.S. Environmental Protection Agency, 2006). In 1999, 

the Kidd Creek mining and metallurgical operation stored 

more than 100 million metric tons of tailings over an area 

of approximately 1,200 hectares (Al and Blowes, 1999).

The mineralogy of the mine waste in germanium- and 

indium-enriched deposits tends to be similar to the mineralogy 

of the deposit, except that the proportion of sulfides, such 

as sphalerite, galena (PbS), and chalcopyrite (CuFeS2 ), is 

reduced relative to typical gangue minerals, such as barite 

(BaSO4 ), calcite (CaCO3 ), dolomite ((Ca,Mg)(CO3)2), pyrite 

(FeS2 ), pyrrhotite (Fe1–x S), and quartz (SiO2) (Kelley and 

others, 1995; Leach and others, 1995; Taylor and others, 

1995). For example, tailings from the Kidd Creek deposit 

contain (in weight percent) quartz and various silicates 

(75 to 85), pyrite (10 to 25), carbonate minerals (7 to 8) 

pyrrhotite (1 to 2), and sphalerite and chalcopyrite (1 to 2) 

(Al and others, 1994). Indium-enriched VMS deposits may 

also contain anhydrite (CaSO4) and gypsum (CaSO4 • 2H2O), 

chlorite (Mg,Fe)3(Si,Al)4O10(OH)2(Mg,Fe)3(OH)6, and 

magnetite (Fe3O4 ) as gangue minerals (Taylor and others, 

1995). Some deposits undergo appreciable weathering before 

discovery, which can result in oxidized zones on the surface 

of orebodies known as gossan. Minerals that precipitate 

from the weathering of ore minerals in gossan commonly 

include amorphous silica (SiO2 ), goethite (FeO(OH)), jarosite 

(KFeIII

3

 (OH)6(SO4)2) and other hydroxysulfate minerals, 

scorodite (FeAsO4 • 2H2O), and secondary sulfide minerals, 

such as chalcocite (Cu2S) and covellite (CuS) (Taylor and 

others, 1995).

Because of their association with sulfide and secondary 

minerals, elements such as Ag, Al, As, Au, B, Ba, Bi, Cd, 

Co, Cu, F, Fe, Ga, Hg, Mn, Mo, Ni, Pb, Sb, Sn, and Zn 

can be found in germanium- and indium-enriched mineral 

deposits (Briskey, 1986; Kelley and others, 1995; Leach and 

others, 1995; Taylor and others, 1995). During oxidizing 

and acidic (pH<3) weathering conditions, which may be 

expected in mine waste with little to no acid-neutralizing 

capacity, many of these accessory elements are expected to 

be mobile (Smith and Huyck, 1999). The acid-neutralizing 

capacity of carbonate-hosted MVT mine waste is expected 

to be greater, however, resulting in the accessory elements 

listed above being less mobile or immobile under oxidizing 

and circum-neutral (5<pH<8) weathering conditions (Smith 

and Huyck, 1999). The exception is zinc, which can make 

up a large percentage of total dissolved metals draining from 

MVT deposits (Plumlee and others, 1999).

Human Health Concerns

Both germanium and indium have no known physiological role in human biochemical functions, and thus are 

considered to be nonessential. Germanium and indium can 

be ingested through food and inhaled with dust in industrial 

environments. For example, workers in an indium ingot 

manufacturing plant who were exposed to insoluble indium 

compounds through inhalation showed higher concentrations 

of indium in plasma and urine relative to a control group 

(Hoet and others, 2012). Germanium toxicity is generally low, 

and germanium does not appear to be carcinogenic; to the 

contrary, the organic germanium compound spirogermanium 

has been shown to destroy cancer cells (Gerber and Léonard, 

1997, and references therein). Excessive ingestion of germanium has taken place owing to the belief that germanium in 

supplements has antioxidant properties. This has led to renal 

failure and renal and multiorgan dysfunction (Glei, 2004, 

and references therein). Unlike solid germanium compounds, 

the gas germane (GeH4) is highly toxic and can be lethal 

at concentrations near 150 ppm or higher (Glei, 2004, and 

references therein). As a result, the U.S. Occupational Safety 

and Health Administration (OSHA) has set a permissible 

exposure limit of 0.2 ppm GeH4

 over an 8-hour workday 

(Occupational Safety and Health Administration, 2013a).

The human toxicity of indium is relatively well 

documented, partly because the short half-life of radioactive 

111In lends itself to use in nuclear medicine to study inflammatory processes and tumors (Madden and others, 2004, 

and references therein). Industrial workers may be exposed 

to insoluble indium compounds, such as indium arsenide, 

indium phosphide, and ITO, all of which have proven to be 

toxic to animals. Chronic (low doses over extended time 

periods) exposure to inhaled indium phosphide, in particular, 

has shown carcinogenic effects in animals (Tanaka and others, 

2010). A few case studies of workers exposed to insoluble 

indium compounds have shown that inhaled indium can cause 

interstitial lung damage (Hamaguchi and others, 2008, and 
Environmental Considerations I19

references therein). Furthermore, among the metal arsenic 

compounds used in the semiconductor industry, indium 

arsenide has proven to cause the greatest impairment to 

lung function (Tanaka, 2004, and references therein). To 

prevent negative effects from occupational indium exposure, 

OSHA has set a limit of 0.1 milligram per cubic meter for 

indium-bearing dust in workplace air over an 8-hour workday 

(Occupational Safety and Health Administration, 2013b).

Mining of zinc-lead-copper deposits, where germanium 

and indium may be byproducts, can potentially mobilize 

elements that are known human toxins. Lead and arsenic 

are perhaps the best-known examples of chemicals that are 

released into the environment through mining and that end up 

in drinking water. Lead is known to affect the neurological 

development of children (Holecy and Mousavi, 2012), 

and arsenic is known to have carcinogenic effects (Gupta 

and others, 2012). Other associated elements also have 

the potential to affect human health when present above 

threshold concentrations in air, drinking water, and soil. The 

current U.S. National Ambient Air Quality Standard for lead 

is 0.15 micrograms per cubic meter (U.S. Environmental 

Protection Agency, 2013b), and the current U.S. primary and 

secondary drinking water standards for lead, copper, and 

zinc are 0, 1, and 5 milligrams per liter (mg/L), respectively 

(U.S. Environmental Protection Agency, 2013a). Canadian 

agricultural soil quality guidelines for copper, lead, and zinc 

are 63, 70, and 200 ppm, respectively (Canadian Council of 

Ministers of the Environment, 1999). These concentrations are 

sometimes exceeded near mine sites and ore processing plants.

Ecological Health Concerns

Relatively few studies have focused on the ecological 

impacts of germanium and indium mobility in the environment. One of several useful endpoints used in toxicity tests 

is the lethal concentration that leads to 50 percent mortality 

(LC50) after exposure to a substance for a certain amount 

of time. Fish tend to be sensitive to low concentrations of 

dissolved metals, and they are therefore useful indicators 

of contamination in aquatic systems. Chronic toxicity 

tests with dissolved germanium (in the form of GeO2

) and 

developing rainbow trout (Oncorhynchus mykiss) revealed 

a mean LC50 value of 0.05 mg/L after 28 days of exposure 

(Birge and others, 1980). Another aquatic organism that is 

commonly used in toxicity studies is the freshwater amphipod 

Hyalella azteca. The 1-week LC50 for germanium incubated 

with H. azteca was 0.21 ppm in soft water and more than 

3.15 ppm in hard water (Borgmann and others, 2005). An 

in vitro study of the effects of indium nitrate (In(NO3)3) on fish 

cells revealed that cell proliferation was significantly reduced 

at indium concentrations of 356 ppm (3-day exposure) and 

677 ppm (1-day exposure) (Zurita and others, 2007). Exposure 

of indium nitrate to the bacterium Aliivibrio fischeri, the 

alga Chlorella vulgaris, and the cladoceran Daphnia magna

resulted in 50 percent of populations exhibiting decreased 

functionality over a range of effective concentrations (EC50 ), 

with the most sensitive EC50 being inhibited bioluminescence 

of A. fischeri at 6 ppm indium (Zurita and others, 2007). 

These indium concentrations are orders of magnitude 

greater than those in reported natural systems (table I4). 

Germanium and indium are usually not included in the suite 

of elements in studies of the chemistry of mine waste water, 

however, so the toxicity endpoint concentrations cannot be 

compared to likely germanium and indium concentrations in 

mining environments.

Plant tissues contain from 50 to 754 ppm germanium and 

1 to 2 ppm indium (Kabata-Pendias and Mukherjee, 2007), 

although neither element is known to have any physiological 

function in plants. In a study of barley seedlings exposed to 

nutrient solutions with varying germanium concentrations, 

roots and shoots accumulated greater amounts of germanium 

with increasing germanium concentrations, and the only toxic 

effects observed were necrosis of primary leaves at concentrations greater than 1.45 ppm germanium (Halperin and 

others, 1995). The effects of indium on plants remain unknown.

The ecological impacts of major metals with which 

germanium and indium are associated are relatively well 

documented. For example, in toxicity tests with aquatic 

invertebrates incubated with lead-zinc mine tailings from the 

Tri-State mining district, Besser and Rabeni (1987) showed 

that increased concentrations of dissolved cadmium, lead, 

and zinc correlated with decreased survival and growth 

(Besser and Rabeni, 1987). Likewise, various fish species 

experienced rapid mortality when exposed to increasing 

dissolved zinc concentrations, but this effect decreased with 

increasing calcium, magnesium, and sodium concentrations 

(De Schamphelaere and Janssen, 2004). High metal concentrations in soils can also cause phytotoxicity. Zinc is a micronutrient for many plant species, but, when highly concentrated 

in soils, it can inhibit a plant’s metabolic functions and cause 

deficiencies in other essential nutrients, such as copper, iron, 

manganese, and phosphorus (Nagajyoti and others, 2010, and 

references therein).

Carbon Footprint

Indium is used in some of the most efficient thin-film 

photovoltaic cells (solar panels) (Miles and others, 2007) 

and has potential to help reduce the carbon dioxide (CO2 ) 

emissions from energy production. The copper-indium–

gallium-diselenide (CIGS) alloy, among others, is becoming 

an increasingly preferred thin-film product because of its 

relatively low material and manufacturing costs (Kaneshiro 

and others, 2010). The CIGS photovoltaic cells are in the 

initial stages of commercialization, but they are expected 

eventually to compete with other forms of energy production 

as economies of scale permit future cost reductions (Miles and 

others, 2007; Rockett, 2010). The recyclability of generated 

products and the potential for leaching of toxic elements from 

improperly disposed of industrial products are subjects that 

no doubt will need to be addressed as the use of thin-film 

photovoltaic cells increases.
I20 Critical Mineral Resources of the United States—Germanium and Indium

Mine Closure

Most recent and new mining operations include closure 

plans that address issues related to the mine footprint. A 

mine’s footprint includes the waste left on site and locally 

impacted soil and water, as well as ecological impacts, such 

as habitat destruction and loss of biodiversity. Following 

mining in copper-lead-zinc deposits, where germanium and 

indium may be byproducts, a common mine closure issue is 

the potential for AMD from the site. Acidic drainage may seep 

from waste piles or tailings ponds, and common remediation 

methods include active water treatment facilities, passive 

limestone-lined channels, or constructed wetlands (Plumlee 

and Logsdon, 1999). The end result of both active and passive 

approaches is eventual precipitation of dissolved metals. 

Precipitated metals tend to be more stable under the prevailing 

anoxic conditions in passive wetland systems, whereas the 

metal-rich precipitates that result from active treatment 

facilities form a sludge that can cause environmental problems 

if not responsibly treated.

At large mines, mine waste is often consolidated into 

pits and submerged under water, forming a tailings pond or 

impoundment. Acid-generating minerals typically are less 

reactive under water because of limited oxygen, but any 

seepage usually needs to be treated. At sites like the Red Dog 

Mine, seepage from the tailings pond will need to be treated 

in perpetuity (U.S. Environmental Protection Agency, 2010). 

The massive tailings impoundment near the Kidd Creek Mine 

is designed to have a conical shape. This limits the amount of 

tailings that may oxidize in near-surface unsaturated zones and 

contribute soluble metals, which allows more control over the 

AMD effluent discharged at the periphery of the impoundment 

(Al and Blowes, 1999). When the potential for AMD exists at 

a mine site, a common mine closure plan includes conducting 

water-quality surveys before, during, and after mining.

Another common mine closure issue related to mining in 

germanium- and indium-enriched deposits is the creation of 

large, dry mounds of mine-waste. These waste piles have the 

potential to become unstable and can be a source of metal-rich 

dust. If using mine waste as backfill into mine workings is 

not an option, mine waste pile stability and dust-generating 

issues can often be addressed through grading and covering 

piles with vegetation. Securing waste piles and prevention 

and treatment of AMD are typically figured into the long-term 

costs of active and proposed metal mining projects.

Problems and Future Research

Weeks (1973) and Brobst and Pratt (1973) pointed out 

that resource studies of minor byproduct metals are limited by 

inadequate data, and that is still true today. The sources of the 

materials that are processed at smelters for byproduct germanium and indium are often either not carefully tracked or the 

information is considered proprietary. Data on the distribution 

of germanium and indium are improving but currently are 

still not sufficient to understand content variations in different 

deposits and especially in different zones in given deposits. 

Additional systematic analytical data are needed to establish 

the overall grades of germanium and indium in different 

deposits, and the potentially economic resources that might 

be available in such deposits. The most promising areas of 

research involve studies of the distribution of germanium 

and indium in copper and zinc ore deposits and coal deposits, 

and the making of a systematic compilation of local and 

regional variations.

Significant concentrations of these elements are not 

currently recovered in the processing of other ores, particularly zinc and (or) copper ores; the addition of more recovery 

facilities and improvement in metallurgical techniques 

could significantly increase supplies. Finally, recycling of 

germanium and indium from electronic products and from 

manufacturing waste is presently a significant contributor to 

the production of these mineral commodities, and improving 

the recycling processes could help increase supplies significantly (Bleiwas, 2010).

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Graphite

By Gilpin R. Robinson, Jr., Jane M. Hammarstrom, and Donald W. Olson

Chapter J of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–J

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter J

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................ J1

Introduction.................................................................................................................................................. J1

Commercial Classifications of Graphite ......................................................................................... J1

Natural Graphite ....................................................................................................................... J3

Synthetic Graphite .................................................................................................................... J3

Uses and Applications ...................................................................................................................... J3

Demand and Availability of Supply ................................................................................................. J4

Graphite Prices and Pricing .................................................................................................... J4

Substitutes for Natural Graphite ............................................................................................ J4

Strategic and Critical Resource Issues ................................................................................ J5

Geology......................................................................................................................................................... J5

Geochemistry...................................................................................................................................... J5

Mineralogy.......................................................................................................................................... J5

Deposit Types ..................................................................................................................................... J6

Amorphous Graphite Deposits ............................................................................................... J6

Deposits of Flake Graphite Disseminated in Metasedimentary Rocks ............................ J7

Vein Deposits Containing Lump or Chip Graphite .............................................................. J15

Mining and Beneficiation Methods ....................................................................................................... J16

Resources and Production ...................................................................................................................... J16

Distribution of Graphite Deposits and Districts .......................................................................... J16

World Graphite Production and Resources ................................................................................ J17

Exploration for New Deposits ................................................................................................................. J19

Environmental Considerations ................................................................................................................ J19

Mine Waste Characteristics .......................................................................................................... J19

Human Health Concerns ................................................................................................................. J20

Ecological Health Concerns ........................................................................................................... J20

Carbon Footprint............................................................................................................................... J21

Regulatory and Environmental Considerations .......................................................................... J21

Problems and Future Research .............................................................................................................. J21

References Cited....................................................................................................................................... J22

Suggested citation:

Robinson, G.R., Jr., Hammarstrom, J.M., and Olson, D.W., 2017, Graphite, chap. J of Schulz, K.J., DeYoung, J.H., Jr., 

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. J1– J24, http://dx.doi.org/

10.3133/pp1802J.
Graphite

By Gilpin R. Robinson, Jr., Jane M. Hammarstrom, and Donald W. Olson

Abstract

Graphite is a form of pure carbon that normally occurs 

as black crystal flakes and masses. It has important properties, 

such as chemical inertness, thermal stability, high electrical 

conductivity, and lubricity (slipperiness) that make it suitable for many industrial applications, including electronics, 

lubricants, metallurgy, and steelmaking. For some of these 

uses, no suitable substitutes are available. Steelmaking and 

refractory applications in metallurgy use the largest amount 

of produced graphite; however, emerging technology uses in 

large-scale fuel cell, battery, and lightweight high-strength 

composite applications could substantially increase world 

demand for graphite.

Graphite ores are classified as “amorphous” (microcrystalline), and “crystalline” (“flake” or “lump or chip”) 

based on the ore’s crystallinity, grain-size, and morphology. 

All graphite deposits mined today formed from metamorphism 

of carbonaceous sedimentary rocks, and the ore type is 

determined by the geologic setting. Thermally metamorphosed 

coal is the usual source of amorphous graphite. Disseminated 

crystalline flake graphite is mined from carbonaceous 

metamorphic rocks, and lump or chip graphite is mined from 

veins in high-grade metamorphic regions. Because graphite 

is chemically inert and nontoxic, the main environmental 

concerns associated with graphite mining are inhalation of 

fine-grained dusts, including silicate and sulfide mineral 

particles, and hydrocarbon vapors produced during the mining 

and processing of ore. Synthetic graphite is manufactured from 

hydrocarbon sources using high-temperature heat treatment, 

and it is more expensive to produce than natural graphite.

Production of natural graphite is dominated by China, 

India, and Brazil, which export graphite worldwide. China 

provides approximately 67 percent of worldwide output of 

natural graphite, and, as the dominant exporter, has the ability 

to set world prices. China has significant graphite reserves, 

and China’s graphite production is expected to increase, 

although rising labor costs and some mine production 

problems are developing. China is expected to continue to be 

the dominant exporter for the near future. Mexico and Canada 

export graphite mainly to the United States, which has not had 

domestic production of natural graphite since the 1950s. Most 

graphite deposits in the United States are too small, low-grade, 

or remote to be of commercial value in the near future, and 

the likelihood of discovering larger, higher-grade, or favorably 

located domestic deposits is unlikely. The United States is a 

major producer of synthetic graphite.

Introduction

Graphite, which is a soft form of elemental carbon, is an 

industrial mineral commodity that is produced only in small 

amounts globally—worldwide production of approximately 

1 million metric tons of graphite concentrate was reported as 

yearly production from 2010 to 2012 (Olson, 2012, 2013). 

The word graphite is derived from the Greek word for writing, 

graphein, which reflects the long use of graphite (mixed 

with clay) for the “lead” in pencils. The unique physical and 

chemical properties of graphite, particularly coarse crystalline 

graphite, make it useful for many industrial applications, and 

for some of those uses, no suitable substitutes are available. 

Although graphite is widely disseminated in many types of 

metamorphic and some igneous rocks, most occurrences have 

no economic importance. The significant deposits of graphite 

are found in carbonaceous sedimentary rocks that have been 

subjected to regional or contact metamorphism and in veins 

precipitated from fluids.

Commercial Classifications of Graphite

Sources of commercial graphite include both natural 

graphite mined from rock and synthetic graphite manufactured from other carbonaceous materials. For commercial 

purposes, natural graphite is classified into the following 

three categories, according to its crystallinity, grain size, and 

morphology: amorphous, crystalline (flake), and crystalline 

(lump or chip). These commodity classes differ in the level 

of purity of the graphite, the proposed industrial use, the 

price, and the geologic setting in which the graphite occurs 

(table J1). Descriptions of (a) the graphite commodity classes 

and their industrial uses, and (b) their associated deposit types, 

geologic settings, and producing deposit locations are given 
J2 Critical Mineral Resources of the United States—Graphite

Table J1. Characteristics of graphite commodities, deposits, and uses, by commodity type.

[Descriptions of graphite commodity classes and their industrial uses, commodity prices, and classes of associated deposit types and geologic settings are 

summarized from information in Klar (1958), Weis (1973), Krauss and others (1988), Sutphin (1991a–c), Simandl and Kenan (1997a–c), Taylor (2006), 

and Olson (2011, 2012, 2013). NA, not applicable; μm, micrometer; cm, centimeter; °C, degree Celsius; m, meter; $, U.S. dollar]

Characteristic Commodity type

Amorphous Flake Lump or chip Synthetic

Deposit type Amorphous Disseminated flake Vein NA

Crystallinity Microcrystalline Crystalline Crystalline Microcrystalline to 

crystalline

Properties 

(crystallinity,

form)

Earthy to compact microcrystalline aggregates; 

grain size is <4 μm

Well-developed crystal 

platelets, with grain size 

between 40 μm and 4 cm 

(but generally ≤ 1 cm), 

and 1 to 150 μm thick

Interlocking aggregates of 

coarse crystals. Available 

as powders to 10-cm pieces

Available in particle sizes 

from 2-μm powders to 

2-cm pieces 

Origin Contact metamorphism, 

often by diabasic or 

granitic intrusions, and 

(or) regional metamorphism of carbonaceous 

sediments, often coal

Regional metamorphism 

of carbonaceous sediments at or exceeding 

amphibolite facies 

conditions

Epigenetic veins and lodes 

formed from metamorphic 

fluids in high-grade

metamorphic rocks,

usually granulites

Produced by heat treatment (graphitization) 

of, or chemical deposition from, hydrocarbon 

materials above

2,100 °C

Orebody Layers, seams, and lenses 

in carbonaceous rock, 

each a few meters thick 

and hundreds of meters 

to several kilometers in 

length; may be folded 

and faulted 

Strata-bound; tabular or 

lense form, as much as 

33 m thick and thousands of meters long. 

Irregular in hinge areas 

of folds. The lenses 

have variable graphite 

content internally and 

between lenses

Vein and fracture-filling 

within or crosscutting 

metamorphic structures 

and rock contacts. Individual veins range from 

0.05 to 3 m thick, although 

usually less than 0.3 m, 

and extend up to hundreds 

of meters, although rarely 

more than tens of meters

NA

Ore grade 

(percent carbon)

50 to 90 (aggregates may 

contain nongraphitic 

carbonaceous material)

Generally 5 to 30,

locally higher

40 to 90 (may require

hand sorting)

NA

Deposit tonnage 

(million metric tons)

0.1 to 500 0.1 to 100 Small; no reliable data

for individual veins

NA

Mine operations Surface or underground 

mines using mechanized and sometimes 

hand methods

Generally open pit surface 

mines. Graphite grades 

and ease of mining 

enhanced by weathering destruction of 

gangue minerals

Mines are typically small, 

labor-intensive, and underground. Mines in Sri Lanka 

are from 30 to 400 m deep. 

Ore is hand sorted, washed, 

and screened

NA

Product grade 

(percent graphite)

60 to 90 75 to 97 90 to 99.9 99.95

Main uses Refractories, steel

industry, paint, coatings, and batteries

Refractories, brake

linings, lubricants,

batteries, and

expandable graphite 

applications

Carbon brushes, brake

linings, and lubricants

Batteries, carbon brushes, 

graphite electrodes, 

nuclear moderator rods 

(porosity unsuitable for 

refractory applications)

Prices 

(per metric ton)

December 2011 prices 

(80 – 85 percent

graphite) ranged 

from $600 to $800 

December 2011 prices 

(90% graphite, 

medium-large flake) 

ranged from 

$1,150 to $2,000

Prices in 2011 for Sri Lankan 

lump and chip graphite 

(99% graphite) ranged 

from $1,700 to $2,070. 

Prices vary by purity, 

grade, and size

December 2011 prices 

(99.9 percent graphite) 

ranged from $7,000 to 

$20,000, which was 

about 4 to 7 times that 

of flake graphite

Major producers

and resources

(in order of production level)

China, Republic of Korea, 

Mexico, Austria

China, Brazil, India, 

Madagascar, Germany,

Austria, Norway, 

Canada, Zimbabwe

Sri Lanka China, Japan, United 

States, Germany 

(Roskill Information 

Services, Ltd., 2012)
Introduction J3

by Klar (1958), Weis (1973), Krauss and others (1988), and 

Taylor (2006). Descriptive models of the geology and characteristics of the deposit types that produce graphite are given by 

Sutphin and Bliss (1990), Sutphin (1991a–c), and Simandl and 

Kenan (1997a–c). A summary of information from these and 

other sources follows.

Natural Graphite

Natural graphite is mined from deposits in metamorphic 

rocks, such as marble, schist, and gneiss, and from accumulations in vein deposits. Natural graphite typically forms as a 

result of metamorphism (regional or contact) of accumulations 

of organic matter in sedimentary rocks. Commercial deposits 

of graphite occur in three types of geologic settings that 

generally correspond to the following commodity classes 

(table J1):

a. “Amorphous graphite” is the commercial designation 

for earthy to compact fine-grained graphite that 

generally results from thermal metamorphism of

coal. Commercial deposits typically contain more 

than 1 million metric tons of ore that is more than 

75 percent carbon. Both the raw ore and the 

commodity may contain nongraphitic carbonaceous 

material in addition to graphite.

b. “Flake graphite” is the commercial designation for 

well-developed crystal platelets of graphite that are 

between 40 micrometers (μm) and 4 centimeters (cm— 

but generally less than or equal to 1 cm) in size and 

that are disseminated in beds of carbonaceous sediments that have been subjected to amphibolite-facies 

or higher grade regional metamorphism. Commercial 

deposits generally contain more than 200,000 metric 

tons of ore that grade greater than 8 percent graphite. 

Disseminated flake graphite deposits are located in 

belts of crystalline metamorphic rock that are predominately Archean to late Proterozoic in age.

c. “Lump or chip” is the commercial designation for 

interlocking aggregates of coarse graphite crystals 

that occur as veins or fracture-fillings in igneous and 

crystalline metamorphic rocks that commonly are 

of Precambrian age. The only commercial deposits 

occur in Sri Lanka where families of veins that 

are up to 3 meters (m) thick and which consist of 

60 to 95 percent graphite are mined to depths of 

30 to 650 m (Touzain and others, 2010). The ore 

may be hand sorted to provide a product grade that 

exceeds 90 percent. No reliable data on the tonnages 

of individual vein deposits are available, but most 

of the deposits are small and likely do not exceed 

100,000 metric tons.

Synthetic Graphite

Most of the graphite used by industries in the 

United States is synthetic. Synthetic graphite of high purity 

is produced by heat treatment (graphitization) of, or chemical 

deposition from, hydrocarbon materials above 2,100 degrees 

Celsius (°C). High-temperature processing is required to 

transform the precursor carbon forms to a graphite structure 

and to vaporize impurities, which include hydrogen, metals, 

nitrogen, organic compounds, and sulfur in the source materials. 

As a result of this treatment, synthetic graphite is more than 

99.9 percent graphite, but it has slightly higher porosity, lower 

density, lower electrical conductivity, and a much higher price 

than natural flake graphite. Synthetic graphite is available in 

particle sizes ranging from 2-μm powders to 2-cm pieces; the 

morphology varies from flakey in fine powders to irregular 

grains and needles in coarser products.

Uses and Applications

Graphite has physical and chemical properties of both 

metals and nonmetals, which make it ideally suited for many 

industrial and technology applications. The metallic properties 

include high thermal and electrical conductivity (table J2). 

The nonmetallic properties include inertness to most chemical 

reagents (strong acids, bases, solvents, and fluxes), high 

thermal resistance, low thermal expansion, and excellent 

cleavage and lubricity (slipperiness). In a nonoxidizing 

atmosphere, graphite remains stable to temperatures above 

3,000 °C. Graphite has a hardness of 1 to 2 on the Mohs scale 

and is thus extremely soft.

Table J2. Selected physical properties of graphite.

[Modified from Krauss and others (1988). g/cm3, gram per cubic centimeter; 

°C, degree Celsius; watt/cm °C, watt per centimeter per degree Celsius; 

Ω-cm, ohm-centimeter]

Property Description

Composition Carbon

Color Gray to black, metallic luster

Hardness (Mohs scale) 0.5 to 1 

Density (g/cm3

) 2.09 to 2.26 

Morphology Hexagonal system; perfect basal

cleavage; usually platy

Melting point About 3,550 °C in nonoxidizing

conditions; decomposes above 

600 °C in oxidizing conditions

Miscellaneous Chemically inert; nontoxic; high thermal and electrical conductivity; high

lubricity (natural lubricant)

Thermal conductivity 

(watt/cm °C)

a axis: 4.0

c axis: 0.8

Thermal expansion 

coefficient (1/ °C)

a axis: 1×10–7

c axis: 140×10–7

Electrical resistivity 

(Ω-cm×104

)

a axis: 1 to 100

c axis: 10,000
J4 Critical Mineral Resources of the United States—Graphite

Some of the major end uses of graphite are in brake 

linings, refractories, and steelmaking. Batteries, brushes for 

electrical motors, foundry materials, fuel cells, and hightemperature lubricants are additional end uses of graphite. 

Graphite forms intercalation compounds with alkali metal 

vapors—usually potassium, lithium, rubidium, and cesium—

where the metal ions fit between the planar carbon sheets of 

the graphite structure. These compounds have applications in 

the energy sector and other technologies. The industrial uses 

and associated commercial value of natural graphite depend 

on the characteristics and crystallinity of the mined graphite 

(table J1).

Fine-grained microcrystalline graphite powder (amorphous graphite) is used extensively in foundry and refractory 

applications, as a source of carbon in steelmaking, and in 

other applications where additions of graphite improve 

the manufacturing process or end product. Coarse-grained 

crystalline graphite in the form of separate lamellar crystals 

(flake graphite) has high value and is used in high-temperature 

lubricants, high-quality foundry and refractory ware, powder 

metallurgy, coatings, and battery and fuel cell applications. 

Crystalline flake graphite accounted for about 50 percent of 

natural graphite usage in the United States in the past decade.

Massive crystalline graphite from vein deposits (lump 

or chip graphite) has the highest purity and grain size of the 

natural graphite sold commercially. Because of its purity and 

crystallinity, many of the highest quality electrical motor 

brushes and other current-carrying carbon products use 

formulations of graphite from vein deposits.

Synthetic graphite is used in many applications that 

require high-purity graphite, including anticorrosion products, 

batteries, carbon brushes, coatings, conductive fillers, electrodes and electrolytic processes, fuel cell bipolar plates, and 

nuclear moderator rods. Synthetic graphite is unsuitable for 

foundry applications because of its increased porosity relative 

to natural graphite. Synthetic graphite is used in more applications in North America than natural graphite and accounts for 

a significant share of the graphite market.

Demand and Availability of Supply

No natural graphite was reported to have been mined 

in the United States in 2010 (Olson, 2012), and the major 

domestic sources of industrial graphite included imports 

of natural graphite, mostly from China, Mexico, Canada, 

Brazil, and Madagascar (in order of tonnage of imports), 

and synthetic graphite manufactured from carbonaceous 

materials. In 2010, 65,400 metric tons of natural graphite 

valued at an estimated $52 million was reported to have been 

imported into the United States for domestic consumption 

and 134,000 metric tons of synthetic graphite valued at an 

estimated $1.07 billion was reported to have been produced 

in the United States (Olson, 2012).

Graphite Prices and Pricing

Graphite is not traded on any commodity exchange. 

Commodity prices are subject to negotiation between buyer and 

seller and depend on carbon content, flake size and distribution 

(for crystalline flake), and the amount and nature of impurities. 

China is the dominant global natural graphite producer and 

has been the world price setter for decades. Four price series 

for graphite commodities based on production and value data 

reported by the U.S. Geological Survey (Olson, 2013) and 

Roskill Information Services, Ltd. (2012) are given in table J1. 

They are (a) microcrystalline (amorphous) graphite powder 

(from Mexico), which is currently in a long-term decline in 

importance because its low-technology end uses are becoming 

obsolete; (b) crystalline flake, which has become the most 

important natural graphite commodity in the past few decades 

and accounts for 80 to 90 percent of the value of U.S. imports 

of natural graphite; (c) lump graphite from Sri Lanka, which 

is a specialty product because its relatively high price results 

in low demand, and (d) synthetic graphite, which is a specialty 

product used primarily for applications that require high purity; 

its high price also limits demand. The price of microcrystalline 

(amorphous) graphite powder (80 to 85 percent carbon) ranged 

from $600 to $800 per metric ton in 2011; that of flake graphite 

(90 percent graphite) ranged from $1,150 to $2,000 per metric 

ton; Sri Lankan lump and chip graphite (99 percent graphite), 

from $1,700 to $2,070, depending on purity, grade, and 

particle size; and synthetic graphite (99.9 percent graphite), 

from $7,000 to $20,000 per metric ton, which is about four 

to seven times that of flake graphite (table J1).

In 2010, graphite prices increased more rapidly for 

crystalline flake graphite than for amorphous graphite owing 

to increased demand for crystalline graphite used in friction 

material applications, such as brake linings, high-quality 

refractories, and lubricants, and in the manufacturing of 

graphite foils and long-life alkaline batteries (Industrial 

Minerals, 2010).

Substitutes for Natural Graphite

The most common substitutes for natural graphite are 

other forms of carbon. Manufactured synthetic graphite 

powders and graphite recovered from discarded foundry and 

manufactured materials can substitute for natural graphite 

in some applications, depending on the relative price. In 

steelmaking and some foundry applications, petroleum 

coke, anthracite, and synthetic graphite can be used instead 

of natural graphite. Synthetic graphite and natural graphite 

powder are commonly substituted for each other, or mixtures 

containing both kinds are prepared for customers (Taylor, 

2006). Crystalline graphite is recovered from some foundry 

and refractory materials, although its recovery cost makes it 

noncompetitive relative to natural graphite for some applications (Weis, 1973; Olson, 2013).
Geology J5

Strategic and Critical Resource Issues

Graphite is considered a critical and strategic mineral 

because of its essential applications in the aerospace and 

energy sectors (such as refractory and foundry applications 

in the steel and metal industries and use in high-temperature 

lubricants, high-strength lightweight composite materials, 

batteries, modern nuclear reactors, fuel cells, electric motors, 

and evolving electronic applications that anticipate rapid 

growth in demand for graphite). The high concentration 

of resource supply from few countries could increase the 

commodity’s susceptibility to market manipulation or 

disruption (Krauss and others, 1988; Taylor, 2006; European 

Commission, 2014). The United States has not produced 

natural graphite since the 1950s; China, Mexico, and Canada, 

in order of import supply, are the dominant sources of graphite 

imports to the United States (Olson, 2013). China has very low 

labor and production costs, and, with aggressive marketing, 

it established dominance in the world market in the 1980s 

(Taylor, 2006; Olson, 2013). China’s graphite production 

is expected to increase and remain dominant in the world 

graphite market for the near future (Olson, 2013). China’s 

rapidly growing domestic market for graphite may restrict 

future exports, however, which has raised concerns about 

possible supply shortages of graphite, particularly crystalline 

flake graphite, needed for anticipated large-scale fuel cell 

and battery applications; such applications could dramatically increase world demand for graphite (Taylor, 2006; 

Olson, 2013; Olson and others, 2016).

Graphite has played an important role in the emerging 

noncarbon energy sector and is used in several new energy 

applications. In energy production applications, graphite is 

used in pebbles for modular nuclear reactors and in highstrength composites for wind, tide, and wave turbines. In 

energy storage applications, graphite is used in bipolar plates 

for fuel cells and flow batteries, in anodes for lithium-ion 

batteries, in electrodes for supercapacitors, in high-strength 

composites for fly wheels, in phase change heat storage, 

and in solar boilers. In energy management applications, 

graphite is used in high-performance thermal insulation and in 

silicon chip heat-dissipation applications. These new energy 

applications commonly use specialty graphite products with 

such properties as extreme purity (>99.9 percent graphite), 

very large or small (nano) particle size, and perfect crystallinity. Current graphite capacity may not be adequate for the 

increasing demands of these new energy applications, which, 

when fully implemented, may require double the current 

graphite supply, especially of high-purity crystalline flake 

graphite (O’Driscoll, 2010; Olson and others, 2016).

Geology

Geochemistry

Carbon is the fourth most abundant element in the 

solar system, and in terms of its abundance in Earth’s 

crust, is ranked about 12th to 17th in abundance, with an 

estimated crustal concentration between 180 and 270 parts 

per million. The abundance of carbon in Earth’s crust is 

difficult to determine because carbon stores exist and cycle 

between rock, sediment, atmosphere, biosphere, and ocean 

reservoirs. The behavior of carbon in its geochemical cycle is 

influenced by the form in which the carbon occurs. Most of 

the carbon in Earth’s crust (approximately 80 to 90 percent) 

is contained in carbonate minerals in carbonate rocks. Most 

of the remaining carbon in the Earth system occurs in living 

and fossil organic matter and as carbon dioxide (CO2 ) in the 

atmosphere or dissolved in the ocean. These remaining carbon 

forms dominate the carbon cycle. Graphite makes up only a 

small proportion of the carbon in Earth’s crust, probably less 

than 0.5 percent, and much of this graphite likely formed 

by high-temperature thermal alteration of organic matter 

from biogenic sources deposited in sedimentary rocks and 

subsurface reservoirs. The mineral graphite is stable and inert 

in the crustal environment and is unchanged under surface 

weathering conditions. Burial and thermal metamorphism 

tends only to recrystallize graphite. Therefore, graphite is 

largely isolated from the overall carbon cycle.

Mineralogy

Graphite is one of four forms of crystalline elemental 

carbon; the others are carbon nanotubes, diamond, and 

fullerenes. Graphite crystallizes in the hexagonal system, with 

rhombohedral symmetry, commonly forming six-sided tabular 

crystal flakes. It occurs naturally in metamorphic rocks and 

in some igneous rocks. Well-crystallized graphite flakes have 

a black metallic luster, whereas microcrystalline material is 

black and earthy with an amorphous appearance.

The crystal structure of graphite consists of carbon 

atoms tightly arranged in parallel-stacked (c axis, fig. J1), 

planar honeycomb-lattice sheets (a axis, fig. J1). When the 

graphite structure is only a 1-atom-thick planar sheet, it is 

called graphene.

The physical properties of graphite are listed in table J2. 

Many properties listed in table J2 vary by crystallographic 

orientation (a-axis and c-axis values, table J2) because of 

differences in bonding within (a axis, fig. J1) and between 

lattice sheets (c axis, fig. J1). The intraplanar (a axis) covalent 

bonding in graphite sheets is strong, but the electrostatic 

interlayer (c axis) forces holding the sheets together are 

weak, resulting in delocalized electrons that are free to move 

between sheets, which separates the sheets and allows them to 
J6 Critical Mineral Resources of the United States—Graphite

Figure J1. Diagram showing the 

arrangement of carbon atoms in graphite, 

which consists of stacks of parallel sheets 

of carbon atoms (red circles in the figure), 

with each sheet containing hexagonal arrays 

of carbon atoms. Each carbon atom within 

a sheet is connected to its three nearest 

neighbors by covalent bonds that separate 

the atoms by a distance of 1.415 angstroms 

(Å)—an angstrom is one ten-billionth of a 

meter. The stacked sheets (shown here in 

their A and B orientation) are held together 

by weak intermolecular van der Waals’ 

forces and are separated from each other 

by a distance of 3.354 Å. The arrangement of 

atoms defined by the dashed lines portrays 

one hexagonal (2H) unit cell of graphite. 

Modified from Kraus and others, 1989.

a-axis

c-axis

A

A

Graphite, fig 1

B

1.415 Å 

3.354 Å

Unit cell

slide across one another easily. This gives graphite a density of 

2.09 to 2.26 grams per cubic centimeter (g/cm 3), perfect basal 

(one-plane) cleavage, softness (0.5 to 1 on the Mohs scale), a 

greasy feel, and self-lubricating properties that make it useful 

as a lubricant. The ratio of force required to induce gliding of 

graphite sheets relative to compression force is low, resulting 

in high lubricity. These delocalized interlayer electrons also 

give graphite its high electrical and thermal conductivity. 

Graphite is the most electrically and thermally conductive of 

the nonmetallic elements (a-axis values, table J2); however, 

the intraplanar covalent bonds are very strong and require high 

energy to break them. These bonding properties make graphite 

chemically inert and physically stable at high temperatures 

(refractory). Graphite resists chemical attack by most reagents 

and is infusible in most common fluxes. Thermal oxidation of 

graphite in the presence of oxygen begins at about 300 °C, and 

the rate increases with temperature; however, it is stable in a 

nonoxidizing atmosphere to above 3,000 °C. Graphite sublimes 

between 3,300 °C and 3,500 °C at 1.033 kilograms per square 

centimeter (kg/cm2) and it melts at approximately 3,550 °C 

at a triple point under a pressure of 88 kg/cm2. Graphite is 

flexible but not elastic. All these properties combined make 

graphite desirable for many industrial applications.

Deposit Types

Natural graphite is a common constituent of metasedimentary rocks and is mined in three geologic settings: 

(a) microcrystalline (amorphous) graphite deposits formed by 

metamorphism of coal, petroleum, or carbon-rich sediments, 

(b) disseminated flakes in metamorphic rock (flake graphite), 

and (c) graphite vein deposits (lump or chip graphite). Similar 

to most industrial mineral commodities, the economics of 

a deposit depend on its location and transportation costs to 

markets in addition to deposit grade, size, and mining characteristics. Impurities in the deposit, such as silica, sulfides, or 

biotite, may limit the economic potential of a deposit.

Amorphous Graphite Deposits

Amorphous graphite deposits, which are formed by 

thermal metamorphism of coal, petroleum pools, or other 

highly carbonaceous sedimentary rocks, contain earthy to 

compact masses of microcrystalline graphite. Deposits may 

consist of multiple layers, seams, and lenses containing 

50 to 90 percent carbon; each deposit is a few meters thick 

and up to a kilometer in length and lies within clastic metasedimentary rocks. The carbon content of amorphous deposits 

is dependent on the amount of carbon present in the original 

sediments, and the grade, tonnage, and purity of the deposit 

depend upon the characteristics of the precursor coal sequence 

or carbonaceous material. Deposits in metamorphosed 

massive coal seams may contain as much as 90 percent 

graphitic carbon, whereas deposits in other carbonaceous 

sediments or impure coal sequences commonly range from 

25 to 60 percent carbon.

Chemical and isotopic biomarkers found in natural 

graphite provide evidence of its origin from ancient biological 

material. Amorphous graphite deposits occur in geologic 

settings and environments where coal and other organicrich sedimentary rocks, such as oil shale, occur, including 

(a) shoreline, fluvial-deltaic, and alluvial fan depositional 

environments in passive continental margin settings, and 

(b) lacustrine and shallow inland seas and margins in foreland, 

continental, sag, or intracontinental rift basin settings. Host 

rocks include conglomerate, metagraywacke, quartzite, and 

schist. The thermal metamorphism is commonly caused by 
Geology J7

crosscutting diabasic or granitic intrusions with associated 

hornfels alteration of host rocks. Hornfels is a hard metamorphic rock formed by contact metamorphism of sedimentary rocks close to intrusive igneous bodies.

Graphite does not form until temperatures exceed 300 °C 

to 400 °C (Landis, 1971). The degree of metamorphism 

controls the degree of graphitization, and the ratio of graphite 

to nongraphitic carbon varies, but typically increases towards 

the heat source. Mineral impurities include meta-anthracite, 

anthracite, quartz, mica, pyrite, and other sulfides. Most 

currently mined deposits typically contain at least 80 percent 

carbon in deposits exceeding 1 million metric tons.

The world’s main sources of amorphous graphite are the 

metamorphosed coal deposits in Sonora, Mexico; Siberia, 

Russia; and the large coal province stretching from China 

into the Korean peninsula (table J3). The deposits in Sonora, 

Mexico, are the most significant deposits of amorphous 

graphite in the Western Hemisphere. They occur in a series of 

beds in an area 30 kilometers (km) long by 15 km wide located 

60 km southeast of Hermosillo. In the district, the orebodies 

average 3 m in thickness, but deformation can produce seams 

up to 7 m wide. Grades exceed 75 percent contained graphite 

and some contain as much as 95 percent. The Moradillas 

deposit (Lourdes area, Mexico) has been reinterpreted as 

formed from a graphitized hydrocarbon protolith owing to its 

vein-like structure and lack of evidence that it is metamorphosed coal (Taylor, 2006). In the United States, small amorphous graphite deposits occur in metamorphosed coal deposits 

in the Narragansett Basin, Rhode Island, and in deposits in 

northern Michigan. An additional U.S. example occurs at 

Raton, New Mexico, where a basalt dike intersects a coal seam.

Deposits of Flake Graphite Disseminated in 

Metasedimentary Rocks

A large proportion of worldwide graphite production is 

derived from deposits of disseminated flake graphite found 

in metamorphic rocks, such as paragneiss (a gneiss derived 

from sedimentary rock), quartzite, and, sometimes, marble. 

These deposits formed when preexisting carbonaceous matter 

in the host sedimentary rock was converted to graphite 

during regional metamorphic conditions at or exceeding 

medium-pressure and medium-to-high-temperature conditions (amphibolite facies). These metamorphic conditions 

are sufficient to crystallize fully ordered graphite (Hoefs and 

Frey, 1976) and recrystallize the rock matrix. The precursor 

host rocks of these deposits are interpreted as occurring in 

depositional settings where organic-rich sediments accumulate 

and are preserved. These settings include sediment-starved 

intracratonic and continental margin basins with low-oxygen 

conditions at depth to accumulate organic sediments, and 

periods of transgression (compare with, for example, Wilde 

and others, 1999), when sea level rises relative to land, 

preserving organic-rich sediments without erosion. The 

sedimentary rocks that are deposited during transgression may 

be deposited directly on existing basement rocks of the crust 

and are termed supracrustal rocks.

The metamorphic conditions sufficient to form these 

deposits occur in the roots of continental mountain belts 

formed by deformation of Earth’s crust, and significant postmetamorphic uplift and erosion are required to expose deposits 

in minable settings. Alumina-rich paragneiss, quartzites, and 

marbles at or near the highest grade of regional metamorphism 

at medium pressure (granulite facies) are favorable host rocks 

for flake graphite deposits because of their granular texture, 

coarse grain size, and low mica content. Supracrustal metasedimentary belts with these metamorphic features are termed 

crystalline metamorphic basement; their worldwide distribution has been mapped by Chorlton (2007). Most flake graphite 

deposits are located in Precambrian crystalline metamorphic 

basement (fig. J2), principally of Neoarchean to Proterozoic 

eras; however, deposits could occur in crystalline basement of 

any age. Precambrian supracrustal metamorphic belts appear 

to be more extensive, and metamorphosed to higher grades, 

than Phanerozoic belts (Chacko and others, 1987).

The principal flake graphite deposits occur as stratabound lenses or layers; individual lenses in deposits are as 

much as tens of meters thick and hundreds of meters long. 

The lenses have variable graphite content, both within 

themselves and from one lens to another. The graphite content 

of a typical deposit varies from about 8 to 15 percent carbon, 

but the grade can vary from as low as 3 percent to about 

60 percent carbon locally in individual deposits and between 

deposits. In Madagascar, one rich lens was reported to contain 

60 percent graphite, and grades in the Kigluaik Mountains 

graphite district, Alaska, exceed 50 percent graphite in some 

areas (Coats, 1944). In paragneiss-hosted deposits, gangue 

(non-ore) minerals include in order of general abundance, 

feldspar, biotite, pyroxene, garnet, and sulfide minerals. 

Gangue minerals in carbonate-hosted deposits include calcite, 

pyroxene, feldspar, garnet, and sulfides. Flake graphite 

in crushed ore is separated from mineral impurities using 

washing, screening, flotation, and, sometimes, leach methods.

The highest graphite grades commonly are associated 

with rock contacts between marble and paragneiss or quartzite, 

lenses in fault zones, in segregations within fold crests, and 

in structures that acted as channels for metamorphic fluids, all 

suggesting structural control of mineralization. Although most 

carbon in these deposits is thought to be present in the original 

sedimentary rocks, some carbon enrichment may be caused by 

processes involving internal or external buffering or mixing of 

metamorphic fluids (Lamb and Valley, 1984; Newton, 1986).

 Crystals in each deposit vary from a fraction of a millimeter to a few centimeters in size, usually reflecting the grain 

size in the parent rock. Most currently mined flake graphite 

deposits typically contain at least 8 to 12 percent graphitic 

carbon in deposits exceeding 0.5 million metric tons. Graphite 

is stable in the weathering environment; deposit grades and 

ease of mining are enhanced by weathering destruction of 

other minerals. Some deposits become so weathered that they 

can be mined with shovels and other hand tools.
J8 Critical Mineral Resources of the United States—Graphite

Table J3. Location, grade, tonnage and other data for selected graphite deposits and districts of the world.—Continued

[The names, locations, and descriptions of most of the deposits are taken from the International Strategic Minerals Inventory (ISMI) graphite inventory 

(Krauss and others, 1988) and the U.S. Geological Survey Mineral Resources Data System (MRDS; U.S. Geological Survey, 2012). Identification number is 

keyed to the deposits shown in figure J2. Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Record types are from the MRDS. Ore type is characterized as amorphous, crystalline, flake, or lump 

(includes chip), as defined in table J1. “Weathering” indicates surface ores that have been deeply weathered, which lowers mining costs. Development status 

of the deposit is given as follows: Current, current or recent mine activity; Past, past producer with no current mining; Prospect, an unmined deposit with 

characterized resources; Occurrence, an unmined deposit with uncharacterized resources. Tonnage of ore has been updated from the values in the MRDS and 

is given in metric tons or, alternatively, is categorized as large (likely exceeds 1 million metric tons but is not well characterized) or small (likely is less than 

100,000 metric tons). %, percent; —, unknown; WGS 84, World Geodetic System of 1984]

Identification 

number Name Country State/Province

Latitude Longitude 

Decimal degrees WGS 84 

1 Quebrada Del Gato Argentina San Juan –31.43 –68.13 

2 Undercliff Australia New South Wales –28.66 152.21 

3 Jack’s Creek Australia Queensland –20.67 147.87 

4 Koppio-Uley mines Australia South Australia –34.80 135.71 

5 Munglinup River Australia Western Australia –33.50 120.85 

6 Doppl-Muehldorf-Zettlitz Austria Niederosterreich 48.38 15.45 

7 Kaisersberg-Trieben Austria Steiermark 47.35 15.07 

8 Itanhem Brazil Bahia –17.10 –40.35 

9 Itapacerica Brazil Minas Gerais –20.43 –45.13 

10 Pedra Azul Brazil Minas Gerais –15.88 –45.13 

11 Bisset Creek Canada Ontario 46.23 –78.07 

12 Coronation Canada Ontario 45.29 –77.94 

13 Graphite Lake-Ryerson-Todd area Canada Ontario 45.73 –79.08 

14 Kirkham-Portland-Timmins area Canada Ontario 44.56 –76.57 

15 Tagliamonte Canada Ontario 46.23 –78.07 

16 Graphex-Graphico-Orwell Mines Canada Quebec 46.14 –75.55 

17 North American Mine Canada Quebec 45.52 –75.55 

18 St-Amime-Lac Des Iles Canada Quebec 46.25 –75.53 

19 Lac Knife Canada Quebec 52.55 –61.18 

20 Pollon Lake area Canada Saskatchewan 56.38 –103.13 

21 Haikou China Hainan 20.08 110.42 

22 Heling China Heilongjiang

[Heilungkiang]

46.31 129.55 

23 Jixi (Liu Mao) China Heilongjiang

[Heilungkiang]

45.28 131.00 

24 Liu Mao China Heilongjiang

[Heilungkiang]

45.30 131.07 

25 Honan deposits China Henan [Honan] 32.50 113.88 

26 Hunan China Hunan 26.00 113.00 

27 Panshi China Jilin [Kirin] 42.93 125.97 

28 Hohot China Nei Mongol

(Inner Mongolia)

40.81 111.62 

29 Shandong Peninsula China Shandong [Shantung] 37.03 120.32 

30 Xing He China — 40.88 113.88 

31 Kolledeye Czech Republic — 49.22 14.45 

32 Velke Vbrno-Konstantin Vvk Czech Republic — 50.13 17.33 

33 Kropfmühl-Cesky Krumlov Germany-Czech 

Republic

Bayern [Bavaria, 

Germany]

48.80 14.32 
Geology J9

Table J3. Location, grade, tonnage and other data for selected graphite deposits and districts of the world.—Continued

[The names, locations, and descriptions of most of the deposits are taken from the International Strategic Minerals Inventory (ISMI) graphite inventory 

(Krauss and others, 1988) and the U.S. Geological Survey Mineral Resources Data System (MRDS; U.S. Geological Survey, 2012). Identification number is 

keyed to the deposits shown in figure J2. Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Record types are from the MRDS. Ore type is characterized as amorphous, crystalline, flake, or lump 

(includes chip), as defined in table J1. “Weathering” indicates surface ores that have been deeply weathered, which lowers mining costs. Development status 

of the deposit is given as follows: Current, current or recent mine activity; Past, past producer with no current mining; Prospect, an unmined deposit with 

characterized resources; Occurrence, an unmined deposit with uncharacterized resources. Tonnage of ore has been updated from the values in the MRDS and 

is given in metric tons or, alternatively, is categorized as large (likely exceeds 1 million metric tons but is not well characterized) or small (likely is less than 

100,000 metric tons). %, percent; —, unknown; WGS 84, World Geodetic System of 1984]

Identification 

number

Record 

type Ore type Operation Development status Tonnage Grade (% of carbon)

1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

16

17

18

19

20

21

22

23

24

25

26

27

28

29

30

31

32

33

District

Site

Site

District

District

District

District

Site

Site

Site

Site

Site

District

District

Site

District

Site

Site

Site

District

Site

Site

Site

Site

District

District

Site

District

District

Site

District

Site

District

Crystalline (flake 

or lump?)

Amorphous

Amorphous

Flake, weathering

Flake

Flake

Amorphous

Lump

Flake, weathering

Flake, weathering

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Amorphous

Amorphous

Flake

Unknown

Amorphous

Amorphous

Crystalline (flake 

or lump?)

Flake

Flake

Flake

Amorphous

Flake

Unknown

Surface

Surface,underground

Surface,underground

Surface,underground

Surface

Underground

Surface

Surface

Surface

Surface

Unknown

Surface

Surface

Surface

Surface

Unknown

Surface

Surface

Unknown

Surface

Unknown

Surface

Unknown

Unknown

Underground

Underground

Surface

Surface

Unknown

Surface,underground

Surface

Underground

Producer

Past producer

Past producer

Past producer

Past producer

Producer

Producer

Producer

Producer

Producer

Prospect

Occurrence

Producer

Producer

Prospect

Producer

Past producer

Producer

Prospect

Past producer

Producer

Producer

Past producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Prospect

Producer

Producer

—

793

2,200

35,030

30,000

1,000,000

1,000,000

2,778

383,000

26,800,000

4,938,000

—

Large

478,000

—

2,200,000

—

—

4,900,000

1,663,000

5,000,000

—

300,000,000

3,000,000

—

—

—

—

5,900,000

—

—

200,000

—

—

32

61

15

25

10

85

40

15.7

11.9

15.58

—

—

9

—

10

—

—

15.8

10.3

3.7

—

85

18.5

—

—

—

—

5

—

—

50

16
J10 Critical Mineral Resources of the United States—Graphite

Table J3. Location, grade, tonnage and other data for selected graphite deposits and districts of the world.—Continued

[The names, locations, and descriptions of most of the deposits are taken from the International Strategic Minerals Inventory (ISMI) graphite inventory 

(Krauss and others, 1988) and the U.S. Geological Survey Mineral Resources Data System (MRDS; U.S. Geological Survey, 2012). Identification number is 

keyed to the deposits shown in figure J2. Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Record types are from the MRDS. Ore type is characterized as amorphous, crystalline, flake, or lump 

(includes chip), as defined in table J1. “Weathering” indicates surface ores that have been deeply weathered, which lowers mining costs. Development status 

of the deposit is given as follows: Current, current or recent mine activity; Past, past producer with no current mining; Prospect, an unmined deposit with 

characterized resources; Occurrence, an unmined deposit with uncharacterized resources. Tonnage of ore has been updated from the values in the MRDS and 

is given in metric tons or, alternatively, is categorized as large (likely exceeds 1 million metric tons but is not well characterized) or small (likely is less than 

100,000 metric tons). %, percent; —, unknown; WGS 84, World Geodetic System of 1984]

Identification 

number Name Country State/Province

Latitude Longitude 

Decimal degrees WGS 84 

34 East Godavari India Andhra Pradesh 17.38 81.85 

35 Khammam India Andhra Pradesh 17.38 81.33 

36 Srikakulam India Andhra Pradesh 18.33 83.10 

37 Visakhapatnam India Andhra Pradesh 18.02 82.93 

38 Khamdih India Bihar 23.97 84.22 

39 Sokra India Bihar 23.97 84.13 

40 Dandatapa area India Odisha 20.80 84.60 

41 Sargipali area India Odisha 20.90 83.08 

42 Titlagarh area India Odisha 20.20 83.37 

43 Tumdibandh-Phulbani area India Odisha 19.83 83.63 

44 Tamatia mines India Rajasthan 23.62 74.50 

45 Madurai India Tamil Nadu [Madras] 10.07 77.83 

46 Chawia Kenya — –3.46 38.38 

47 Oldoinyo-Nyiro Kenya — 0.75 37.00 

48 Yonghung-Kaechon area Korea, North Hamgyong-Namdo 39.83 127.43 

49 Songjin deposits Korea, North Hwanghae-Namdo 40.67 129.20 

50 Wolmyong-Kaerim-Pongmyong Korea, Republic of Ch’Ungch’Ong-Bukto 36.41 127.75 

51 Yongwon-GunJa-Pyongtack-Shihung Korea, Republic of Ch’Ungch’Ong-Namdo 36.91 127.67 

52 Ampangadiatany Madagascar Toamasina –19.41 48.89 

53 Ambatomitamba Sahanovo area Madagascar Toamasina –18.35 49.10 

54 Perinet-Ambatovy area Madagascar — –18.93 48.45 

55 Vatomandry area Madagascar — –19.34 48.94 

56 Telixtlahuaca Mexico Oaxaca 17.33 –96.86 

57 Las Trincheras Mexico Sonora 30.40 –111.53 

58 Lourdes area Mexico Sonora 28.60 –110.50 

59 Tonichi Mexico Sonora 28.58 –109.57 

60 Skaland-Senja Norway — 69.45 17.29 

61 La Galgada-La Limena-San Carlos Mines Peru Ancash –8.46 –78.13 

62 Baia De Fier Romania — 45.23 23.75 

63 Botogolsk Russia Buryatiya 52.46 100.75 

64 Boyarsk Russia Buryatiya 51.85 106.10 

65 Kureyka Russia Krasnodarskiy Kray 66.48 87.17 

66 Noginskoje Russia Krasnodarskiy Kray 64.50 91.25 

67 Soyusnoye Russia — 47.91 130.93 

68 Tayginsk Russia — 55.63 60.65 

69 Gumbu graphite mine South Africa — –22.32 30.67 

70 Malonga graphite mine South Africa — –22.65 30.88 

71 Kahatagaha-Kolongaha mines Sri Lanka — 7.57 80.53 
Geology J11

Table J3. Location, grade, tonnage and other data for selected graphite deposits and districts of the world.—Continued

[The names, locations, and descriptions of most of the deposits are taken from the International Strategic Minerals Inventory (ISMI) graphite inventory 

(Krauss and others, 1988) and the U.S. Geological Survey Mineral Resources Data System (MRDS; U.S. Geological Survey, 2012). Identification number is 

keyed to the deposits shown in figure J2. Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Record types are from the MRDS. Ore type is characterized as amorphous, crystalline, flake, or lump 

(includes chip), as defined in table J1. “Weathering” indicates surface ores that have been deeply weathered, which lowers mining costs. Development status 

of the deposit is given as follows: Current, current or recent mine activity; Past, past producer with no current mining; Prospect, an unmined deposit with 

characterized resources; Occurrence, an unmined deposit with uncharacterized resources. Tonnage of ore has been updated from the values in the MRDS and 

is given in metric tons or, alternatively, is categorized as large (likely exceeds 1 million metric tons but is not well characterized) or small (likely is less than 

100,000 metric tons). %, percent; —, unknown; WGS 84, World Geodetic System of 1984]

Identification 

number

Record 

type Ore type Operation Development status Tonnage Grade (% of carbon)

34

35

36

37

38

39

40

41

42

43

44

45

46

47

48

49

50

51

52

53

54

55

56

57

58

59

60

61

62

63

64

65

66

67

68

69

70

71

District

District

District

District

District

District

District

District

District

District

District

District

District

Site

District

District

District

District

Site

District

District

District

Site

Site

District

Site

Site

District

Site

Site

District

Site

Site

Site

Site

Site

Site

District

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake, lump

Flake

Flake, lump

Flake, weathering

Flake, weathering

Amorphous

Flake

Amorphous

Flake, weathering

Flake, weathering

Flake, weathering

Flake, weathering

Flake, weathering

Flake, weathering

Amorphous

Amorphous

Amorphous

Flake

Amorphous

Amorphous

Lump, flake

Flake

Amorphous

Amorphous

Flake

Flake

Flake

Amorphous

Lump

Surface

Surface,underground

Unknown

Surface

Surface

Surface

Surface

Surface

Surface

Surface

Surface

Surface

Surface

Surface

Surface,underground

Unknown

Underground

Surface

Surface

Surface

Surface

Surface

Surface

Underground

Underground

Unknown

Underground

Unknown

Surface,underground

Underground

Unknown

Surface,underground

Underground

Surface

Surface

Unknown

Surface

Underground

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Past producer

Prospect

Producer

Producer

Producer

Producer

Past producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Producer

Past producer

Past producer

Producer

Past producer

Producer

Producer

Producer

Producer

Past producer

Past producer

Producer

—

—

—

—

—

—

—

—

—

—

1,120,000

130,000

—

1,200,000

—

—

12,000,000

1,400,000

—

2,000,000

—

2,000,000

13,400,000

—

2,000,000

2,000,000

2,000,000

—

58,000

Small

Large

87,000,000

10,000,000

8,200,000

1,116,000

—

20,000,000

100,000

—

—

—

—

—

—

—

—

—

—

14.1

19.5

—

13

—

—

80

3.5

—

9

—

7

4

—

80

75

30

—

70

—

5

88

85

18

3

—

63.5

60
J12 Critical Mineral Resources of the United States—Graphite

Table J3. Location, grade, tonnage and other data for selected graphite deposits and districts of the world.—Continued

[The names, locations, and descriptions of most of the deposits are taken from the International Strategic Minerals Inventory (ISMI) graphite inventory 

(Krauss and others, 1988) and the U.S. Geological Survey Mineral Resources Data System (MRDS; U.S. Geological Survey, 2012). Identification number is 

keyed to the deposits shown in figure J2. Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Record types are from the MRDS. Ore type is characterized as amorphous, crystalline, flake, or lump 

(includes chip), as defined in table J1. “Weathering” indicates surface ores that have been deeply weathered, which lowers mining costs. Development status 

of the deposit is given as follows: Current, current or recent mine activity; Past, past producer with no current mining; Prospect, an unmined deposit with 

characterized resources; Occurrence, an unmined deposit with uncharacterized resources. Tonnage of ore has been updated from the values in the MRDS and 

is given in metric tons or, alternatively, is categorized as large (likely exceeds 1 million metric tons but is not well characterized) or small (likely is less than 

100,000 metric tons). %, percent; —, unknown; WGS 84, World Geodetic System of 1984]

Identification 

number Name Country State/Province

Latitude Longitude 

Decimal degrees WGS 84 

72 Zavalyevskiy Ukraine — 48.20 30.03 

73 Clay-Coosa-Chilton graphite district United States Alabama 33.28 –85.84 

74 Christophosen Creek United States Alaska 65.03 –165.63 

75 Graphite Creek United States Alaska 65.04 –165.54 

76 Imuruk Basin Graphite United States Alaska 65.04 –165.55 

77 Kigluaik Mountains graphite district United States Alaska 65.04 –165.54 

78 Southern California area United States California 34.65 –118.47 

79 Black Lady Nos. 1 and 2 United States California 36.90 –118.07 

80 Eureka Plumbago-Morgan mines United States California 38.02 –120.38 

81 Skinner Ranch United States California 38.19 –122.59 

82 Unnamed United States California 36.28 –121.56 

83 Unnamed United States California 41.21 –123.77 

84 Wedge United States California 32.80 –115.98 

85 Graphite Basin United States Colorado 38.59 –106.38 

86 Unnamed graphite deposit United States Colorado 38.67 –105.98 

87 Graphite mines United States Connecticut 41.86 –73.37 

88 Lead Hill Mine graphite deposits United States Connecticut 41.94 –72.18 

89 Unnamed graphite mine United States Georgia 34.12 –84.74 

90 Detroit-Northern Graphite quarries United States Michigan 46.63 –88.35 

91 Black Diamond Carbon Graphite United States Montana 46.32 –111.63 

92 Crystal Graphite Mine United States Montana 45.10 –112.51 

93 Boston-Goshen-Osgood Mines area United States New Hampshire 43.27 –72.10 

94 Stone Mills Graphite Mine United States New Jersey 40.68 –74.88 

95 Adirondack Mining & Milling Mine United States New York 43.56 –73.46 

96 Essex-Warren area United States New York 43.90 –73.47 

97 Champlain Graphite Mine United States New York 43.54 –73.48 

98 M. B. Hooper Graphite Mine United States New York 43.55 –73.50 

99 Pope Mills-Rossie deposits United States New York 44.47 –75.55 

100 Rowland Graphite Mine area United States New York 43.60 –73.97 

101 Silver Leaf Graphite Co. Mine United States New York 43.59 –73.43 

102 Dillinger Mine and mill United States North Carolina 35.90 –82.18 

103 Cranston-Penners Ledge Mines United States Rhode Island 41.77 –71.46 

104 Copper Cliff Mine United States South Dakota 44.12 –103.84 

105 Burnet-Llano district United States Texas 30.78 –98.36 

106 Rabbit Creek area United States Wyoming 42.21 –105.23 

107 Taskazgan Uzbekistan — 40.81 63.38 

108 Lynx Mine Zimbabwe — –16.43 29.30 
Geology J13

Table J3. Location, grade, tonnage and other data for selected graphite deposits and districts of the world.—Continued

[The names, locations, and descriptions of most of the deposits are taken from the International Strategic Minerals Inventory (ISMI) graphite inventory 

(Krauss and others, 1988) and the U.S. Geological Survey Mineral Resources Data System (MRDS; U.S. Geological Survey, 2012). Identification number is 

keyed to the deposits shown in figure J2. Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Record types are from the MRDS. Ore type is characterized as amorphous, crystalline, flake, or lump 

(includes chip), as defined in table J1. “Weathering” indicates surface ores that have been deeply weathered, which lowers mining costs. Development status 

of the deposit is given as follows: Current, current or recent mine activity; Past, past producer with no current mining; Prospect, an unmined deposit with 

characterized resources; Occurrence, an unmined deposit with uncharacterized resources. Tonnage of ore has been updated from the values in the MRDS and 

is given in metric tons or, alternatively, is categorized as large (likely exceeds 1 million metric tons but is not well characterized) or small (likely is less than 

100,000 metric tons). %, percent; —, unknown; WGS 84, World Geodetic System of 1984]

Identification 

number

Record 

type Ore type Operation Development status Tonnage Grade (% of carbon)

72

73

74

75

76

77

78

79

80

81

82

83

84

85

86

87

88

89

90

91

92

93

94

95

96

97

98

99

100

101

102

103

104

105

106

107

108

Site

District

Site

Site

Site

District

District

Site

District

Site

Site

Site

Site

District

Site

Site

Site

Site

Site

Site

Site

District

Site

Site

District

Site

Site

District

District

Site

Site

District

Site

District

Site

Site

Site

Flake, weathering

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Flake

Lump or flake

Lump or flake

Flake

Amorphous

Flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Lump or flake

Flake

Amorphous

Flake

Flake

Flake

Flake

Flake

Surface

Unknown

Unknown

Unknown

Surface

Unknown

Surface

Surface

Unknown

Surface

Unknown

Unknown

Unknown

Underground

Unknown

Surface

Surface

Surface

Surface

Surface

Unknown

Surface

Unknown

Surface

Surface-Underground

Surface

Underground

Underground

Surface-Underground

Surface

Surface

Unknown

Unknown

Unknown

Underground

Surface

Underground

Producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Past producer

Producer

Producer?

100,000,000

300,000

65,000

—

—

10,000,000

—

—

—

—

—

—

—

—

—

Small

Small

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

400,000

—

2,300,000

6,700,000

5.5

10

52

—

—

10

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

5

—

15

26.2
J14 Critical Mineral Resources of the United States

—Graphite

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Crystalline metamorphic bedrock

EXPLANATION

Graphite, fig 02

Amorphous

Graphite deposit and identifier

Crystalline

1

2

3

4

5

6 7

8

9

10

11–18

19 20

22

21

23

24

25

26

27

28 29

30

33

31 32

34

35 36

38

37

39

40

41 43 42 44

45

46

47

49

48

50 51

53

52

54

55

56

57

58 59

60

61

61

62

63 64

65 66

67

68

69

70

71

72

73

74757677

78

78

79

80 81

82

83

84

85

86

90

89

91 92

94

93, 95–101

102 103

104

105

106 107

108

87–88

ifier number in table J3. World map showing locations of major graphite deposits and districts in the world, by commodity type. Numbers refer to the ident Figure J2.

Distribution of crystalline metamorphic basement is from Chorlton (2007).
Geology J15

The world’s main sources of crystalline flake graphite are 

the deposits in Brazil, Canada, China, India, and Madagascar. 

Probably the largest resources of high-grade crystalline 

flake graphite in the world are in China (Jixi district) and 

the island country of Madagascar. The deposits in both areas 

occur in belts of Neoarchean to late Proterozoic micaceous 

gneiss and schist that are interpreted to be associated with a 

series of mountain-building events related to the formation 

of the Gondwana supercontinent about 600 million years ago 

(Wilde and others, 1999), named the Pan-African orogeny. 

Pan-African age supracrustal metamorphic belts also occur in 

areas in Western Australia, eastern India, and Sri Lanka; these 

areas also host significant flake graphite deposits located in the 

most productive graphite-bearing regions of the world (fig. J2).

Graphite-rich horizons in Madagascar occur over a 

distance of more than 800 km in the eastern half of the 

island. Individual deposits are graphite-rich layers that 

range from 3 m to more than 35 m in thickness and extend 

up to a few kilometers. Ore grades average 4 to 10 percent 

graphite, but grades are reported to be as high as 60 percent 

in some areas. The flake graphite deposits in the State of 

Minas Gerais, Brazil, are mined from graphite-bearing soils 

overlying extensively weathered host rock with grades of 

12 to 16 percent graphite; the ore is mined by front-end 

loader or shovel. Shandong Province in China is a major 

flake graphite producer. The flake graphite deposits in the 

vicinity of Jixi, China, are mined from deformed lenses and 

discordant layers of graphite schist and gneiss in paragneiss; 

graphite content ranges from 15 percent in the gneiss to up to 

45 percent in the schists, with local flake graphite resources 

exceeding 360 million metric tons (Wilde and others, 1999). 

Little is known about the deposits in Russia and India.

Large disseminated flake graphite deposits occur at a 

number of places in the Grenville Province metamorphic belt 

in Canada’s Quebec and southeastern Ontario Provinces, and 

small deposits associated with Grenville series marble, gneiss, 

and pegmatites also occur in the Adirondack Mountains 

west and northwest of Ticonderoga, New York (Alling, 

1918). The Grenville Province is a southwest-trending belt 

of deformed high-grade supracrustal metamorphic rocks of 

Mesoproterozoic age that is exposed across 2,000 km from 

Labrador, through Quebec, into southeastern Ontario, and 

continues in the subsurface a further 3,000 km to Texas and 

Mexico. The Lac Knife graphite deposit at Fermont, Quebec, 

has high-grade ore hosted by migmatized quartz-feldspar 

(biotite) gneiss. The host rock is the southern extension of 

carbonaceous shales and sandstones that have been deformed 

and metamorphosed in the crosscutting Grenville Province 

metamorphic belt.

Deposits of flake graphite in the United States have 

been mined in Alabama, Alaska, and Texas. The deposits of 

the Clay-Coosa-Chilton graphite district, Alabama, occur 

in two parallel layers and lenses in the Ashland quartz-mica 

schist that contain 1 to 5 percent disseminated flake graphite 

(Pallister and Thoenin, 1948). The layers and lenses range 

from 6 m to more than 35 m in thickness, dip gently, and occur 

in a northeast-trending 1- to 3-km-wide belt that extends for 

more than 90 km. The deposits are weathered to depths of 

35 m. The deposits in the Burnet-Llano district, Texas, occur 

in the Precambrian Packsaddle schist and show a wide range 

of graphite content and flake size (Needham, 1946). The 

graphite flake deposits in the Kigluaik Mountains graphite 

district, Alaska (Coats, 1944) represent the largest known 

domestic graphite resource, but are located in a rugged and 

remote area with high mining costs.

Vein Deposits Containing Lump or Chip Graphite

Crystalline graphite of the lump or chip commodity type 

is generally found in well-defined veins composed almost 

entirely of graphite that cut high-grade metamorphic rocks or 

as accumulations along intrusive contacts of pegmatites with 

marbles or paragneiss. The veins form in high-grade regional 

metamorphic environments where metasedimentary belts are 

invaded by igneous rocks. In some areas, the veins are localized in granulite-facies domains within the broader regional 

metamorphic belt (Chamberlain and Rumble, 1988).

Graphite-bearing veins form within or crosscut metamorphic structures, rock contacts, and folds. The veins range 

from thin films to massive bodies that are more than 3 m thick 

and may extend for hundreds of meters, although they usually 

extend for less than tens of meters. The veins consist largely 

of coarse, platy, or needlelike interlocking crystals of graphite. 

Mineral impurities depend largely upon the host rock; quartz, 

feldspar, pyroxene, and calcite are common. Ore grades 

range from 40 to more than 90 percent graphitic carbon; 

the percentage largely depends on the amount of wallrock 

contained with the ore material. Hand sorting may be required 

to provide lump concentrates of high purity.

Graphite vein deposits are found in similar settings as 

disseminated flake graphite deposits, and they may be spatially 

associated with them. The Sri Lankan graphite deposits, 

which are the only economically significant examples of the 

vein-type deposits, occur in high-grade metamorphic rocks 

of Neoarchean to Proterozoic age. Most vein deposits and 

prospects occur in crystalline metamorphic basement rocks of 

Neoarchean to late Proterozoic age, although deposits may be 

of any age. The graphite in these deposits occurs as veins in 

fractures and structures that are emplaced near or after peak 

metamorphic conditions.

It is clear that the carbon in the vein deposits has been 

transported by metamorphic or replacement processes, 

presumably aided by metamorphic or other fluids; however, 

the nature of the processes responsible for the precipitation of 

graphite in the veins is uncertain and controversial, and may 

vary between deposits. Most veins are hosted in high-grade 

metamorphic rocks; however, graphite precipitation may 

take place after the most intense deformation and thermal 

conditions of regional metamorphism have been achieved, as 

suggested by the localization of veins in brittle structures that 

are not disturbed by penetrative deformation associated with 

metamorphism and veins that crosscut syntectonic intrusives 
J16 Critical Mineral Resources of the United States—Graphite

(Rumble and Hoering, 1986; Rumble and others, 1986). In 

some veins, low-temperature minerals, such as prehnite, occur 

with graphite. Textural evidence and age determinations on 

zircon inclusion rims in graphite (Zeitler and others, 1990), 

however, indicate that some veins were emplaced during or 

soon after peak metamorphic conditions and that the hydrothermal systems responsible for vein formation may also be 

responsible for the high-grade metamorphic conditions found 

in their immediate vicinity (Chamberlain and Rumble, 1988). 

Carbon isotope studies of the vein deposits show a wide 

range in isotopic composition between and within districts; 

this variation may be explained by varying carbon sources 

from two crustal reservoirs: (a) reduced biogenic carbon, and 

(b) carbonate minerals (Rumble and Hoering, 1986). The most 

probable processes resulting in graphite precipitation in veins 

likely include rock-fluid redox reactions, such as oxidation 

of methane (CH4 )-bearing fluids or reduction of CO2-bearing 

fluids by wall rock (internal or external buffering), cooling 

of hot fluids saturated with respect to graphite, and mixing 

of fluids with differing CH4 and CO2 contents (Rumble and 

others, 1986).

The United States has small graphite vein deposits in 

New Hampshire, Connecticut, the Adirondacks of New York, 

and Montana; these deposits are largely of only scientific 

interest. The Crystal Graphite Mine near Dillon, Montana, is 

the largest known graphite vein deposit in the United States. 

At this deposit, veins up to 60 cm thick and 15 m long occur in 

fractures in gneiss and pegmatite (Bastin, 1912); the fractures 

and veins formed after the peak of metamorphism in these 

rocks (Ford, 1954).

Mining and Beneficiation Methods

Natural graphite is mined from both open pit and 

underground mine operations (table J3). Production from open 

pit operations is less expensive and is preferred where the 

overburden can be removed economically. Most crystalline 

flake deposits are mined using open pit methods; the exception is some steeply dipping bodies with high-grade minable 

lenses containing 15 percent or more contained carbon 

(table J3). Graphite is stable in the weathering environment, 

and intensely weathered crystalline flake deposits, such as 

those in Brazil, Kenya, Madagascar, and Oaxaca, Mexico, 

can be easily and profitably mined using open pit methods. 

Underground mining methods are used in some amorphous 

deposits in China, Europe, the Republic of Korea, and Mexico, 

and vein deposits in Sri Lanka, where the ore intervals are 

deep and of high grade (>80 percent contained carbon). 

Beneficiation processes for graphite vary from simple hand 

sorting and screening of high-grade ore at some amorphous 

graphite deposits and at the Sri Lankan vein operations to 

multistage crushing, screening, washing, and flotation cycles 

required to produce high-quality and high-purity graphite flake 

and powder products. No refining of amorphous graphite ore 

is needed for most uses.

Resources and Production

Distribution of Graphite Deposits and Districts

The world map in figure J2 shows the location of 

108 natural graphite deposits and districts characterized as: 

(a) amorphous deposits that result from thermal or contact 

metamorphism of highly carbonaceous sediments, usually 

graphitized coal beds, or (b) crystalline deposits of either 

disseminated flake graphite, which have resulted from regional 

metamorphism of carbonaceous sediments under conditions 

exceeding amphibolite facies, or coarse graphite aggregates 

(lump or chip) in fracture-filling veins cutting igneous and 

metamorphic rocks. Both of the crystalline deposit types 

(flake, or lump and chip) are mostly located in crystalline 

metamorphic bedrock (fig. J2), consisting primarily of coarsegrained quartzofeldspathic metasedimentary rocks at high 

metamorphic grades (Chorlton, 2007).

Table J3 describes the features of the 108 natural graphite 

deposits and districts shown in figure J2 that are known to have 

produced graphite or are significant prospects with potential 

for future development. A number of deposits and districts 

with past production, particularly those in the United States, 

have grade and tonnage characteristics that render the deposits 

subeconomic under current conditions; these are included to 

identify broad areas and regions that may be future sources 

of graphite production or may contain undiscovered deposits. 

The names, locations, and descriptions of most of the deposits 

in table J3 are taken from the International Strategic Minerals 

Inventory (ISMI) graphite inventory (Krauss and others, 1988) 

and the U.S. Geological Survey Mineral Resources Data 

System (MRDS) (U.S. Geological Survey, 2012). Figure J3 

shows the deposit grade (percent carbon) and tonnage reported 

for some of the deposits in table J3, characterized by deposit 

type. The lowest tonnage and grade deposits in figure J3 are 

0 

10 

20 

30 

40 

50 

60 

70 

80 

90 

100 

1 10 100 1,000 10,000 100,000 1,000,000 

Grade, in percent carbon 

Deposit tonnage, in thousand metric tons

Crystalline 

Amorphous 

Deposit type

EXPLANATION

Graphite, fig 3.

Figure J3. Plot of grade (carbon) and tonnage for some 

of the amorphous and crystalline graphite deposits listed 

in table J3, by deposit type. The table also gives additional 

resource data, and sources.
Resources and Production J17

subeconomic under current market conditions; typical ranges 

of commercial grades and deposit tonnage, by commodity 

type, are given in table J1.

World Graphite Production and Resources

Natural graphite is produced by more than 20 nations, but 

world production is dominated by China, India, and Brazil. 

Figure J4 shows the average level and amounts of natural 

graphite produced worldwide by area from 2006 to 2010, 

based on data from Olson (2011). During this period, China 

accounted for approximately 67 percent of worldwide production of natural graphite, and established itself as the dominant 

exporter with the ability to set world prices. China became a 

major global supplier of graphite in the 1980s owing in part to 

very low production costs and aggressive marketing. India and 

Brazil are significant worldwide exporters of graphite whereas 

Mexico and Canada export mainly to the United States.

Global graphite resources are large relative to annual 

global consumption, and undoubtedly extensive, but their 

extent is poorly known because resources of industrial mineral 

commodities typically are not fully delineated and reported far 

in advance of development. Complete estimates of graphite 

resources are likely not available for any single major graphite 

district in the world. Table J4 tabulates conservative estimates 

for recoverable graphite resources for a number of major 

graphite-producing nations. Reserves of natural graphite (that 

is, the identified, delineated, and reported world resources that 

are considered to warrant economic exploitation at the time 

of reserve determination) are equivalent to 81 million metric 

tons. Approximately 26 percent of reserves are crystalline 

(flake and vein type [lump or chip]) graphite materials for 

which demand is increasing. Other identified resources beyond 

reserves (“Other identified resources” in table J4) are about 

an additional 1.4 billion metric tons, leading to an estimate of 

total identified graphite resources of approximately 1.5 billion 

metric tons worldwide. Approximately one-half of this total 

resource estimate is flake graphite. Future exploration is 

likely to result in the discovery of world resources that are 

many times this estimate; however, many new discoveries are 

likely to be located in remote areas with high mining costs 

and limited access to infrastructure and industrial markets that 

use graphite. The Lac Knife deposit in Quebec, Canada, is an 

example of a relatively recent discovery in a remote setting 

(Bonneau and Raby, 1990).

Future discoveries of amorphous and flake graphite 

deposits in the conterminous United States are likely to be 

limited by favorable geology and preexisting exploration 

coverage. The coal fields of the United States are well 

explored and delineated, and past igneous activity in the coalfields is either absent or of such limited scale that the chance 

of finding amorphous graphite deposits of commercial size and 

grade is negligible. Furthermore, with the exception of Alaska, 

crystalline metamorphic rocks that could potentially host flake 

and vein (lump or chip) graphite deposits have limited extent 

and exposure in the United States, and most of the crystalline 

basement that is present is well characterized. The identified 

flake graphite deposits in the Kigluaik Mountains of Alaska 

indicate potential for additional deposits in surrounding areas; 

however, these are likely to occur in remote areas with rugged 

terrain and limited infrastructure, and such deposits would be 

a great distance from graphite markets and have high mining 

costs. The weathered low-grade crystalline flake deposit 

districts in Alabama and Texas may provide future domestic 

sources of graphite if flake graphite prices increase, because of 

the ease of mining the deposits and their proximity to markets.

China 639 kt

India 134 kt 

Brazil 78 kt

North Korea 30 kt

Canada 27 kt 

European Union 21 kt 

All others 

(not including

Russia) 27 kt

Figure J4. Pie chart showing average annual natural 

graphite production for the period 2006–10, by country 

or region and amount (in thousand metric tons [kt]). 

The tonnages are estimated based on data reported in 

Olson (2011).
J18 Critical Mineral Resources of the United States—Graphite

Table J4. Estimates of world graphite resources, by country, commodity type, and resource category, in 

thousand metric tons of recoverable graphite.

[Resources reported for recoverable amorphous, flake, or crystalline (vein-type) commodity categories. Resource categories: 

(a) Reserves are demonstrated resources of recoverable natural graphite considered to warrant economic exploitation at the time 

of reserve determination; (b) Other identified resources are estimates of characterized recoverable resources beyond those reported 

as reserves; only part of this estimate would be economic; and, (c) Total identified resources includes reserves and other identified 

resources. Data are adapted from Olson (2013) and Taylor (2006) and rounded to two significant figures. NA, not available]

Country Type Reserves

Other

identified

resources

Total

identified

resources1

Austria Amorphous 50 1,100 1,200

Brazil Flake 360 3,400 3,800

Canada Flake 1,500 4,200 5,700

China Amorphous 55,000 NA NA

Flake 6,000 350,000 360,000

Czech Republic Flake 900 12,000 13,000

Germany Flake 130 690 820

India Flake 735 13,000 14,000

Madagascar Flake 940 180,000 180,000

Mexico Amorphous 3,100 10,000 13,000

Flake 106 320 430

North Korea Amorphous 1,000 30,000 31,000

Flake 700 1,400 2,100

Norway Flake 200 260 460

Republic of Korea Amorphous 20 2,100 2,100

Flake 6 74 80

Romania Amorphous 300 NA 300

Russia, with Ukraine Amorphous 1,000 560,000 560,000

Flake 6,400 94,000 100,000

Sri Lanka Crystalline, vein 50 400 450

Flake 1,800 7,000 8,800

United States2 Amorphous 0 5,900 5,900

Flake 0 280 280

Zimbabwe Flake 600 1,200 1,800

Other Flake 280 920 1,200

Total Amorphous 60,000 750,000 810,000

Total Flake 21,000 670,000 690,000

Total Crystalline, vein 50 400 450

Total By resource category 81,000 1,400,000 1,500,000

1

 Total identified resources includes reserves and other identified resources.

2Other identified resources includes Alabama, Alaska, and New York. 
Environmental Considerations J19

Exploration for New Deposits

Prospecting for graphite deposits consists primarily 

of outcrop examination, trenching, and sampling, usually 

followed by drilling. Because of its chemical inertness, 

graphite resists weathering, and outcrops containing graphite 

and residual soils containing conspicuous graphite flakes may 

be evident. Knowledge of areal geology and the geologic 

features and settings of different graphite deposit types and 

associated deposits, such as coal, can be used to identify 

promising exploration targets. The Lac Knife flake graphite 

deposit was initially discovered during routine geologic 

mapping. When interest developed, the prospect was trenched 

by shovel, followed by further surface excavation, and later 

systematically characterized at depth by more than 99 drill 

holes. All this activity occurred during a more than 30-year 

time period (Bonneau and Raby, 1990).

Because of the high electrical conductivity of graphite 

(in deposits where individual graphite flakes are touching), 

a variety of electromagnetic methods have been used to 

search for deposits, principally flake graphite deposits. 

Electro-magnetic geophysical methods measure variations in 

the electrical conductivity of the ground caused by changes 

in mineral content, properties, or subsurface features. The 

methods rely on the process of electromagnetic induction 

that describes how a conductive material, such as graphite, 

will interact with a magnetic field. Surveys are conducted by 

using either surface or airborne methods; all methods use a 

transmitter and receiver coil spaced in standard configurations. 

The different methods use different operating frequencies to 

provide a range of depth penetrations and resolution needed 

for different applications. High-frequency electromagnetic 

systems are best for locating small shallow targets; lower 

frequency systems are used to investigate deeper subsurface 

conditions or define regional targets. The relative merits of 

various geophysical techniques in detecting graphite deposits 

are discussed by Goosens (1982, p. 136).

Graphite flakes may be found in stream sediment samples 

collected during exploration geochemical surveys. The presence of sulfides and vanadium-bearing garnet (goldmanite) 

may also be an indicator. There is a positive correlation 

between carbon, uranium, and vanadium contents in some 

deposits (Li and others, 1985), and positive vanadium, nickel, 

and (or) uranium anomalies in soil associated with graphite 

beds were reported by Tichy and Turnovec (1978) and may 

be considered a geochemical indicator of graphite deposits. 

These geochemical features probably relate to the depositional 

environment and trace element content of the protolith and 

likely vary among deposits. Therefore, it may be difficult to 

use these features effectively in grassroots exploration for 

graphite deposits.

Environmental Considerations

Because of the chemically inert, nontoxic nature of 

graphite, the main environmental concerns associated with 

graphite mining are (a) the presence of other minerals that can 

occur with graphite, and (b) inhalation of graphite particles 

or fine-grained silica minerals in graphite during mining and 

processing. For example, the iron sulfide minerals pyrite and 

pyrrhotite are present in amounts ranging from a trace to 

several percent in some graphite deposits. These minerals can 

cause acid-rock drainage if exposed to air and water in waste 

rock or tailings. Graphite in soils and stream or river sediments is inert and poses no known terrestrial or aquatic risks.

Mine Waste Characteristics

The volumes and mineralogical characteristics of mine 

wastes depend on the size and type of deposit, as well as 

the mining and processing practices employed. For open pit 

mines, the amount of overburden that must be stripped prior 

to the onset of mining contributes to the mine waste. Ore 

processing for disseminated flake graphite typically includes 

milling, flotation, screening, and drying. The resulting mine 

waste and flotation tailings are composed of the minerals 

that make up the host rocks, typically schists and gneisses 

composed mainly of quartz and feldspars. Other silicate 

minerals in these rocks include pyroxenes, amphiboles, 

micas, garnet, and sillimanite. These are all moderate- or 

slow-weathering minerals (Jambor, 2003). Quartz persists 

in the environment, whereas feldspars weather to clays. 

Fast-weathering, potentially acid-producing sulfide minerals, 

such as pyrite and pyrrhotite, are reported at some deposits. 

The yellow iron sulfate mineral jarosite, which is an indicator 

of acidic conditions, is a common weathering product of 

pyritic, gneiss-hosted graphite deposits (Simandl and Kenan, 

1997b). Calcite and other carbonate minerals may be present 

in gangue and can provide acid neutralization if present in 

sufficient amounts.

The largest known flake graphite deposit, the 

Zavalyevskiy deposit in Ukraine, has total resources on the 

order of 100 million metric tons with reserves of 6.4 million 

metric tons containing 5 to 7 percent graphite (Zavalyevskiy 

Graphite Ltd., 2013). The ore is kaolinized gneiss that contains 

garnet, biotite, chlorite, pyroxene, quartz, and sillimanite. 

Both garnet and crushed stone products are recovered 

from the ore. The graphite ore is processed by flotation to 

produce a high-purity graphite concentrate (86 to 97 percent 

graphite) and ash (10 to 13 percent graphite). Graphite 

orebodies in graphite schist at the Liu Mao Mine in China 

are elevated in vanadium (0.2 percent vanadium pentoxide 

[V2

O5 ]), titanium (0.5 percent titanium dioxide [TiO2]), and 

uranium (0.004 percent U). These elemental concentrations 
J20 Critical Mineral Resources of the United States—Graphite

are attributed to garnet intergrown with graphite in the deposit 

(Wilde and others, 1999). In Canada, a composite grab sample 

for metallurgical testing at the Trout Lake (also known as 

Treelined Lake) graphite prospect in southern Ontario contains 

1.8 percent sulfur, which is attributed to 4.5 percent pyrrhotite 

in the sample; the preliminary test indicated that pyrrhotite 

removal by screening and flotation would be required to obtain 

a salable graphite product (Kuehnbaum and Zebruck, 2002). 

At the Bissett Creek flake graphite deposit in Ontario, both 

ore and waste rock are recognized as potential acid generators 

(Systèmes Gèostat International Inc., 2007). The graphite 

occurs in a calcareous biotite-amphibole-quartzofeldspathic 

gneiss with disseminated graphite, pyrite, pyrrhotite, and trace 

chalcopyrite. Graphite concentrations range from about 1 to 

10 percent; the sulfide minerals constitute 1 to 5 percent of 

the rock. Mine plans call for underwater tailings storage and 

the addition of carbonate-bearing material to the waste dump 

to neutralize acid potential. In contrast, at the Eagle deposit in 

British Columbia, high-purity flake graphite is quarried from 

quartz-rich sands, and slightly alkaline sand and aggregate 

byproducts are marketed as construction materials.

Amorphous graphite deposits range from about 

300 to 800 million metric tons of ore and have higher average 

graphite grades than other types of graphite deposits (table J1). 

The ore is essentially the entire graphitized unit. The associated waste primarily is any overburden that must be removed 

to access the deposit. Mineral impurities in amorphous 

graphite deposits reflect the compositions of the protolith coal 

or sediment (Simandl and Kenan, 1997a). In the Bohemia 

region of the Czech Republic, elevated concentrations of 

vanadium and nickel anomalies are associated with amorphous 

graphite deposits (Tichy and Turnovec, 1978). Waste material 

described as graphite gangue at the Jixi deposit in China 

contains zinc, nickel, and mercury (Liu and Man, 2007).

Vein graphite from Sri Lanka, which is significant for its 

purity and high degree of crystallinity, is mined underground 

at the Bogala Mine (370 m deep) and Kahatagaha-Kolongaha 

Mine (650 m deep). Total inferred reserves for these two 

mines are about 400,000 metric tons. At these mines, graphite 

veins are blasted with dynamite. When the fumes clear, 

the ore is hauled to the surface and then hand cobbed and 

graded; no further processing is required, so no tailings or 

large waste piles are produced (Ranasinha and Uragoda, 

1972). Hydrothermal minerals associated with the graphite 

from Sri Lanka include apatite, biotite, calcite, chalcopyrite, 

pyrite, and quartz. The average chemical contents reported 

for different vein graphite morphologies from the Bogala 

and the Kahatagaha-Kolongaha Mines reported by Touzain 

and others (2010) are as follows: 93 to 99 percent carbon, 

0.06 to 1.06 percent iron oxide (Fe2O3), 0.05 to 0.45 percent 

silicon dioxide (SiO2 ), 0.01 to 0.19 percent sulfur, and 

0.004 to 0.82 percent calcium oxide (CaO).

Human Health Concerns

The primary human health concern associated with 

graphite mining is inhalation of dust and fumes generated 

during mining and processing. Graphite is not considered to 

be toxic and is not a listed carcinogen, although crystalline 

silica (a common impurity in graphite) is considered to be 

carcinogenic. Graphite is considered a nuisance dust. The 

time-weighted average (TWA) recommended exposure level 

set by the U.S. National Institute for Occupational Safety and 

Health (NIOSH) is 2.5 milligrams per cubic meter.

Chronic exposure to graphite dust is associated with 

pneumoconiosis, a group of lung diseases associated with 

chronic exposure to mineral and metallic dusts. Graphite 

pneumoconiosis was recognized in the 1940s in graphite 

workers in Sri Lanka based on similarities to pneumoconiosis 

in coal miners (Uragoda, 1997). The fact that graphite always 

contains some quartz (typically about 2 percent but up to 

10 percent) raised concerns that graphite pneumoconiosis was 

actually a slow developing form of silicosis (Ranasinha and 

Uragoda, 1972). A review of 605 cases of graphite pneumoconiosis reported in the literature and experimental studies on 

animals concluded that although many years of occupational 

exposure to pure graphite may cause disease, most studies 

indicate that mixed-dust inhalation is a more likely cause of 

lung disease (Hanoa, 1983). Implementation of dust abatement 

protocols in mining and processing has resulted in decreased 

incidence of lung disease in graphite workers since the 1970s.

Ecological Health Concerns

The environmental fate and effects of dispersed graphite 

flakes were addressed in a study of the use of graphite flakes 

in a dust cloud to obscure visual and electromagnetic observation of military operations under battlefield conditions (Driver 

and others, 1993). Graphite mixed with fog oil and dispersed 

as an aerosol provides a chemically inert obscurant used to 

protect movements of troops and equipment. Dispersion of 

airborne plumes of these mixtures can deposit graphite on 

soil, vegetation, and water surfaces or pose inhalation risks 

to wildlife. Although obscurant releases are short-term events 

(~30 minutes), airborne distribution and surface deposition 

of flake graphite can occur many kilometers downwind of the 

release site. Deposited graphite persists in the environment, 

so the repeated use on military training sites warranted an 

evaluation of potential ecological impacts. These studies are 

not directly applicable to graphite mining and processing, but 

they do demonstrate that graphite flakes likely pose little or no 

chemical risk to the environment. The series of experiments 

described by Driver and others (1993) showed that (a) graphite 

flakes are not toxic to soil invertebrates (oligochaeta, or earthworms), (b) no adverse effects on terrestrial plants (corn and 

cucumber) were noted in soils amended with graphite flakes in 
Problems and Future Research J21

concentrations up to 0.5 percent, by weight, and (c) a 48-hour 

acute aquatic toxicity test toward daphnia (water fleas) using 

graphite suspensions was toxic; however, iron contaminants 

in the graphite may explain the toxicity. The long-term effects 

of obscurant flake graphite on exposed avian and mammalian 

wildlife are unknown.

Carbon Footprint

The term “carbon footprint” refers to the amount of 

carbon dioxide (CO2 ) emitted in a 1-year period. According 

to GrafTech International Holdings, Inc. (2013), natural 

graphite has a substantially lower carbon footprint for both 

CO2 emissions and energy consumption during mining and 

processing than other mineral commodities for which data are 

available, such as aluminum or copper, on an equivalent mass, 

volume, or yearly production basis. Cost- and energy-efficient 

technologies that could reduce CO2 to synthetic graphite could 

lower CO2 emissions and produce inert graphite, providing 

an industrial option to sequester carbon from the atmosphere 

(Xu and others, 2005).

Regulatory and Environmental Considerations

The environmental requirements for natural graphite, 

which is an inert, nontoxic substance, are limited to dust 

control and control of organic vapors that may arise during 

mining or processing of graphite ores or the fabrication of 

graphite products. Given the current and likely future absence 

of graphite mining in the United States, no other domestic 

regulatory or environmental requirements are anticipated.

Problems and Future Research

Major shifts in U.S. consumption of natural graphite by 

end use are underway and include a long-term decrease in the 

use of amorphous graphite for refractory and other applications and an increase in the use of flake graphite, particularly 

for emerging energy and high-technology applications (Taylor, 

2006). In addition, shifts in the global graphite supply and 

exports are anticipated as China’s rapidly growing domestic 

market for graphite, coupled with their developing mine 

production problems and rising labor costs, may serve to 

limit Chinese exports in the future, particularly of crystalline 

flake graphite; meanwhile, anticipated large-scale fuel cell 

and battery applications could dramatically increase world 

demand for graphite (Taylor, 2006; Olson, 2013; Olson and 

others, 2016). For the near future, however, China’s graphite 

production is expected to increase and its dominance of world 

production to continue (Olson, 2013). Possible disruptions to 

global supplies are likely to be temporary or sporadic because 

identified reserves in China are sufficient to support new 

mines and additional production. In addition, exploration for 

new flake graphite deposits is ongoing, and additional deposits 

are near development in Canada and elsewhere (Olson, 2013; 

Olson and others, 2016).

Other than scientific studies on some vein graphite 

deposits of scientific (not commercial) interest, little recent 

geologic characterization and study has been undertaken on 

graphite deposits in the United States, and exploration for new 

deposits has not taken place. The lack of exploration is chiefly 

because known deposits are considered noncommercial, 

large areas of the country lack the geologic conditions and 

settings thought necessary to form economic graphite deposits, 

graphite supplies have been reliably available from foreign 

sources for many years, and synthetic graphite is increasingly 

used for many applications. As earlier noted by Weis (1973), 

research and development of better techniques to beneficiate 

low-grade flake graphite ores would be needed to support a 

viable domestic industry.

Studies on the temporal variation of carbon deposition 

and preservation and the behavior of organic matter and 

graphite under metamorphic and hydrothermal conditions 

might provide additional insights into the occurrence and 

distribution of high-grade flake graphite deposits. Many of 

the important flake graphite deposits worldwide occur in 

Precambrian crystalline metamorphic rocks, and most of these 

occur in Pan-African age supracrustal metamorphic belts. 

In addition, many of these deposits have grades exceeding 

15 percent contained carbon, which exceeds the typical carbon 

content range of organic-bearing shales (Vine and Tourtelot, 

1970). It is not known whether these deposits represent 

unusual protoliths—for example, saprolitic alginate (oilshale), unusual periods or regions of biological productivity 

and preservation of organic matter, or if they record some 

carbon enrichment owing to mechanical deformation or 

precipitation of graphite by processes involving internal or 

external buffering or mixing of metamorphic fluids.
J22 Critical Mineral Resources of the United States—Graphite

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Lithium

By Dwight C. Bradley, Lisa L. Stillings, Brian W. Jaskula, LeeAnn Munk, and 

Andrew D. McCauley

Chapter K of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–K

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter K

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................K1

Introduction..................................................................................................................................................K1

Historical Overview ...........................................................................................................................K1

Lithium as a Critical Element ............................................................................................................K2

Uses and Sources ..............................................................................................................................K2

Geology.........................................................................................................................................................K5

Geochemistry and Mineralogy ........................................................................................................K5

Deposit Types .....................................................................................................................................K5

Lithium-Cesium-Tantalum Pegmatite Deposits ....................................................................K8

Lithium-Enriched Granites .......................................................................................................K8

Lithium Brine Deposits in Closed Basins ..............................................................................K8

Lithium in Other Brines ..........................................................................................................K12

Lithium-Clay Deposits ............................................................................................................K12

Lithium-Zeolite Deposits ........................................................................................................K12

Resources and Production ......................................................................................................................K13

Production.........................................................................................................................................K13

Reserves and Other Identified Resources ...................................................................................K13

Undiscovered Resources ...............................................................................................................K13

Exploration for New Deposits .................................................................................................................K14

Exploration for Lithium-Cesium-Tantalum Pegmatite Deposits ................................................K14

Exploration for Lithium-Brine Deposits in Closed Basins .........................................................K14

Exploration for Lithium in Sedimentary Rocks (Clays and Zeolites) ........................................K14

Environmental Considerations ................................................................................................................K15

Sources and Fate in the Environment ..........................................................................................K15

Mine Waste Characteristics ..........................................................................................................K16

Human Health Concerns .................................................................................................................K16

Ecological Health Concerns ...........................................................................................................K16

Carbon Footprint...............................................................................................................................K16

Problems and Future Research ..............................................................................................................K17

Acknowledgments....................................................................................................................................K17

References Cited.......................................................................................................................................K17

Suggested citation:

Bradley, D.C., Stillings, L.L., Jaskula, B.W., Munk, LeeAnn, and McCauley, A.D., 2017, Lithium, chap. K of Schulz, K.J., 

DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic 

and environmental geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. K1– K21, 

http://dx.doi.org/10.3133/pp1802K.
Lithium

By Dwight C. Bradley, Lisa L. Stillings, Brian W. Jaskula, LeeAnn Munk, and Andrew D. McCauley

Abstract

Lithium, the lightest of all metals, is used in air treatment, 

batteries, ceramics, glass, metallurgy, pharmaceuticals, and 

polymers. Rechargeable lithium-ion batteries are particularly 

important in efforts to reduce global warming because they 

make it possible to power cars and trucks from renewable 

sources of energy (for example, hydroelectric, solar, or 

wind) instead of by burning fossil fuels. Today, lithium is 

extracted from brines that are pumped from beneath arid 

sedimentary basins and extracted from granitic pegmatite 

ores. The leading producer of lithium from brine is Chile, and 

the leading producer of lithium from pegmatites is Australia. 

Other potential sources of lithium include clays, geothermal 

brines, oilfield brines, and zeolites. Worldwide resources of 

lithium are estimated to be more than 39 million metric tons, 

which is enough to meet projected demand to the year 2100. 

The United States is not a major producer at present but has 

significant lithium resources.

Introduction

Historical Overview

Lithium (from the Greek lithos, for stone) is the first 

metal in the periodic table. It was recognized as an element in 

1817 when the Swedish chemist Johan Arfvedson analyzed 

the mineral petalite. Useful amounts of lithium metal were first 

isolated in 1855 by August Matthiessen and, independently, 

by Robert Bunsen, after whom the Bunsen burner was named 

(Emsley, 2011). By this time, lithium had been found in other 

silicate minerals (lepidolite and spodumene) and had been 

detected in trace quantities in plants, seawater, spa waters, and 

the human body (Emsley, 2011). In 1869, Dmitiri Mendeleev 

correctly positioned it adjacent to the alkali metal sodium in 

his then-revolutionary periodic table of elements.

Petalite, lepidolite, and spodumene are found in granitic 

pegmatites. The first significant lithium mining began in 

1898 at the Etta pegmatite in South Dakota. In the early 

1900s, other pegmatite mines in California, South Dakota, 

and New Mexico also began lithium production, and the 

United States dominated the global supply (Norton, 1973). The 

process of extracting lithium from brine was also first developed 

in the United States. The earliest production of lithium from 

brine took place in 1938 as a byproduct of sodium carbonate 

production at Searles Lake, California. A larger lithium-brine 

deposit at Clayton Valley, Nevada, began production in 1966. 

Today, brine resources dominate the lithium industry.

In the 1970s, when the U.S. Geological Survey (USGS) 

published Professional Paper 820, “United States Mineral 

Resources” (Brobst and Pratt, 1973), lithium was a relatively 

obscure mineral commodity that did not even warrant an 

entire chapter of its own (Norton, 1973). Since then, the global 

resource picture for lithium has changed considerably. In 

1973, the United States was still the world’s leading lithium 

producer; today, U.S. production lags far behind that of Chile, 

Australia, and several other nations. Additionally, two new 

classes of ore deposits—lithium clays and lithium zeolites—

have been recognized. Between 1975 and 2005, world lithium 

production increased by a factor of about five (fig. K1), and 

more growth is anticipated.

Lithium Figure K1

Pegmatite

1900 1925

Add to caption: Modified from Mohr 

and others (2010)

0

10

20

30

5

15

25

1950 1975

Year

Deposit type

not specified

Lithium production, in thousand metric tons

Brine

2000

Figure K1. Graph showing world lithium production from 

1900 to 2007, by deposit type and year. The layers of the graph 

are placed one above the other, forming a cumulative total. 

Modified from Mohr and others (2010).
K2 Critical Mineral Resources of the United States—Lithium

Many advances in understanding lithium geology 

were spurred by the 1973 oil embargo. Lithium was—and 

still is—seen as a strategic mineral commodity that could 

contribute toward U.S. energy independence. At that time, 

the USGS launched a major research effort on lithium 

resources that lasted for several years and involved dozens 

of researchers. Initial results were presented in a multichapter 

USGS Professional Paper (Vine, 1976), and final results 

were summarized by Vine (1980). After about 1980, national 

concerns about energy self-sufficiency ebbed, and studies 

in lithium-resource geology fell into neglect for more than a 

quarter century.

A recent revival of interest in lithium has been driven by 

the threat of global warming from the burning of fossil fuels, 

combined with the recognition that a transition to electric and 

hybrid vehicles, which are powered by lithium batteries, could 

substantially reduce the Nation’s carbon footprint and its reliance on foreign oil (U.S. Department of Energy, 2010). Recent 

summaries of the global lithium resource picture include those 

by Evans (2008, 2012, 2014), Yaksic and Tilton (2009), Mohr 

and others (2010), Gruber and others (2011), and Kesler and 

others (2012).

Lithium as a Critical Element

Lithium has been listed as one of the critical or nearcritical elements in various recent studies based largely on its 

importance in green technologies. It was one of the 15 mineral 

commodities studied by participating countries in the 

International Strategic Minerals Inventory (ISMI) of the 1980s 

and 1990s (Anstett and others, 1990). The National Research 

Council (2008) concluded that lithium, although not critical 

at the time of the 2008 assessment, could potentially become 

critical depending on the level of growth of new uses, such as 

hybrid vehicle batteries. In the U.S. Department of Energy’s 

Critical Materials Strategy (2010), lithium is included as 1 

of 16 key elements. The 2011 joint report of the American 

Physical Society Panel on Public Affairs and the Materials 

Research Society on energy-critical minerals included lithium 

as 1 of 29 key elements for green technology.

Uses and Sources

Lithium has many uses. Worldwide lithium consumption 

in 2013 by end use was estimated to be as follows (fig. K2): 

ceramics and glass, 35 percent; rechargeable batteries, 

29 percent; lubricating greases, 8 percent; continuous casting 

mold flux powders, 6 percent; air treatment, 5 percent; and 

polymer production, 5 percent; primary batteries, 2 percent; 

primary aluminum production, 1 percent; and other uses, 

9 percent (Merriman, 2014).

Ceramics and glass are expected to continue to be the 

leading end uses for lithium for years to come. In glassmaking 

(Garrett, 2004), lithium makes the process more economical 

by reducing the melting point and decreasing viscosity. 

Ceramics

and glass 

35%

Rechargeable

batteries 

29%

Other

uses

9%

Lithium Figure K2

Continuous casting

mold flux powders 

6%

Air treatment 

5%

Polymer production 

5%

Primary aluminum production 

1%

Lubricating

greases 

8%

Primary batteries

2%

Figure K2. Pie chart showing major end uses of lithium as a 

percentage of world consumption in 2013. Data are courtesy 

of David Merriman, Roskill Information Services, Ltd., written 

commun., May 2014.

The addition of 0.1 to 4 percent lithium oxide (Li2O), to 

various glasses increases hardness and reduces thermal 

expansion. A common application is in glass stovetops; where 

ordinary window-glass would shatter, lithium-bearing glass 

can withstand large temperature swings and rough use. In the 

ceramics industry, lithium is used to make porcelain enamels, 

glazes, and tiles; as in glassmaking, a small amount of lithium 

lowers the melting temperature and reduces thermal expansion 

in the resulting ceramic product (Garrett, 2004).

Production of batteries, which currently rank second 

among the end uses of lithium, will likely see increased 

production in the next few years as the production of electric 

and hybrid vehicles increases. Production will result in 

increased capacity utilization of the vehicle battery plants that 

are currently being built worldwide. Battery manufacturing has 

the largest growth potential of any sector of the lithium industry; 

several major automobile companies are developing lithium-ion 

batteries (fig. K3H). Demand for rechargeable lithium batteries 

now exceeds that of rechargeable nonlithium batteries for use 

in cellular telephones, cordless tools, MP3 players, and portable 

computers and tablets. Nonrechargeable lithium batteries are 

used in calculators, cameras, and watches (Jaskula, 2013).

Other uses of lithium are not widely appreciated by 

consumers. Lithium is a minor but key ingredient in the 

electrolytic refining of aluminum. Lightweight lithium-aluminum 

alloys have applications in the aerospace industry. For example, 

the fuel tank of the NASA space shuttles used an aluminumlithium alloy (Garrett, 2004). Lithium, as lithium stearate or 

similar compounds, is widely used as a high-performance lubricant 

in automotive, aircraft, industrial, marine, and military applications 

(Garrett, 2004). As of 1993, 60 percent of all industrial lubricants 

contained lithium, typically at concentrations of 1 to 2 percent. 

Lithium chloride and bromide are among the best available 
Introduction K3

substances for absorbing water; thus, these salts are used in air 

conditioning systems to reduce humidity while simultaneously 

cooling the air. Since the 1940s, pharmaceuticals containing 

lithium compounds have been used to control bipolar disorder. 

Finally, the lighter of two lithium isotopes, 6Li, was used in the 

production of tritium (3H) after World War II by the U.S. Atomic 

Energy Commission and later by the U.S. Department of Energy 

in the production of nuclear weapons (Garrett, 2004).

Lithium is traded in three forms: mineral concentrates, 

mineral compounds, and refined metal. Lithium minerals—

mainly spodumene, petalite, and lepidolite—are mined from 

pegmatites and are used mostly as feedstock for glasses and 

ceramics. Most lithium compounds (for example, lithium 

carbonate, lithium chloride, and lithium hydroxide) are 

obtained from brines. Lithium metal (fig. K3G) is obtained by 

electrolysis from lithium chloride.

The prices of lithium metal and lithium compounds are 

not published; however, approximate values can be gleaned 

from a variety of industry sources. Battery-grade lithium 

carbonate is currently valued at between $6,500 and $7,000 

per metric ton. In 2010, Japan imported lithium metal valued 

at $33.36 per pound, or $73,546 per metric ton.

Lithium is the element of choice in the various 

applications listed above. Substitutes do exist, however, in 

batteries, ceramics, greases, and manufactured glass. For 

example, calcium and aluminum soaps can substitute for 

stearates in greases; calcium, magnesium, mercury, and 

zinc can be used as anode materials in batteries; sodic and 

potassic fluxes can substitute for lithium in ceramics and 

glass manufacturing; and composite materials of boron, glass, 

or polymer fibers can be used in place of aluminum-lithium 

alloys (Jaskula, 2013).

Figure K3. Photographs showing some sources and uses of lithium. A, Molds of giant spodumene crystals in the Etta 

lithium-cesium-tantalum (LCT) pegmatite, Black Hills, South Dakota; note miner for scale. From Schaller (1916). B, Graphic 

granite, which is an intergrowth of quartz and feldspar that is characteristic of pegmatites. This granite is part of the Havey

pegmatite in Maine. By Dwight C. Bradley. C, Elbaite, also known as watermelon tourmaline, which is a lithium-bearing 

gemstone. In 1973, a single gem-filled pocket in a pegmatite in Maine produced about 2 metric tons of tourmaline that would 

be valued today (for gems and mineral specimens) at about $40 million. This sample is from the Dunton Mine in Newry, 

Maine. Courtesy of the Mineralogical & Geological Museum at Harvard University (speciman number MGMH 125525. Dunton 

Mine, Newry, Maine. Copyright 2012. President and Fellows of Harvard College; all rights reserved). D, Salar de Atacama 

in Chile, where lithium brines are pumped from the shallow subsurface. By Luca Galuzzi (www.galuzzi.it)/CC-BY-SA-2.5

(http://commons.wikimedia.org/wiki/File:Miscanti_Lagoon_San_Pedro_de_Atacama_Chile_Luca_Galuzzi_2006.jpg). 

A B

C D
K4 Critical Mineral Resources of the United States—Lithium

Figure K3.—Continued E, Lithium-brine evaporating ponds at Clayton Valley, Nevada. The dark rectangular pond at 

the upper center of the image is 1.7 kilometers long. By Doc Searls/CC-BY-2.0 (http://commons.wikimedia.org/wiki/

File:Chemetall_Foote_Lithium_Operation.jpg). F, Drill core through a sequence of Pliocene-age lake sediments containing 

lithium clay (hectorite) from Fish Lake Valley, Nev. (for location, see fig. K5, inset). Markers indicate depth in feet below 

the ground surface. Hectorite is present in the tan and gray layers of siltstone and claystone. By Lisa L. Stillings. G, Lithium 

metal is so light that it floats in oil, and so reactive that it cannot be stored in air or water. By W. Oelen/CC-BY-SA-3.0 

(http://commons.wikimedia.org/wiki/File%3ALithium_element.jpg). H, View into a Tesla Model S® electric car showing the 

lithium-ion battery covering much of the chassis. Various battery configurations in electric vehicles use between about 

10 and 22 kilograms of lithium. By Oleg Alexandrov/CC-BY-SA-3.0 (http://commons.wikimedia.org/wiki/File:Tesla_Motors_

Model_S_base.JPG).

E

F

G H
Geology K5

Geology

Geochemistry and Mineralogy

In most geologic settings, lithium is found only as a trace 

element, measured in parts per million (ppm). Earth’s mantle 

contains 1.6 ppm lithium and the oceanic crust contains 

4.3 ppm (Sun and McDonough, 1989). The upper continental 

crust (that is, rocks typically at Earth’s surface on land) 

contains an average of 20 ppm lithium (McLennan, 2001); at 

this abundance, lithium ranks about 30th among the elements, 

behind copper but ahead of lead, tin, and silver.

Lithium is classed as a large-ion lithophile element: as a 

cooling magma starts to crystallize minerals, lithium stays in 

the remaining melt until near the end. Active plate tectonics 

over Earth’s history have concentrated lithium in the continental crust through the partial melting of the mantle beneath 

mid-ocean ridges and volcanic arcs. The melt, or magma, rises 

and then cools to become new rock in Earth’s crust, bringing 

much of the available lithium with it. Among the common 

rock or sediment types, the highest lithium concentrations 

are in shales (which average 66 ppm), deep-sea clays (which 

average 57 ppm), and low-calcium granites (which average 

40 ppm) (Faure, 1998 and references therein). These trace 

concentrations are insufficient for an ore deposit or even 

for the formation of minerals in which lithium is part of the 

chemical formula. When present only in trace concentrations 

such as these, lithium atoms substitute for other metals 

(typically magnesium) in common rock-forming minerals.

Lithium minerals form only when rare combinations 

of favorable factors line up. Most of the known lithium 

minerals are found in coarsely crystalline granites known 

as lithium-cesium-tantalum (LCT) pegmatites. In terms of 

lithium resources, the most important minerals are spodumene and petalite (both lithium aluminum silicates) and the 

pink mica lepidolite (potassium lithium aluminum silicate). 

The main lithium mineral in sedimentary rocks is the clay 

hectorite. More information on lithium minerals is provided 

in table K1.

Lithium is extremely soluble. During weathering of 

rocks, it tends to be removed in solution and carried to the 

sea by rivers. Thus, lithium would be expected to have 

built up in the oceans in the same way that a buildup of 

sodium has made the oceans salty. Yet, remarkably, seawater 

contains less than 1 ppm lithium. The likely explanation 

is that seawater lithium is scavenged in trace quantities by 

clay minerals and accumulates in sea-floor oozes.

Deposit Types

Present and potential sources of lithium worldwide are 

broken down by deposit type, as follows: closed-basin brines, 

58 percent; pegmatites (including lithium-enriched granites), 

26 percent; lithium-clays (hectorite), 7 percent; and oilfield 

brines, geothermal brines, and lithium-zeolites (jadarite), 

3 percent each (Evans, 2012). Figures K4 and K5 show the 

worldwide distributions of lithium deposits.

Table K1. Commercially and (or) scientifically important lithium-bearing minerals.

[Chemical compositions are from Barthelmy (2014)]

Mineral name Chemical formula

Lithium

(weight

percent)

Geologic setting

Amblygonite (Li,Na)Al(PO4

)(F,OH) 3.44 Pegmatite intrusions in orogenic belts; solid solution with 

montebrasite

Elbaite Na(Li,Al)3

Al6(BO3)3Si6O18(OH)4 1.89 Tourmaline group. Pegmatite intrusions in orogenic belts. 

Includes the gem form watermelon tourmaline

Eucryptite LiAlSiO4 5.51 Pegmatite intrusions in orogenic belts

Hectorite Na0.3(Mg,Li)3

Si4O10(OH)2 0.54 Hydrothermal alteration of volcanic ash in arid, closed basins

Jadarite LiNaB3

SiO7(OH) 3.38 Hydrothermal alteration of volcanic ash in arid, closed basins

Lepidolite K(Li,Al)3

(Si,Al)4O10(F,OH)2 3.58 Pegmatite intrusions in orogenic belts

Montrebasite LiAl(PO4

)(OH,F) 4.74 Pegmatite intrusions in orogenic belts

Petalite LiAlSi4

O10 2.09 Pegmatite intrusions in orogenic belts

Spodumene LiAlSi2

O6 3.73 Pegmatite intrusions in orogenic belts. Gem forms are triphane 

(yellow), kunzite (pink), and hiddenite (green)

Zabuyelite Li2

CO3 18.79 Evaporite mineral (rare)
K6 Critical Mineral Resources of the United States—Lithium 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10°

20°30°40°50°

80°

70°

70°

80°

60°

Whabouchi Tanco

BlackHills

Big Whopper

Altai

Lipovy Log

Elba Covas de Barrosa

Khnefissat

Taghawlor Jiajika

Kenticha

Manono

Bikita Pilgangoorla Wodgina Rubicon Volta Grande

San Diego Kings Mountain

Bald

Deans Greenbushes

Lithium Figure K4, world map of pegmatites

Alto Ligonha

Sahany Valley

Zavatinskoe

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

LCT pegmatitesLithium-enriched granites

EXPLANATION

ndicate pegmatites World map showing locations of selected lithium-cesium-tantalum (LCT) pegmatites and lithium granites. The larger red squares i Figure K4.

containing major lithium and (or) tantalum resources. Adapted from Bradley and others (2016).
Geology K7

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Salar de Uyuni

Salar del Hombre Muerto

Salar de Atacama

PARAGUAY

ARGENTINA

CHILE

BOLIVIA

Kings Valley

Hector

Fish Lake Valley

Salton Sea

Searles

Lake

UTAH

ARIZONA

NEW

MEXICO

NEVADA

OREGON IDAHO

CALIFORNIAClayton Valley

Smackover

Zhabuye Lake

Jadar Basin

Bigadiç Taijanier Lake

DXC

Lithium brines

Lithium-enriched strata

Closed-basinOilfieldGeothermalLithium claysLithium zeolites

EXPLANATION

Lithium Figure K5, brine clay, and lithium-zeolite World map showing locations of selected closed-basin lithium-brine, lithium-enriched oilfield brine, geothermal brine, lithium- Figure K5.

deposits. The locations of the closed-basin brines are from a tabulation by Munk and others (2016).
K8 Critical Mineral Resources of the United States—Lithium

Lithium-Cesium-Tantalum Pegmatite Deposits

Granitic pegmatites are igneous rocks distinguished by an 

extremely coarse but variable grain size and by other unique 

textures, particularly what are known as graphic intergrowths 

(fig. K3B; London, 2008). Pegmatites are composed primarily 

of the major minerals in granite, namely quartz, albite (sodium 

feldspar), potassium feldspar, and the white mica muscovite, 

plus smaller amounts of garnet, tourmaline, biotite, and 

apatite. In addition to the lithium minerals petalite, lepidolite, 

and spodumene, LCT pegmatites may contain the cesium 

ore mineral, pollucite; the tantalum ore mineral, columbitetantalite; the beryllium ore mineral, beryl; and the tin ore 

mineral, cassiterite. More information on tantalum, beryllium, 

and tin resources in LCT pegmatites can be found in Foley and 

others (this volume, chap. E), Schulz and others (this volume, 

chap. M), and Kamilli and others (this volume, chap. S). LCT 

pegmatites may also yield gemstones and high-value museum 

specimens of rare minerals. Gemstones include the beryl 

varieties aquamarine, emerald, and heliodor; the lithiumbearing tourmaline variety elbaite (“watermelon tourmaline”) 

(fig. K3C ); and the spodumene varieties hiddenite, kunzite, 

and triphane. LCT pegmatites are also mined for feldspars, 

muscovite, and ultrapure quartz. The diversity of mineral 

species in the most complex LCT pegmatites is impressive; 

at the Tanco deposit in Manitoba, Canada, for example, 

105 minerals have been identified (London, 2008). LCT 

pegmatites are famous among geologists and mineral collectors for their huge crystals. The largest spodumene crystal yet 

found is 14 meters (m) long (from South Dakota) (fig. K3A); 

the largest beryl crystal, 18 m; and the largest potassium 

feldspar crystal, 49 m (Rickwood, 1981). These finds are all 

the more remarkable because pegmatites are believed to have 

crystallized in a matter of days to years (London, 2008).

The main producers of lithium from pegmatites are 

Australia, Brazil, China, Portugal, and Zimbabwe. The world 

has hundreds of known deposits (fig. K4), and some of these 

(for example, in Quebec, Canada) contain large reserves of 

lithium that have not been mined. The United States is not 

currently producing lithium from pegmatites, although large 

reserves remain in the Kings Mountain pegmatite district in 

North Carolina. Other domestic deposits are too small to be 

viable under current or likely market conditions. Kesler and 

others (2012) estimated lithium resources for 20 of the world’s 

most important pegmatite deposits or districts.

Bradley and McCauley (2013) and Bradley and others 

(2017) have summarized tectonic factors in pegmatite genesis. 

LCT pegmatites are found in the metamorphic-igneous 

hinterlands of orogenic belts and are a consequence of plate 

convergence. Most LCT pegmatites formed during collision 

between continents or microcontinents and are associated with 

aluminum-rich granites that originated by melting of 

metamorphosed sedimentary rocks. Dozens of pegmatites 

in the Appalachians formed during the protracted collision 

between Africa and North America that took place 370 million 

to 275 million years ago.

LCT pegmatites can be dated using isotope geochronology. The age is a function of the ratio of the abundances of 

a radioactive isotope and its decay product. In pegmatites, the 

minerals columbite-tantalite, monazite, and zircon are dated 

by exploiting the decay of uranium-238 to lead-206. LCT 

pegmatites from six continents have now been dated (Bradley 

and McCauley, 2013). The oldest one, which is located in 

Australia, formed at about 2,829 mega-annum (Ma, or million 

years before present) and the youngest, in Italy, at about 7 Ma 

(McCauley and Bradley, 2014). The global age distribution is 

episodic, with a few peak times separated by long intervals of 

few or no pegmatites (fig. K6). The age peaks at 2,638; 1,809; 

962; 485; 309; 274; and 97 Ma correspond to times when 

Earth’s continents were being assembled into supercontinents 

(Bradley, 2011).

Even the biggest LCT pegmatites are much smaller 

than typical granitic plutons, which commonly have areas of 

greater than 10 square kilometers (km2 ). The Greenbushes 

pegmatite in Western Australia, which is large for an LCT 

pegmatite, is 3 kilometers (km) long and a few hundred meters 

across (Partington and others, 1995). On a regional scale, 

LCT pegmatites tend to occur in districts alongside fertile 

granites, with the highest concentration of lithium and other 

incompatible elements in the more distal pegmatites (fig. K7). 

Individual pegmatites take various shapes, ranging from 

tabular dikes and sills to irregular bodies.

Lithium-Enriched Granites

Some muscovite-bearing granites include zones that are 

enriched in the elements lithium, tantalum, tin, and fluorine. 

At the Yichun Mine in Jiangxi Province, China, the top of a 

biotite-muscovite granite grades into muscovite granite and 

then into lepidolite granite, which has been mined for lithium 

and tantalum (Schwartz, 1992). Lithium-enriched granites are 

closely related to LCT pegmatites, and the two have not been 

distinguished from one another in recent global assessments of 

lithium resources (for example, Kunasz, 2006; Evans, 2008; 

Yaksic and Tilton, 2009; Naumov and Naumova, 2010; and 

Gruber and others, 2011).

Lithium Brine Deposits in Closed Basins

Closed-basin brine deposits contain an estimated 

58 percent of the world’s identified lithium resources (Evans, 

2012). Lithium brine deposits are accumulations of saline 

groundwater that are enriched in dissolved lithium. The locations of the world’s major lithium brine deposits are shown on 

the map in figure K5. The producing deposits have average 

lithium concentrations that range from 160 to 1,400 ppm and 

estimated resources of 0.3 to 6.3 million metric tons of lithium 

(Munk and others, 2016). The producing deposits are located 

in Asia, North America, and South America, and lie within the 

northern and arid latitudinal belts on either side of the equator. 

These deposits share a number of characteristics, including 
Geology K9

0

0 1,000 2,000 3,000 4,000

0.2

0.4

0.6

0.8

1.0

Lithium, in million metric tons

Lithium Figure K6

0 

2 

4 

6 

8 

10 

2,638 

1,809 

962 

309 

274 

97 

Number of LCT pegmatites Number of detrital zircons

Age, in million years before present

Phanerozoic Proterozoic Archean

A. Age distribution of LCT pegmatites

B. Age distribution of detrital zircons from modern river sands

C. Age distribution of lithium resources in pegmatites

485 

n=21,655

1,200

1,100

1,000

900

800

700

600

500

400

300

200

100

0

Figure K6. A, Histogram of the number of lithiumcesium-tantalum (LCT) pegmatites formed per 

50-million-year interval as a function of geologic time

showing their episodic genesis (McCauley and Bradley,

2013, 2014). B, Histogram showing the age distribution

of detrital zircon in modern sand as a function of time.

Maxima are interpreted as times when supercontinents

were being assembled and minima as times when

they were together. LCT pegmatites are thus linked

to supercontinent assembly. From Voice and others

(2011). C, Histogram showing global lithium resource in

pegmatites per 100-million-year interval as a function

of geologic time. Note the broad correspondence

between all three plots. Modified from McCauley

and Bradley (2014).

the following: (a) an arid climate; (b) closed basin containing 

a salt lake or salt flat; (c) tectonically driven subsidence; (d) 

associated igneous or geothermal activity; (e) lithium-bearing 

source rocks; (f) one or more adequate aquifers, to host the 

brine reservoir; and (g) sufficient time to concentrate a brine. 

Key aspects of the proposed lithium-brine deposit model are 

shown in figure K8.

The lithium brine deposit in Clayton Valley, Nev., is a 

well-studied example that can be used to highlight some of 

the key characteristics of this kind of deposit. Lithium has 

been produced from brine in Clayton Valley since the 1960s. 

It is not a particularly notable basin—it has a relatively small 

footprint of about 100 km2 and a total lithium resource of 

about 0.3 million metric tons—but it is better understood 

than the rest.

Clayton Valley is one of about 150 basins in the Basin 

and Range Province of the Western United States that formed 

when the western 1,000 km of North America was profoundly 

reshaped because of interactions with adjacent plates in the 

Pacific Ocean. A computer animation of this process in available on the Internet (Atwater, 2011). In the past 15 million 

years, Nevada has been stretched in an east-west direction 

to about twice its former width. The basins are one of many 

consequences of this process. The basins are relatively flat 

regions where the Earth’s former surface has subsided over 

geologic time, and the resulting depressions have been 

filled with sediment, typically hundreds to thousands of 

meters thick.

The Basin and Range Province is dry. Its aridity is 

partly owing to its position in the northerly of Earth’s two 

low-latitude dry belts and partly to a double rain-shadow 

cast by two of the highest mountain ranges in the contiguous 

United States (the White Mountains and the Sierra Nevada, the 

latter of which reaches 4,421 m, or 14,505 feet). The region 

is so dry that most of the basins are closed; that is, surface 

waters (mostly snowmelt) can flow in but not out. Inflowing 

waters bring in dissolved calcium, magnesium, potassium, 

sodium, and less abundant elements, such as lithium. These 

elements have built up over time, either in permanent lakes 

(such as Utah’s Great Salt Lake) or in ephemeral lakes that 

sometimes dry out completely, leaving a salar (salt pan or 

playa) covered by evaporite minerals. Halite (sodium chloride, 

or common rock salt) is the best known of these minerals. As a 

salt lake dries out, lithium behaves somewhat differently than 

other dissolved metals because it is more soluble. Rather than 

crystallizing an evaporite mineral, lithium remains in solution 

and ends up in a bitter-tasting residual brine. This brine is 

denser than fresh water and sinks into the subsurface where 

it collects in one or more aquifers (zones or layers of porous 

and permeable rock or sediment). At Clayton Valley, lithiumenriched brine is pumped from six different aquifers that all lie 

within a few hundred meters of the surface.

The lithium in Clayton Valley is thought to have come 

from various sources; some of it was weathered from rocks 

or sediments by rain, snowmelt, or groundwater, and some 
K10 Critical Mineral Resources of the United States—Lithium

Increasing fractionation, volatile enrichment,

complexity of zoning, and extent of replacement

Lithium Figure K7

Fault

Be 

Pegmatitic granite Li, Be, Nb>Ta 

Li, Be, Ta, Sn, Cs, Rb

Limit of pegmatite halo

Parental granite

Metasedimentary

or meta-igneous

country rock

NOT TO SCALE

EXPLANATION

Pegmatite bodies with

 characteristic enrichments 

Li, Be, Ta, Sn, Cs, Rb 

Li, Be, Nb>Ta

Be

Figure K7. Schematic cross section showing the concentric arrangement of lithium-cesium-tantalum (LCT)

pegmatites (small purple, blue, and green bodies) around a parental granite pluton. In this model, common 

pegmatites form near the parent, whereas pegmatites with enrichments in incompatible elements (indicated 

by chemical symbols) and corresponding rare minerals form farther away. Modified from Galeschuk and 

Vanstone (2005) and Trueman and Černý (1982). Be, beryllium; Cs, cesium; Li, lithium; Nb, niobium; Rb, 

rubidium; Sn, tin; Ta, tantalum

was brought in by hydrothermal waters that rose from anomalously hot rocks below the basin (fig. K8). Before intense 

pumping disrupted the groundwater system beneath Clayton 

Valley, these hydrothermal waters came to the surface at two 

hot springs.

The arid climate needed to form the brine is also a key to 

production, because the first step in the extraction of lithium 

is solar evaporation. In Clayton Valley, brine carrying an 

average of 160 ppm lithium is pumped to the surface and 

evaporated in a succession of nine artificial ponds that cover a 

total area of 16 km2 (fig. K3E ) (Garrett, 2004). In each pond 

in the chain, brine enters at one end, loses some of its water 

during the ensuing weeks or months, and is drained or pumped 

from the other end into the next pond. Multiple ponds are 

used so as to keep separate the various evaporite minerals that 

crystallize out in sequence. It takes nearly 2 years to achieve 

a concentrate that is both enriched in lithium (to 5,000 ppm) 

and depleted in other, more abundant, elements. This “mother 

liquor” is pumped from the last pond to a chemical plant for 

production of lithium carbonate and lithium hydroxide. A 

more complete description of the extraction of lithium from 

brine is provided in Garrett (2004).

The same general conditions that made the Clayton 

Valley brines have created even larger deposits in the Andes 

Mountains of South America and in Tibet in central Asia 

(fig. K5). The world’s most significant lithium district is the 

Salar de Atacama, on the Altiplano of Chile (fig. K3D), which 

is one of the driest places on Earth. The Salar de Atacama is a 

huge salt flat that covers an area of about 2,100 km2, which is 

roughly two-thirds the size of Rhode Island. It is surrounded 

by active volcanoes and cut by active faults. The salar is fed 

by a combination of snowmelt from nearby mountains and 

hydrothermal fluids from below. At an average of 1,400 ppm 

lithium, the brine is far richer than at Clayton Valley. The 

total reserve is more than 20 times greater, at an estimated 

7.5 million metric tons of lithium.
Geology K11

Lithium sequestered around

hot springs in clays 

Lithium weathered from older rocks

Lithium delivered

in windblown dust

Lithium in airfall tuffs

from nearby volcano

NOT TO SCALE

Preconcentrated lithium

from exhumed basinal strata 

Evaporation

ponds

Lithium delivered in solution

from connected basin

 Convecting

basinal brines

Lithium delivered

 in magmatic

 fluids 

Evaporative

concentration

of lithium

Mechanisms for concentrating lithium in brines

• Evaporation of water from brines

• Hydrothermal fluids react with aquifer and liberate lithium

Mechanisms for removing lithium from brine pool

• Brine spills out of basin

• Brine leaks out from bottom of basin

• Lithium minerals crystallize from saturated brine

• Lithium clays crystallize from hydrothermal fluids

• Lithium brines are trapped in fluid inclusions in halite 

Potential sources of lithium in brine 

1. Older bedrock

2. Primary magmatic or hydrothermal fluids

3. Volcanic ash

4. Loess

5. Exhumed basin deposits (recycled lithium)

6. Regional groundwater flow

Salar with

lithium brine

Deep

groundwater

Neogene volcano

Salar

Proximal alluvial fan

Other modern basin fill

Older basin fill, now exhumed

Bedrock, undifferentiated

Fault

Lithium Figure K8

EXPLANATION

Magma or advected

 hot rocks at depth

1

1

1

6

2

6

5

2

3

4

Figure K8. Conceptual ore-deposit model for lithium brine. An extensional basin is shown, but basins in active 

contractional and strike-slip tectonic settings are also prospective, the requirement being that the basin is closed. 

From Bradley and others (2013) and Munk and others (2016).
K12 Critical Mineral Resources of the United States—Lithium

Another notable deposit is the Salar de Uyuni in Bolivia. 

With an estimated 9 million metric tons of lithium at an 

average concentration of about 320 ppm (Jaskula, 2013), it is 

likely the largest accumulation of lithium in the world (Gruber 

and others, 2011). Despite its size, it is not yet in production. 

One difficulty is an abundance of dissolved magnesium, which 

hampers the evaporative concentration of lithium by causing it 

to precipitate prematurely as an impurity in magnesium salts.

The most noteworthy of several lithium brine systems 

in China is at Zabuye Lake, Tibet. It is the only known 

closed basin in which a lithium salt (specifically, Li2CO3, 

or zabuyelite) precipitates as part of the evaporite mineral 

sequence, and the only basin in a recent global compilation 

where volcanism or hydrothermal activity has not been 

reported (Munk and others, 2016). The brines have an average 

concentration of 700 ppm and an estimated resource of about 

1.5 million metric tons of lithium (Gruber and others, 2011).

Lithium in Other Brines

Deep oilfield brines may contain up to several hundreds 

of parts per million lithium. Evans (2008) reported high 

tonnages of lithium contained in oilfield brines in Arkansas, 

North Dakota, Oklahoma, Texas, and Wyoming, with lithium 

concentrations of up to 700 ppm. One of the best-known of 

the oilfield brines is contained in the Smackover Formation in 

the Gulf Coast area of the south-central United States (Collins, 

1976). In some places, the brine contains up to 692 milligrams 

per liter (mg/L) lithium. The brine occupies pore space 

in an approximately 200-m-thick limestone at depths of 

1,800 to 4,800 m. The brine has been interpreted to be trapped 

seawater that was subsequently hydrothermally enriched in 

lithium and other trace elements (Garrett, 2004).

Oilfield brines have two drawbacks as potential lithium 

resources. First, they typically occur at much greater depths 

(greater than 1 km) than closed-basin brines. Second, unless 

they happen to be located in an arid climate, recovery of 

lithium using the expedient and inexpensive method of solar 

evaporation will not be feasible (Bradley and others, 2013).

Geothermal brines are another potential source of 

lithium. These fluids traditionally derive their value from 

their contained heat, which can be converted to mechanical 

energy—but some geothermal fluids contain anomalous 

dissolved metals, including lithium. Reportedly, lithium is 

now being recovered by Simbol, Inc. from a geothermal brine 

in the Salton Sea area along the California-Mexico border 

(Simbol, Inc., 2013).

Lithium-Clay Deposits

A small subset of the world’s clay deposits is enriched in 

lithium. Lithium-bearing clay deposits contain an estimated 

7 percent of the world’s lithium resources (Evans, 2012). 

Analysis of lithium-enriched clays from several occurrences in 

the Western United States revealed that they all show enrichments in magnesium, fluorine, and lithium—a combination 

attributed to hydrothermal processes (Stillings and Morissette, 

2012). Hectorite (fig. K3F), which is a member of the smectite 

family of clay minerals, appears to be the most widespread, 

but Stillings and Morissette (2012) also found lithiumenriched clays of the illite family, as well as illite-smectite 

mixtures. Hectorite is a silicate of sodium, magnesium, and 

lithium and contains about one-half percent lithium (table K1). 

Lithium-clay deposits, prospects, and occurrences are shown 

in figure K5.

The most significant lithium-clay resource so far 

discovered is the Kings Valley hectorite deposit, located 

within the McDermitt caldera complex, which is an extinct 

volcanic center in northern Nevada. The McDermitt area 

had extensive volcanism during the Miocene; at least five 

collapsed vents and resurgent domes have been recognized 

within the complex. The lithium clays occur in hydrothermally 

altered, volcanic-derived sediments of lakes that occupied the 

caldera (Rytuba and Glanzman, 1979). Recovery of lithium 

by leaching the clay with sulfuric acid has proven feasible 

(Eggleston and Hertel, 2008). The deposit is being developed 

for production, but lithium production has not yet begun 

(Western Lithium Corp., 2015).

Clays that happen to contain lithium have been mined in 

Arizona, California, and elsewhere in Nevada. These deposits 

were or are mined not for their lithium content but rather for 

their physical properties, which make them good additives 

to, in order of value, drilling muds, paints, and cosmetics. 

Hectorite gets its name from such a deposit at Hector, Calif., 

where lake beds of volcanic ash were altered by hydrothermal 

fluids that came up along a fault zone (Ames and others, 

1958). At least two other deposits in the Southwestern 

United States have been similarly interpreted (Norton, 1965; 

Wilkerson and others, 2001). In Turkey, the world-class 

Bigadiç borate deposit, which formed in the hydrothermally 

altered sediments that fill a rift-related lacustrine basin, 

contains associated hectorite (Helvaci and others, 2004).

Lithium-Zeolite Deposits

The only documented lithium-zeolite deposit is from a 

Neogene basin system in the Balkan region of Eastern Europe 

(fig. K5). Miocene-age lake beds in the Jadar basin include 

oil shale, carbonate rocks, evaporites, and tuff (Obradović 

and others, 1997). These strata have been authigenically 

overgrown by massive layers of jadarite, which is a recently 

recognized boron-lithium silicate mineral of the zeolite 

family (Stanley and others, 2007). The jadarite layers are 

reportedly several meters thick. This single jadarite deposit 

contains an estimated 3 percent of the world’s lithium resource 

(Evans, 2012).
Resources and Production K13

Resources and Production

Production

Global annual production of lithium has been increasing 

since the mid-1900s; in 2012, it was about 37,000 metric tons 

(fig. K1; Jaskula, 2013). The world has hundreds of lithium 

deposits, prospects, and occurrences (figs. K4 and K5) but only 

a few of them are currently producing. The present-day supply 

of lithium comes from pegmatite deposits and closed-basin 

brines. The main pegmatite producers are Australia, Brazil, 

China, Portugal, and Zimbabwe. The main brine producers are 

Argentina, Chile, China, and the United States. The various 

other deposit types have potential as future contributors.

The United States imports most of the lithium that it 

consumes. In 2012, the country’s net import reliance was 

greater than 70 percent (Jaskula, 2013).

Reserves and Other Identified Resources

Global lithium resources total more than 39 million metric 

tons, which is nearly three thousand times current annual 

production. Of this resource, the USGS currently estimates 

there to be approximately 13 million metric tons of reserves; 

that is, resources that are demonstrated (measured and indicated categories) to be and are currently economically recoverable. These figures have changed in recent years, and various 

experts have reported a wide range of values based on somewhat different methods and assumptions (Gruber and others, 

2011). An important additional aspect of the global lithium 

outlook is recycling, which was not figured into the above 

resource estimates. Lithium-ion battery recycling is expected 

to play a major role in the supply of lithium in the medium to 

long term. Lithium can be recycled indefinitely, whereas oil can 

be used only once. Resources of lithium are expected to satisfy 

demand into the next century (Evans, 2012).

Although current domestic lithium production is small, 

the United States is endowed with a substantial lithium 

resource. Gruber and others (2011) estimated the domestic 

lithium resource to be about 10.2 million metric tons. This 

includes the Kings Mountain pegmatite district in North 

Carolina (5.9 million metric tons of lithium), the Kings 

Valley hectorite deposit in Nevada (2.0 million metric tons), 

and various closed-basin, geothermal, and oilfield brines 

(2.3 million metric tons).

Undiscovered Resources

Exploration for lithium deposits has never been as intense

as it has been for other metallic resources, such as gold and

copper, so it is likely that large, economically viable lithium 

deposits remain to be discovered. Now that a potentially very 

large lithium battery market exists for hybrid and electric 

vehicles, more exploration is being done, and new deposits 

will likely be discovered. The USGS contributes to this effort 

by developing mineral deposit models, which provide a 

conceptual basis for modern mineral exploration. New models 

have been developed for both LCT pegmatites (Bradley and 

McCauley, 2013; Bradley and others, (2017) and lithium brines 

(Bradley and others, 2013; Munk and others, 2016).

Plots of grade versus tonnage are shown in figure K9 

for lithium in pegmatites and brines; these give an indication 

of the sizes of deposits that might remain to be found. The 

two parameters, deposit grade and deposit tonnage, vary 

across a wide range of values for most of the common lithium 

deposit types.

1 10 100 100

100

10

1,000

10,000

1,000 10,000 100,000

1.0

0.1

10

Deposit tonnage, in million metric tons of resource Deposit tonnage, in million metric tons of brine

Lithium Figure K9

Whabouchi

Tanco

Great Salt Lake

Atacama

Clayton Valley

Greenbushes

Grade, in weight percent Li O2

Grade, in parts per million Li

A. Lithium resources in LCT pegmatites B. Lithium resources in closed-basin brines

Figure K9. Plots of lithium grade and tonnage for selected world deposits. These plots give an indication of the sizes 

of deposits that might remain to be found for A, lithium-cesium-tantalum (LCT) pegmatites, and B , closed-basin brines. 

Data for the LCT pegmatites are from Bradley and others (2017), and data for the closed-basin brines are from Munk 

and others (2016). Some of the deposits discussed in the text are labeled.
K14 Critical Mineral Resources of the United States—Lithium

Exploration for New Deposits

Exploration for Lithium-Cesium-Tantalum 

Pegmatite Deposits

The search for new LCT pegmatite districts begins with 

the knowledge that all LCT pegmatites were emplaced into 

orogenic belts (Bradley and McCauley, 2013). The global age 

distribution (fig. K6) provides a first-order filter in the search 

for new LCT pegmatites in unexplored regions. The potential 

for giant deposits seems greatest in Archean (2,500 Ma 

and older) collision zones. Most LCT pegmatites intruded 

metamorphosed sedimentary rocks, typically at conditions 

of medium pressure and medium to high temperatures 

(greenschist to amphibolite facies) (Černý, 1992). New 

deposits are likely to be found within a few kilometers of a 

“fertile” granite, which generally are large enough (typically 

several square kilometers in area) to be known already from 

routine geologic mapping. Fertile granites are identified 

by the presence of distinctive minerals, such as muscovite, 

tourmaline, and garnet; by anomalously high concentrations 

of such trace elements as cesium, lithium, rubidium, tin, and 

tantalum; and by low concentrations of the major elements 

calcium, iron, and magnesium (Selway and others, 2005). 

Around a fertile granite, the more distal LCT pegmatites are 

more likely to be enriched in lithium (fig. K7). Weathering 

of LCT pegmatites can result in both soil anomalies and 

indicator minerals. Smith and others (1987) demonstrated that 

arsenic, beryllium, antimony, and tin form a 12-by-20-km 

halo in lateritic soils around the Greenbushes pegmatite; 

niobium, tantalum, and boron form a smaller, 1-by-5-km halo. 

Cassiterite, tantalite, elbaite, and spessartine are sufficiently 

dense and durable to serve as heavy indicator minerals that 

can be found by panning unconsolidated sediments.

Exploration for Lithium-Brine Deposits in 

Closed Basins

The most critical factor that determines if a sedimentary 

basin can accumulate lithium brine is whether or not the 

basin is closed. All the lithium-brine basins in figure K5 are 

closed, and all have a saline lake or a salar (that is, a salt flat 

or salt-encrusted depression). Closed basins form because of 

tectonic processes, but they are maintained only where, over 

longer timespans, evaporation exceeds precipitation. If the 

long-term rate of precipitation in a basin increases sufficiently, 

lake water will eventually overtop some point along the 

drainage divide and drain away, carrying with it any dissolved 

lithium. Mean annual temperature does not appear to be a 

critical factor; values range through an order of magnitude 

from about 0 degrees Celsius (°C) to 23 °C (Munk and others, 

2016). Judging from the present-day population of basins, the 

favorable zones lie between about 19° and 37° north or south 

latitude (Bradley and others, 2013; Munk and others, 2016). 

Lithium-brine resources are found in a range of active tectonic 

settings, from extensional, to strike-slip, to contractional 

(Bradley and others, 2013; Munk and others, 2016). All but 

one of the lithium-brine basins show evidence for elevated 

heat flow, as revealed by the presence of young volcanoes, 

hot springs, and (or) geothermal resources. Evidence of older 

hydrothermal activity is seen in several instances and should 

be regarded as a favorable characteristic. For example, at 

Clayton Valley, some Miocene to Pliocene basinal deposits 

were hydrothermally enriched in hectorite; this probably took 

place a few million years ago, before the lithium-enriched 

basinal strata were uplifted and exposed. Hectorite concentration haloes have been reported from several basins that lack 

lithium brines in the U.S. Basin and Range Province (Vine, 

1980). In such cases, a hectorite halo may be all that remains 

of a former brine system that has since leaked away. Like 

lithium brines, borate deposits appear to be hybrids involving 

both hydrothermal and evaporative processes; their presence 

is a favorable indicator for lithium brines. Source rocks, such 

as felsic, glassy tuffs that have abundant and readily leached 

lithium, are favorable indicators of a nearby lithium brine 

deposit (Price and others, 2000).

Exploration for Lithium in Sedimentary Rocks 

(Clays and Zeolites)

The depositional environments of deposits, prospects, 

and occurrences in the Basin and Range Province share the 

following characteristics: they are all found in lake sediments, 

the sediments contain ash or other volcanic materials, and 

there is evidence of alteration by hydrothermal fluids (AsherBolinder, 1982, 1991). Borate deposits share the same general 

characteristics, and, accordingly, the presence of borates is a 

favorable indication for lithium clays. In the Basin and Range 

Province, borate deposits were extensively prospected in the 

late 1800s and early 1900s, and many deposits and prospects 

were found. Lithium clays are not as obvious in these settings 

as are the common borate minerals; significant lithium 

resources may thus have been overlooked.

The lithium zeolites at Jadar, Serbia, compose the only 

known deposit of this type. Accordingly, exploration guidelines must be based on this occurrence, which formed in an 

extensional tectonic regime and is hosted in Miocene-age lake 

beds that include oil shale, carbonate rocks, evaporites, oil 

shale, and tuff (Obradović and others, 1997).
Environmental Considerations K15

Environmental Considerations

Sources and Fate in the Environment

Lithium in the environment can be derived from both 

natural and anthropogenic sources. As noted already, lithium is 

present in trace quantities in most rocks and natural waters and 

is a major component only in certain rare pegmatite minerals, 

clays, zeolites, and brines. Potential anthropogenic lithium 

sources include mine wastes, and after disposal, the wide 

range of manufactured products.

In soils, lithium is found in detrital and authigenic 

minerals and, to a lesser extent, in the organic fraction 

(Heier and Billings, 1978; Ribas, 1991; Schrauzer, 2002). 

Its concentration depends upon the parent rock (table K2); 

a study of Scottish soils shows the highest lithium concentrations are found in soils that develop over claystones and 

metamorphosed felsic rocks. The top 10 centimeters of a soil 

profile can be poorer in lithium than lower layers because 

lithium is leached by weathering solutions and can be 

incorporated into underlying clay minerals (Mitchell, 1964; 

Ribas, 1991). Shacklette and others (1973) observed that soils 

from the Western United States are more enriched in lithium 

than are those in the Eastern United States, with mean values 

of 23.3 ppm lithium and 17.3 ppm lithium, respectively. 

Cannon and others (1975) found a difference between the 

lithium contents of sediments in hydrologically open and 

closed basins in the Basin and Range Province of California 

and Nevada. In open basins, the lithium concentrations are 

relatively low, ranging from 0 to 150 ppm. In contrast, the 

range of lithium concentrations in hydrologically closed basins 

is 30 to 2,000 ppm, and lacustrine clays have a median value 

of 700 ppm lithium. The difference is owing to hydrological 

flow patterns; in open basins, dissolved salts are continually 

carried off by outflowing waters.

Although lithium is not required for the growth of plants, 

they are known to assimilate it, typically in the range of 

0.2 to 30 ppm (Aral and Vecchio-Sadus, 2008). More extreme 

values have also been reported; for example, matrimony 

vine (Lycium barbarum) from Arizona contained 1,120 ppm 

lithium (dry weight) (Sievers and Canon, 1973). A study of 

plants from 20 basins in Nevada and California demonstrated 

that native plants grown in arid basins absorb more lithium 

than those in more humid regions. Cannon and others (1975) 

reported median lithium concentrations of 22.8 ppm (dry 

weight) from Magnoliopsida (also known as dicotyledons) 

in arid basins compared with 1.3 ppm from humid regions of 

the United States (Bertrand, 1952, 1959). Cannon and others 

(1975) reported that plants growing in hot-spring discharge 

zones appear to have died from accumulation and precipitation 

of evaporative salts; these “fossilized root crowns” of salt can 

contain 1,000 ppm lithium.

Being a highly soluble element, lithium is commonly 

found as a dissolved species in both groundwater and surface 

water. This property makes lithium useful as a conservative 

tracer in hydrogeological studies (Bencala and others, 1990). 

There is evidence, however, that lithium may be removed 

from solution through sorption onto suspended clays. For 

example, in Saline Valley, Calif., 98 percent of lithium is lost 

from solution during surface flow from the spring to the playa 

(Lombardi, 1963). Lithium substitutes readily for magnesium 

in the clay structure because of the similar ionic radii of 

the two elements, and this isomorphic substitution is likely 

responsible for the affinity of lithium for clays. Plant uptake 

is another mechanism by which lithium can be removed from 

solution, especially soil solution. There is a much greater 

uptake of lithium by plants in acid soils, as acidity increases 

the solubility of metals (Schrauzer, 2002; Lenntech, 2007).

The use of lithium batteries is anticipated to increase 

in the future in response to the increasing demand for 

electric vehicles. The increased use of lithium in batteries is 

in turn expected to result in increased recycling of lithium 

from batteries. Recent press reports of explosions and fires 

associated with lithium batteries (for example, Mouawad and 

Drew, 2013) could cause concern about lithium recycling. 

Spent lithium batteries contain virtually no lithium metal, 

however, because it has all been converted to lithium 

oxide by the time the battery stops working (National 

Electrical Manufacturers Association, 2001). Lithium oxide 

is unreactive; unlike lithium metal, it does not produce 

hydrogen gas upon exposure to water, and it does not catch 

fire. Both HMIS (Hazardous Materials Identification System) 

and NFPA (National Fire Protection Association) give it a 

“0” rating for flammability and physical (reactivity) hazard 

risks (Alpha Aesar, 2012; Sigma Aldrich, 2014). It is a 

health hazard (a “3” rating by HMIS and NFPA) because 

it is corrosive and can cause serious eye and skin damage 

upon contact. Disposal of spent batteries in a modern 

municipal solid waste facility is not likely to present a lithium 

hazard, however. Lithium oxide is slightly to moderately 

soluble—a maximum of 6.67 g of Li2O can dissolve in 

100 g of water (H2O) at 0 oC (ESPI Metals, 1993).

Table K2. Lithium concentrations (in parts per million) in soils developed on various types of bedrock.

[Sources: Mitchell (1964), as reported in Heier and Billings (1978)]

Igneous Metamorphic Sedimentary

Olivine gabbro Andesite Granite Serpentinite Quartzite Granitic gneiss Quartz mica schist Shale Sandstone

30 50 7 30 15 70 200 60 20
K16 Critical Mineral Resources of the United States—Lithium

Mine Waste Characteristics

Mine-waste characteristics vary with the type of lithium 

deposit. Only two types of deposits—pegmatites and closedbasin brines—have produced mine wastes.

The mining history of lithium pegmatites may affect the 

mineralogy of tailings and waste piles that remain. Tin was 

the original target of mining at Kings Mountain (Kesler, 1942, 

1955, 1961), suggesting that lithium-bearing minerals may 

still remain in dumps from that era. Mineral processing of the 

pegmatite minerals involved crushing, wet grinding, sieving, 

gravity concentration, flotation, and collection with fatty acid 

amines. Tailings were discharged to storage ponds (Garrett, 

2004) that likely contain some lithium. In 1969, mineral 

processing was changed to improve the recovery of lithium. 

The new process involved calcining (heating) the spodumene 

concentrate after collection, then leaching the calcine with hot 

water. Yet even with this newer process, all the lithium still 

was not recovered.

Another source of lithium to waste and tailings piles can 

be through dissolution of the lithium minerals in the grinding 

and sieving steps of processing. For example, spodumene is 

fairly insoluble in water and dilute acids, but a small amount 

of dissolution has been observed during ore processing and 

concentration (Aral and Vecchio-Sadus, 2008). The dissolved 

lithium will be transported with the processing water and may 

concentrate in tailings ponds that have concentrations ranging 

up to 13 ppm.

The production of lithium by evaporation of closed-basin 

brines produces various evaporite minerals in the sequence of 

ponds. Details vary among the deposits owing to differences 

in brine chemistry. Precipitates at Clayton Valley include 

clay, calcium carbonate, calcium sulfate, hydrated magnesia, 

potassium-sodium sulfate, and potassium-sodium chloride 

(Garrett, 2004).

Human Health Concerns

Lithium is not considered an essential element for life, 

but it is present in most organisms in trace quantities (Léonard 

and others, 1995; Lenntech, 2007). The normal human intake 

is about 2 milligrams per day (mg/d) (Léonard and others, 

1995), and the human body contains about 7 milligrams (mg). 

Ingested lithium is absorbed from the intestines and excreted 

through the kidneys, and the half-life in the body is about 

24 hours (Freeman and Freeman, 2006; Aral and VecchioSadus, 2008).

In the realm of human health, lithium is best known as 

the leading treatment for the debilitating symptoms of bipolar 

disorder. Modern use of lithium compounds for this purpose 

dates from pioneering work after World War II in Australia 

by John Cade (Mitchell and Hadzi-Pavlovic, 2000). About 

4 percent of people—or nearly 300 million out of the present 

world population—will suffer from bipolar disorder at some 

point in their lives (Ketter, 2010). The number who treat the 

condition with lithium is unknown but undoubtedly is in the 

millions. A typical dosage is 1,800 mg/d of lithium carbonate 

(Drugs.com, 2013).

Lithium can be toxic. In fact, the United States banned the 

use of medicinal lithium between 1949 and 1970 for this reason 

(Strobusch and Jefferson, 1980). Aral and Vecchio-Sadus 

(2008) summarized the various harmful effects and noted that 

a dose of 5 g of lithium can be fatal to humans. No lithium 

poisoning has been reported from industrial applications, 

however (Aral and Vecchio-Sadus, 2008).

Ecological Health Concerns

Lithium affects metabolism, neuronal communication, 

and cell proliferation in many organisms in addition to 

humans. It disturbs the development of invertebrates (Léonard 

and others, 1995), and, in rats, it has been reported to reduce 

the number in a litter and the weight of the offspring and to 

cause incomplete ossification (Domingo, 1994). One of several 

useful endpoints used in toxicity tests is the concentration that 

leads to 50 percent mortality (lethal concentration 50, or LC50) 

after exposure to a substance for a certain amount of time. 

For earthworms exposed to soil containing lithium, the LC50

value was 70 milligrams per kilogram (mg/kg) after 7 weeks 

of exposure (Fischer and Molnár, 1997). The sensitivity of 

fish to lithium was measured in a series of 96-hour tests, 

which revealed an LC50 value that ranged from 13 mg/L for 

fathead minnows to greater than 100 mg/L for other fish 

species. Lithium concentrations of 1.2 mg/L immobilized the 

tiny crustacean Daphnia magna in 64 hours, and a dose of 

1.7 mg/L of lithium prevented the formation of embryos in 

fish eggs (Kszos and Stewart, 2003).

Carbon Footprint

Lithium’s potential importance to the world’s carbon 

budget derives from its use in lightweight, rechargeable 

car batteries. Over the course of a hypothetical lifetime of 

150,000 miles (240,000 km), today’s average gasolinepowered car will produce about 63,450 kilograms of carbon 

dioxide based on values from the U.S. Environmental 

Protection Agency, Office of Transportation and Air Quality 

(2011). This greenhouse gas will go directly from the tailpipe 

to the atmosphere, where it will contribute to global warming; 

some of it will later end up dissolved in seawater, where it will 

contribute to further ocean acidification, harming or killing 

fish and shellfish. The carbon dioxide output of an electric 

car depends largely on the means of charging its battery, but 

in all cases—even if the electricity comes from a coal-fired 

powerplant—the output of carbon dioxide is lower than for a 

gasoline-powered car (Notter and others, 2010). In the bestcase scenario, where the battery is recharged from renewable 

hydroelectric, wind, tidal, or solar power, the carbon dioxide 

output (beyond infrastructure costs) is negligible.
References Cited K17

Problems and Future Research

As is the case for most of the critical elements, lithium 

resources have not been as thoroughly studied as some 

conventional metallic resources, such as copper and lead. 

Basic research on the better-known lithium deposit types 

(pegmatites and brines) as well as on the newer deposit types 

(clays and zeolites) would help identify new resources and 

could ensure that a transition from fossil fuels to renewable 

energy sources is not limited by the supply of lithium.

For LCT pegmatites, the overarching question is why 

some orogenic belts are well-endowed with these deposits, 

whereas other seemingly similar belts are not? The connections between LCT pegmatites and tectonics have yet to be 

unraveled at the level of detail that would aid in exploration. 

Although most LCT pegmatites are found in orogenic belts 

that formed during continental collision, there is no obvious 

explanation for why particular pegmatites formed where and 

when they did. To address this problem would require tight 

geochronology for the pegmatite, for the orogenic belt in 

which it occurs, and for any associated sedimentary basins. 

Such an effort is now technologically possible.

A similar question applies to lithium brines—Why have 

these deposits formed in only a small subset of closed basins 

in arid regions? In the U.S. Basin and Range Province, for 

example, about 150 basins share the same basic characteristics, but only a few have lithium brines. Another key question 

involves the timescale of lithium-brine genesis—Are lithiumenriched brines like fossil fuels, or are they renewable on short 

time scales?

Lithium-clay (hectorite) and lithium-zeolite (jadarite) 

deposits have not been systematically studied. Mineral-deposit 

models would be useful to guide exploration efforts for both 

types. Lithium-clay deposits, which can be studied at the 

surface, are also of interest as potential examples of fossil 

lithium-brine systems that once operated at depth.

Lithium is neither an ecological nor a human health 

concern when it is present at ambient conditions. As global 

lithium consumption rises during the next few decades, 

however, environmental issues related to the production, 

routine daily use, and disposal of lithium products will 

warrant additional research.

Acknowledgments

We would like to thank technical reviewers Nora Foley 

and Klaus Schulz and volume editor John DeYoung for 

suggestions that significantly improved the manuscript. 

Keith Labay helped create the deposit location maps and 

Suzanne Nicholson provided additional editorial comments 

after the technical review was complete.

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Manganese

By William F. Cannon, Bryn E. Kimball, and Lisa A. Corathers

Chapter L of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–L

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter L

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract.........................................................................................................................................................L1

Introduction...................................................................................................................................................L2

Uses, Applications, and Consumption .............................................................................................L2

Strategic and Critical Resource Issues ..........................................................................................L2

Manganese Commodities ..................................................................................................................L2

Geology..........................................................................................................................................................L4

Geochemistry.......................................................................................................................................L4

Mineralogy...........................................................................................................................................L5

Deposit Types ......................................................................................................................................L6

Land-Based Deposits ................................................................................................................L6

Manganese Deposits in Marine Sedimentary Rocks .................................................L6

Manganese Deposits Without Iron Enrichments ...............................................L6

Iron-Related Manganiferous Sedimentary Deposits .........................................L9

Secondary Enrichment (Supergene) Deposits ............................................................L9

Seabed Deposits ........................................................................................................................L9

Ferromanganese Nodules .............................................................................................L10

Ferromanganese Crusts ................................................................................................L11

Resources and Production .......................................................................................................................L12

Manganese Reserves ......................................................................................................................L12

Other Identified Resources .............................................................................................................L12

Land-Based Resources ..........................................................................................................L12

Seabed Resources ..................................................................................................................L14

Ferromanganese Nodules .............................................................................................L15

Ferromanganese Crusts ................................................................................................L15

Resources in the United States ......................................................................................................L16

Land-Based Resources ..........................................................................................................L16

Aroostook County, Maine ..............................................................................................L16

Cuyuna Range, Minnesota ............................................................................................L17

Seabed Resources ..................................................................................................................L17

Undiscovered Resources ................................................................................................................L17

Exploration for New Deposits ..................................................................................................................L18

Environmental Considerations .................................................................................................................L18

Sources and Fate in the Environment ...........................................................................................L18

Mine Waste Characteristics ...........................................................................................................L21

Human Health Concerns ..................................................................................................................L21

Ecological Health Concerns ............................................................................................................L22

Carbon Footprint................................................................................................................................L22

Mine Closure......................................................................................................................................L23

Problems and Future Research ...............................................................................................................L23

References Cited........................................................................................................................................L24

Suggested citation:

Cannon, W.F., Kimball, B.E., and Corathers, L.A., 2017, Manganese, chap. L of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, 

and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and 

prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. L1– L28, http://dx.doi.org/10.3133/ 

pp1802L.
Manganese

By William F. Cannon, Bryn E. Kimball, and Lisa A. Corathers

Abstract

Manganese is an essential element for modern industrial 

societies. Its principal use is in steelmaking, where it 

serves as a purifying agent in iron-ore refining and as an 

alloy that converts iron into steel. Although the amount of 

manganese consumed to make a ton of steel is small, ranging 

from 6 to 9 kilograms, it is an irreplaceable component in the 

production of this fundamental material. The United States has 

been totally reliant on imports of manganese for many decades 

and will continue to be so for at least the near future. There 

are no domestic reserves, and although some large low-grade 

resources are known, they are far inferior to manganese ores 

readily available on the international market. World reserves 

of manganese are about 630 million metric tons, and annual 

global consumption is about 16 million metric tons. Current 

reserves are adequate to meet global demand for several 

decades. Global resources in traditional land-based deposits, 

including both reserves and rocks sufficiently enriched in 

manganese to be ores in the future, are much larger, at about 

17 billion metric tons. Manganese resources in seabed deposits 

of ferromanganese nodules and crusts are larger than those 

on land and have not been fully quantified. No production 

from seabed deposits has yet been done, but current research 

and development activities are substantial and may bring 

parts of these seabed resources into production in the future. 

The advent of economically successful seabed mining could 

substantially alter the current scenario of manganese supply 

by providing a large new source of manganese in addition to 

traditional land-based deposits.

From a purely geologic perspective, there is no global 

shortage of proven ores and potential new ores that could 

be developed from the vast tonnage of identified resources. 

Reserves and resources are very unevenly distributed, 

however. The Kalahari manganese district in South Africa 

contains 70 percent of the world’s identified resources and 

about 25 percent of its reserves. South Africa, Brazil, and 

Ukraine together accounted for nearly 65 percent of reserves 

in 2013. The combination of total import reliance for manganese, the mineral commodity’s essential uses in our industrialized society, and the potential for supply disruptions because 

of the limited sources of the ore makes manganese among the 

most critical minerals for the United States.

Manganese is the 12th most abundant element in Earth’s 

crust. Its concentration varies among common types of rocks, 

mostly in the range of from 0.1 to 0.2 percent. The highest 

quality manganese ores contain from 40 to 45 percent manganese. The formation of these ores requires specialized geologic 

conditions that concentrate manganese at several hundred 

times its average crustal abundance. The dominant processes 

in forming the world’s principal deposits take place in the 

oceans. As a result, most important manganese deposits occur 

in ancient marine sedimentary rocks that are now exposed on 

continents as a result of subsequent tectonic uplift and erosion. 

In many cases, other processes have further enriched these 

manganiferous sedimentary rocks to form some of today’s highest

grade ores. Modern seabed resources of ferromanganese nodules

cover vast areas of the present ocean floor and are still forming 

by complex interactions of marine microorganisms, manganese 

dissolved in seawater, and chemical processes on the seabed.

Manganese is ubiquitous in soil, water, and air. It occurs 

most often in solid form but can become soluble under acidic 

conditions. Manganese mining, like any activity that disturbs 

large areas of Earth’s surface, has the potential to produce 

increases in manganese concentrations that could be harmful to 

humans or the environment if not properly controlled. Although 

manganese is an essential nutrient for humans and most other 

organisms, overexposure can lead to neurotoxicity in humans. 

Workers at manganese mining and processing facilities have the 

greatest potential to inhale manganese-rich dust. Without proper 

protective equipment, these workers may develop a permanent 

neurological disorder known as manganism. Each manganese 

mine is unique and presents its own suite of potential hazards 

and preventative measures. Likewise, various nations have their 

own sets of standards to ensure safe mining, isolation of mine 

waste, treatment of mine waters, and mine closure and restoration. Interest in mining trace metals contained in ferromanganese 

nodules and crusts on the seabed has increased rapidly in the 

past decade. Prime areas for future research include overcoming 

the technological challenges presented by mining as deep as 

6,500 meters below sea level and understanding and mitigating 

the potential impacts of seabed mining on marine ecosystems.
L2 Critical Mineral Resources of the United States—Manganese

Introduction

Manganese, which is a ferrous metal, is an irreplaceable 

element in modern industrial economies, and it is consumed 

in large quantities by all industrialized nations. Manganese 

is important because of its desulfurizing, deoxidizing, and 

alloying properties, as well as its other chemical properties. An 

abundant and stable long-term supply of manganese is critical 

to the United States and other industrialized nations, as well as 

to developing countries. The United States has no manganese 

production or reserves of manganese ore and is totally reliant 

on imports of this mineral commodity.

Uses, Applications, and Consumption

In metallurgical applications, manganese is used 

predominantly as an alloying addition in steel. It is used 

as well in refining iron ore to metallic iron prior to the 

steelmaking process. Manganese has no known substitutes in 

the overall conversion of iron ore to steel, so steel cannot be 

produced without manganese. Steel and cast iron production 

together provide the largest market for manganese (historically 

accounting for 77 to 90 percent of manganese consumption 

in the United States), although manganese is also used as an 

alloy with nonferrous metals, such as aluminum and copper.

Nonmetallurgical applications of manganese include 

battery cathode production (mainly synthetic manganese 

dioxide); soft ferrites (manganese-zinc ferrites) used in electronics; micronutrient additives in fertilizers and animal feed 

(manganese sulfate and manganous oxide); water treatment 

chemicals (potassium permanganate and manganese dioxide); 

and other chemicals (manganese dioxide), such as those used 

as a colorant for automobile undercoat paints, bricks, frits, 

glass, textiles, and tiles. The product “manganese violet” 

is used for coloration in cosmetics, glazes used by artists, 

plastics, and powder coatings.

In the United States, manganese is imported in the 

following three principal forms: ore (minimally processed 

natural manganese concentrations), and the alloys ferromanganese and silicomanganese (both of which are products 

of the refining of natural ores). Because the United States 

imports nearly all the manganese that it consumes and because 

U.S. manufacturers typically do not maintain large stocks 

of manganese, the consumption of both ore and alloys can 

be estimated from the imports. For the period 2007 to 2011, 

the U.S. average apparent consumption of manganese from 

manganese dioxide, ferromanganese, and manganese metal 

was 739,000 metric tons of contained manganese. Steelmaking 

accounted for 78 percent of U.S. consumption; batteries, 

another 7 percent; and other uses, including chemicals, 

nonferrous alloys, and cast iron, the remaining 15 percent 

(Corathers, 2013a).

Strategic and Critical Resource Issues

The U.S. Department of Defense rates manganese as one 

of the most critical mineral commodities for the United States 

because it is both essential for industry and has no substitutes, 

and because of the potential for and the likely effects of 

supply disruptions (fig. L1; National Research Council, 2008). 

Geopolitical concerns have led to a long-term goal of identifying 

and ensuring supplies that are not susceptible to disruption.

On a global scale, manganese reserves—that is, material 

that can be mined economically under present-day conditions—

are more than adequate to supply the world’s industries 

with manganese for many decades, but those materials 

are unequally distributed geographically. Only a relatively 

few nations produce large amounts of manganese (three 

countries account for 60 percent of production in 2011). The 

United States is 100 percent reliant on imports. Many industrialized nations, including many countries in Europe, as well 

as Canada and Japan, are also largely reliant on imports for 

their supply of manganese.

Identified manganese resources include identified 

economic deposits (reserves) as well as a substantially larger 

amount of naturally occurring manganese-rich rocks that are 

not proven to be economically minable today but that could 

become economically viable in the future. Some types of 

deposits (including volcanogenic deposits, hydrothermal vein, 

and karst-related deposits), mines, and ore-forming processes 

have local importance but are only of minor significance from 

a national and global perspective.

Manganese Commodities

Most manganese materials in the United States are used 

in metallurgical processes. The major manganese commodities 

for this purpose are manganese ore and manganese alloys.

Manganese ore is the primary source of manganese used 

in the manufacture of manganese ferroalloys. Virtually all 

manganese ores are subjected to some form of beneficiation 

to achieve a concentrated product that has greater manganese 

content and fewer undesirable impurities. The manganese 

content of the more commonly used and traded ores, concentrates, nodules, and sinter for metallurgical purposes is in the 

approximate range of 38 to 55 percent. A manganese content 

of 48 percent is considered standard as a pricing basis.

From 2007 to 2011, global manganese ore production 

averaged 13.2 million metric tons of contained manganese, 

and production was led by, in order of percent of global 

production, South Africa, Australia, and China (table L1; 

fig. L1B). From 2008 to 2011, the United States imported a 

total of 1.9 million metric tons of manganese ore on a grossweight basis. The leading sources of these imports were, in 

order of the quantity supplied, Gabon, Australia, South Africa, 

and Brazil (fig. L2A). Between 2008 and 2012, the average 
Introduction L3

South Africa

21%

Australia

20%

China

Gabon 18%

11%

Brazil

7%

India

7%

Others

16%

0

2,000 

4,000 

6,000 

8,000 

10,000 

12,000 

14,000 

16,000 

18,000 

2007 2008 2009

Year

2010 2011

Contained manganese, in thousand metric tons

World production

Apparent consumption

EXPLANATION

A

Manganese Fig 1

B

Figure L1. A, Bar chart showing world production of manganese 

ore (contained manganese) from 2007 to 2011 compared with world 

apparent consumption. Average annual production for the period 

was 13.2 million metric tons. B, Pie chart showing distribution of 

manganese ore (contained manganese) production, by country 

and percent of world total, for the same period (Corathers, 2013a; 

International Manganese Institute, 2012).

South Africa

51%

Other

21%

Republic

of Korea

7% Ukraine

8%

China

13%

Gabon 

61%

Other

6%

Brazil

5% South

Africa

7%

Australia

21%

South Africa

42%

Other

10%

Australia

11%

Norway

14%

Georgia

23%

Mangenese Fig 2.

C. Silicomanganese (gross weight)— 285,000 metric tons

A. Manganese ore (gross weight)—470,000 metric tons

B. Ferromanganese (gross weight)— 319,000 metric tons

Figure L2. Pie charts showing the sources (by country) 

and the annual average amounts (by percentage of the total 

supplied) of gross weight of U.S. imports of A, manganese ore; 

B, ferromanganese; and C, silicomanganese for the period 

2008–11. The tonnage numbers shown are the average 

annual gross weight of the imports. Data are from Corathers 

(2011–16).

Table L1. Estimated world manganese ore reserves in 

2012, in thousand metric tons of contained manganese.

[Data are from Corathers (2013b); NA, not available]

Country Reserves

United States None

Australia 97,000

Brazil 110,000

Burma NA

China 44,000

Gabon 27,000

India 49,000

Kazakhstan 5,000

Malaysia NA

Mexico 5,000

South Africa 150,000

Ukraine 140,000

Other countries Small

World total 630,000
L4 Critical Mineral Resources of the United States—Manganese

annual amount of manganese ore that was reported to the 

U.S. Geological Survey (USGS) as having been consumed 

domestically was 480,000 metric tons (gross weight).

The key manganese ferroalloys are ferromanganese, 

which is subdivided into standard (high-carbon), mediumcarbon, and low-carbon grades, and silicomanganese. 

High-carbon ferromanganese is smelted directly in either a 

blast furnace or an electric submerged-arc furnace (SAF), and 

silicomanganese is smelted in an SAF. Silicomanganese is the 

manganese ferroalloy that is used predominantly throughout 

the world, including in the United States. Silicomanganese 

is smelted in a similar manner to that of high-carbon ferromanganese, except that a more siliceous charge, usually in the 

form of quartz, quartzite, or an ore having high silica content, is 

used. When an SAF is used, high-carbon ferromanganese and 

silicomanganese can be produced interchangeably. Production 

of refined grades of manganese ferroalloys, such as mediumand low-carbon ferromanganese, involves two additional stages 

using such equipment as a converter or a direct arc furnace. 

From 2006 to 2010, global ferromanganese and silicomanganese production (by gross weight) averaged 5.4 million metric 

tons per year and 9.2 million metric tons per year, respectively.

As of 2013, there were two producers of manganese 

alloys in the United States—Eramet Marietta, Inc. and Felman 

Productions Inc. Eramet Marietta generally produces silicomanganese and all grades of ferromanganese at its plant in 

Marietta, Ohio. Felman Productions produces silicomanganese 

at its plant in Letart, West Virginia. Even with this domestic 

production, the United States imported a total of 1.3 million 

metric tons of ferromanganese and 1.1 million metric tons of 

silicomanganese (gross-weight basis) from 2008 to 2011. The 

leading sources of these imports, by country, are shown in 

figure L2 (B and C).

 The information and discussion presented in this chapter 

are based entirely on published scientific and technical literature 

and contain no new conclusions or interpretations. Rather, this 

chapter presents a summary of the vast literature on the geology 

and resources of manganese that are the result of more than a 

century of research on this important mineral commodity.

Geology

Geochemistry

Manganese is the 12th most abundant element in Earth’s 

crust and is thus considered to be a relatively abundant metal. 

Average crustal rocks contain about 0.1 percent manganese, 

but manganese concentration varies according to rock 

type. Mafic rocks, such as basalt, typically contain about 

0.16 percent manganese, whereas granitic rocks contain 

about 0.06 percent. Manganese ores contain a minimum of 

15 percent manganese, but most ores are considerably higher 

grade, ranging up to about 50 percent manganese. The formation of manganese ore, therefore, requires a geochemical 

system that concentrates manganese by at least 150 times its 

average crustal abundance and by as much as 500 times for 

the higher grade ores. An additional specialized set of conditions is required to produce strong manganese enrichment 

without concentrating iron, with which manganese shares 

many chemical similarities and is commonly compositionally 

correlated. Manganese ores have a much higher manganese-toiron ratio than do typical crustal rocks. Virtually all manganese 

ores have formed as a result of the transport of manganese in 

water solutions, at least for short distances, and precipitation 

of manganese minerals out of those solutions in sufficiently 

high concentrations and volumes to form minable orebodies. 

Understanding the factors that control manganese solubility in 

water and the behavior of manganese-bearing water at scales 

from meters to that of ocean basins is the key to understanding 

manganese metallogenesis.

Manganese is a potentially mobile element and occurs 

in three oxidation states in nature (+2, +3, +4). The solubility 

and mobility of manganese in various solutions is strongly 

controlled by the acidity (pH) and oxidation potential (Eh) of 

manganese-bearing solutions, which include surface water, 

ocean water, and high-temperature fluids in various subsurface 

environments. In general, manganese solubility is highest 

in more acidic (lower pH) conditions and less oxidizing 

(lower Eh) conditions in which dissolved manganese exists 

primarily in the most soluble +2 oxidation state. The transport 

of large quantities of manganese to sites of ore formation, 

therefore, generally requires the existence of oxygen-poor 

(low Eh) ocean water or surface water. This is in contrast to 

the typical state of modern surface and ocean water, in which 

the oxidation potential is sufficiently high to make manganese 

essentially insoluble. The variability in chemical behavior 

of manganese under differing pH and Eh conditions results 

in many situations in which manganese can be concentrated 

to potentially economic concentrations. A full review of the 

geochemistry of manganese as it relates to manganese ore 

formation is beyond the scope of this summary report; for 

greater detail and specificity than is provided here, see for 

example, Roy (1988), Varentsov (1996), Nicholson and others 

(1997), and Maynard (2003, 2010).

The vast majority of manganese ores have formed at 

the ambient conditions of Earth’s surface, both subaerial and 

submarine, and the discussion below concentrates on the 

geochemical behavior of manganese in these settings. Many 

manganese deposits are a direct result of deposition of manganese-rich sediment layers on the seabed, both in present-day 

oceans and oceans of the geologic past. In addition, secondary 

enrichment of manganese-rich marine sedimentary rocks by 

later continental weathering (supergene ores) has produced 

many high-grade manganese ore deposits. In both cases, 

large-scale transport of dissolved manganese in seawater and 

deposition of manganese from seawater are crucial components of manganese metallogeny. The most significant sources 

of dissolved manganese in seawater are (a) dissolved manganese in rivers entering the oceans, (b) manganese released 

from marine sediments by post-depositional alteration, and 

(c) manganese from deep-sea hydrothermal vents.
Geology L5

On continents, most manganese occurs in trace amounts 

in reduced (+2) form in silicate and carbonate minerals. It 

can be released by continental weathering where some of 

it is dissolved in surface water and eventually transported 

by rivers to the world’s oceans. Because surface waters are 

generally well oxidized, the solubility of manganese is very 

low. The Mississippi River, for instance, shows manganese 

solubility that varies seasonally but averages only a few 

micrograms per liter (Shiller, 1997). Although this is a low 

concentration, it is still much higher than manganese in open 

oceans. Substantially more manganese is delivered by rivers 

to the oceans as fine-grained suspended sediments, which 

either contain manganese within minerals or have manganese 

adsorbed to their surface. Although this manganese does 

not immediately affect the dissolved manganese pool of the 

oceans, it is an important flux of manganese to seabed sediments from which it is later mobilized by reducing reactions 

in the upper layers of the sedimentary sequence.

Seabed hydrothermal vents are also a significant source 

of manganese. The character of these fluids is summarized 

by Tivey (2007). Vent fluids commonly have manganese 

concentrations measured in thousands of micrograms per liter, 

which is several orders of magnitude greater than typical river 

waters. Such vents are arrayed along active spreading centers 

in all the world oceans.

The formation of marine sedimentary manganese 

deposits, in which the contained manganese is delivered to 

the site of sedimentation by seawater, generally requires that 

such seawater be less oxidized than typical seawater in order 

to sustain an elevated solubility of manganese. Such seawater 

can exist both as layers of reduced oxygen (oxygen minimum 

zones) in open oceans and as highly reduced sulfidic seawater 

that can form in the deepwater mass in stratified oceans in 

restricted basins. Recent reviews of manganese mineralization (Maynard, 2003, 2010) provide further discussion 

of these conditions, as does the discussion below regarding 

deposit types.

Mineralogy

Manganese occurs as a trace element in many minerals, 

but the ore minerals are limited to a suite of manganese 

oxides, hydroxides, carbonates, and silicates. Manganese ores 

are commonly fine-grained or microscopic mixtures of several 

different manganese minerals whose accurate characterization 

has long been a problem. They can seldom be accurately 

identified solely by visual examination. Modern analytical 

techniques have led to revisions of older data on the composition and mineral structure of many manganese minerals. 

This has led, in turn, to revised descriptions in more recent 

literature of the mineralogy of many manganese deposits.

 A full list of manganese ore minerals classified by 

deposit type is given in Dorr and others (1973), and details 

of chemical composition, crystal structure, and relative 

abundance are provided by Maynard (2003). In spite of 

the complexity of manganese minerals and their variable 

occurrence in individual ore deposits, a few generalizations 

allow characterization of many aspects of the ore mineralogy. The most common manganese minerals within ore 

deposits are manganese oxides, followed by manganese 

carbonates. Manganese carbonates, however, are the most 

voluminous manganese minerals when all manganeseenriched rocks—many of which are sub-ore grade—are 

considered. Among oxides, pyrolusite (a simple manganese 

oxide) and cryptomelane (a potassium-bearing oxide) are 

the most widely reported, although not necessarily the most 

voluminous. The carbonate minerals are rhodochrosite (a 

manganese carbonate) and kutnahorite (a calcium-manganese 

carbonate). Rhodochrosite is by far the most widely recognized manganese carbonate, but kutnahorite is abundant in the 

vast Kalahari manganese ores of South Africa. In addition, the 

manganese hydroxide manganite is a very widely occurring 

ore mineral, as is the manganese silicate mineral braunite. 

Descriptions of some of the characteristics of these most 

abundant of the manganese minerals are given below.

Rhodochrosite.—The carbonate mineral rhodochrosite 

(MnCO3) is the most widely reported manganese mineral in 

known deposits, including currently subeconomic deposits. It 

is the principal ore mineral at Molango, Mexico, and at many 

deposits in China. Production of a usable manganese product 

from rhodochrosite requires processing to convert manganese 

carbonate to manganese oxides. Rhodochrosite deposits are 

also important as protores (lower grade precursors for higher 

grade supergene deposits) in which oxidation and leaching at 

Earth’s surface has converted the original carbonate mineral 

to manganese oxides and hydroxides and commonly increased 

the manganese concentration to ore grade.

Braunite.—Braunite (Mn2+Mn3+

6 (SiO4)O8) is generally 

classified as a silicate mineral, although it contains less than 

10 weight percent silicon dioxide, or silica (SiO2). It is a major 

ore mineral in the Kalahari manganese district in South Africa; 

in Groote Eylandt, Australia; and in several districts in India. 

It is also a major mineral in volcanogenic manganese deposits, 

which are not discussed in detail in this chapter.

Cryptomelane.—Cryptomelane (K(Mn4+,Mn2+) 8

O16), a 

potassium-bearing manganese mineral, is a highly oxidized 

ore mineral in several major manganese-producing regions, 

including in Brazil, Gabon, and South Africa.

Manganite.—Manganite (MnO(OH)) is a manganese 

hydroxide that forms commonly by oxidation at or near 

Earth’s surface as a result of circulating groundwater. It is a 

significant mineral in major supergene deposits, such as those 

in Gabon and Ghana, as well as in parts of the Kalahari district 

in South Africa.

Pyrolusite.—Pyrolusite (MnO2) is a compositionally 

simple manganese oxide. It forms in a variety of oxidizing 

conditions, especially in cases of higher acidity (lower pH). 

It is a major ore mineral in many significant mining districts, 

including in Brazil, Gabon, Ghana, the Republic of Georgia, 

and South Africa.
L6 Critical Mineral Resources of the United States—Manganese

Deposit Types

Over the years, many classifications of manganese 

deposits have been proposed, typically based on details of 

mineralogy, type of associated rocks, and the geometric form 

of the deposits. For this summary, we use a simple classification of (a) land-based deposits consisting mostly of ancient 

marine sediments and zones of secondary enrichment developed within them that can be further subdivided based on the 

geologic environment in which they formed and the processes 

responsible for manganese concentrations, and (b) modern 

seabed deposits, which are found widely on the modern ocean 

floor. This discussion excludes several deposit types, such as 

volcanogenic, karst, and hydrothermal vein deposits, which 

are widespread but generally small. Although they may be 

important locally, their total manganese production potential 

is small relative to the three major deposit types (marine 

sedimentary, secondary enrichment, and seabed) and has only 

a small effect on national and global supply and demand. The 

global distribution of the major manganese deposits, their size, 

and deposit type are shown in figure L3.

Land-Based Deposits

Manganese Deposits in Marine Sedimentary Rocks

The vast majority of land-based manganese resources 

occur as extensive layers of manganese-rich sedimentary 

rocks, some of which formed as long as 2.5 billion years ago. 

These rocks formed on ancient seabeds (fig. L4) and have 

since become part of continents through tectonic processes of 

uplift and continental accretion. They may be of sufficiently 

high manganese content to constitute ores themselves, or they 

may be protores in which additional natural concentrating 

mechanisms increased the manganese content enough to 

form commercial ores. A recent compilation of the chemical 

composition of sedimentary manganese deposits of the world 

indicates that average manganese grade is about 24 percent 

and iron content is 4.3 percent (Maynard, 2010). These 

figures indicate that average sedimentary manganese deposits 

are low-grade ore, at best. Most ores currently being mined 

are the higher grade portion of these deposits or zones of 

secondary enrichment developed within them. These averages 

also illustrate an essential feature of sedimentary manganese 

deposits—the extremely efficient separation of iron from 

manganese during deposit formation. Whereas manganese is 

concentrated, on average, roughly 250 times its average crustal 

abundance in marine sedimentary deposits, iron occurs at 

concentrations near its average abundance.

Two types of sedimentary manganese deposits can be 

distinguished based on the nature of rocks with which they 

are interlayered and, by inference, the character of ocean 

water from which they were precipitated: (a) manganiferous 

sediments that occur independent of iron concentrations; and 

(b) manganiferous sediments interlayered with ferruginous 

strata, including the vast Kalahari deposits of South Africa that 

occur as interlayers in a banded iron formation. The essential 

difference between the two is the degree of oxygen depletion 

in ocean waters that varies from anoxic or suboxic (in which 

the solubility of both manganese and iron are enhanced—for 

instance, the Kalahari manganese district in South Africa), 

to euxinic and sulfidic (in which manganese solubility is 

enhanced but iron solubility is depressed—for instance, 

Groote Eylandt in Australia).

Manganese Deposits Without Iron Enrichments

Many sedimentary manganese deposits of the world 

occur in sedimentary sequences that are devoid of iron enrichments and that are themselves not enriched in iron compared 

with the average crustal abundance of manganese. The larger 

of such deposits include Molango in Mexico, Groote Eylandt 

in Australia, the deposits of the Black Sea region, and many 

deposits in China. These relatively low-iron deposits are 

thought to have formed in shallow marine settings adjacent to 

stratified oceans that contained a low oxygen deepwater mass, 

as illustrated in figure L5. In such water masses, dissolved 

hydrogen sulfide is locally present (euxinic conditions), which 

causes iron to be precipitated as sulfide minerals in black shale 

and leaves very low concentrations of dissolved iron in the 

seawater. Because there is no comparable manganese sulfide 

phase under these conditions, dissolved manganese concentrations (Mn2+) remain high (Cannon and Force, 1983; Force 

and Cannon, 1988). This type of ocean water is a reservoir 

from which manganese deposits can be formed by precipitation of the dissolved manganese caused by the mixing of 

the manganese-rich deep ocean water with the more oxidized 

surface waters. Because of the increased Eh conditions, 

the water becomes saturated in dissolved manganese, and 

manganese-rich sediments can be precipitated.

These conditions occur most often in nearshore shallow 

water on continental shelves (fig. L5). Consequently, manganese beds are interlayered with other sediments typical of 

those settings. The character of these types of manganese 

deposits is quite variable, depending on local conditions at the 

eventual deposition sites. For instance, the variety in the types 

of manganese ores described for the Groote Eylandt deposit 

are related to local variations in the seabed topography and 

distance from the coastline (Frakes and Bolton, 1992).

A modern analog for the creation of manganese deposits 

without iron enrichment is the Black Sea (Cannon and Force, 

1983; Force and Cannon, 1988). Deep euxinic waters of the 

Black Sea contain concentrations of manganese greater than 

300 micrograms per liter (μg/L), whereas oxidized surface 

water contains less than 1 μg/L. The concentration of iron 

in the euxinic water is less than 5 μg/L (Maynard, 2010). 

Thus, these deep waters are poised to precipitate manganeserich sediments devoid of iron enrichments when oxidized. 

Circulation of water in the Black Sea produces upwelling 

of deep waters onto a shallower shelf to the north where 

manganese-rich sediments are being precipitated on the 

present seabed.
Geology L7

180° 135° 90° 45° 0° 45° 90° 135° 180°

40°

30°

20°

10°

0°

10°

20°

30°

40°

80°

70°

60°

50°

50°

60°

70°

80°

Kalahari

Cuyuna

Range

Aroostock County

Molango

Serra do Navio

Azul

Urucum

Morro 

da Mina

China

Groote Eylandt

Imini

Tambao

Nsuta Moanda

Western Transvaal

Postmasburg

Urkut Varna

Nikopol

Bolshe Tokmak

Tchiatura

Gujarat

Karnataka

Goa Andhra

Pradesh

MadhyaPradesh

Odisha Clarion- Clipperton zone Pacific prime crust zone

Manganese fig 3

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Marine sedimentary, in thousand metric tons

<100,000100,000 to 10,000,000>10,000,000

Secondary enrichment, in thousand metric tons

<100,000100,000 to 10,000,000

Permissive areas for ferromanganese nodulesPermissive areas for ferromanganese crustsClarion-Clipperton zonePacific prime crust zone

Seabed deposits Land-based deposits

Large subeconomic deposits  in the United States

Many small to moderate size  sedimentary deposits—Total  resource not quantified

EXPLANATION

as subeconomic deposits in World map showing the location, relative size, and type of the major terrestrial manganese deposits listed in table L2 as well 

. Areas of greatest potential for seabed mining of ferro manganese nodules and crusts are shown by 

Figure L3.the United States. Deposits in China have been grouped togethercolored outlines. Seabed data are from Hein and others (2013).
L8 Critical Mineral Resources of the United States—Manganese

Figure L4. Photographs showing examples of contrasting types of manganese ore. A, Thinly layered ore from Molango, Mexico.

The manganese mineral is rhodochrosite, which is a manganese carbonate. The delicate laminations are the result of deposition 

in relatively deep, quiet seawater below the zone of surface wave agitation. B, Pisolitic ore from Groote Eylandt, Australia. The 

approximately spherical masses (pisolites) are mostly pyrolusite, which is a manganese oxide. The spheres were likely formed in 

shallow water by progressive growth of new outer layers as the masses were rolled on the seabed by wave agitation. Photographs 

by William F. Cannon, U.S. Geological Survey.

Sea level Island, peninsula, or shoal

Oxidized surface water Dissolved Mn and Fe low

Low-oxygen deep water;

dissolved Fe low; precipitated 

as Fe sulfide in black shale

Dissolved Mn high

Black shale

Gray shale

Mainland

Older rocks

Mn oxides Mn carbonates Sandstone

Zone of coastal upwelling and mixing of deep water with oxidized surface water. Deep Mn-enriched 

water becomes oxidized and precipitates manganese carbonate or oxides depending on local conditions

2 to 10 KILOMETERS

NOT TO SCALE

Figure 5 Manganese

100 to 200 METERS

Figure L5. Schematic diagram of the oceanic conditions necessary to form sedimentary manganese deposits that are not enriched 

in iron. Deep seawater that is very depleted in oxygen and hydrogen sulfide reacts with iron and precipitates iron sulfides, which 

results in low concentrations of dissolved iron, whereas the dissolved manganese is enriched. Upwelling and mixing of this deep 

water with oxidized surface water causes precipitation of manganese oxide and carbonate minerals, depending on the local 

conditions. Diagram is modified from Cannon and Force (1983). In a variant of this model, deep water that is more oxidized and 

does not contain hydrogen sulfide results in high concentrations of both dissolved manganese and iron and in the precipitation of 

ferruginous manganese. Fe, iron; Mn, manganese

A B
Geology L9

Iron-Related Manganiferous Sedimentary Deposits

Many occurrences of manganiferous sedimentary 

deposits occur interlayered with iron-rich strata, commonly in 

banded iron formations that are found widely in Precambrian 

sequences in many parts of the world. These include the vast 

manganese deposits of the Kalahari district in South Africa 

and the Urucum mining district in Brazil, and many deposits 

in India. There is voluminous literature on the Kalahari 

manganese deposits. Relatively recent summaries and the 

references therein contain the essential facts and interpretations of the occurrence and origin of these remarkable 

deposits (Tsikos and Moore, 1997; Tsikos and others, 2003). 

In the Kalahari, three manganese layers are within the banded 

iron formation known as the Hotazel iron formation. The lower 

of the three is the principal source of ore and varies from 

15 to 45 meters (m) thick. The hallmark of such deposits is the 

interlayering of iron- and manganese-rich layers, which indicate that the deposit formed from an ocean in which both iron 

and manganese were enriched, but that the two metals were 

differentiated from each other at the site of deposition, perhaps 

by subtle changes in the oxidation state or acidity of the ocean 

water. In the Kalahari, for instance, the manganese layers have 

iron contents that are near the average crustal abundance but 

are enriched in manganese from 200 to nearly 400 times the 

average crustal abundance (Beukes and Gutzmer, 1996). The 

iron-rich layers, on the other hand, are enriched above the 

average crustal abundance by as much as fivefold in iron and 

from twofold to twentyfold in manganese.

The mutual enrichment of both iron and manganese, 

although in varying proportions between iron and manganese 

layers, seems inconsistent with the sulfidic ocean model 

in which manganese is enriched to the near exclusion of 

iron. Rather, a stratified ocean in which deeper waters are 

suboxic to anoxic, but not sulfidic, is more consistent with 

the origin of these deposits. In anoxic water, the solubility of 

both manganese and iron is enhanced, and such waters are 

capable of delivering those metals to sites of deposition by 

such processes as coastal upwelling onto continental margins, 

which is the mode of deposition favored by nearly all recent 

interpretations of these deposits. In the Kalahari, although 

the primary sedimentary accumulations of manganese are, 

in part, ore-grade material, secondary enrichment by later 

hydrothermal processes has increased manganese grades 

in some areas to form especially desirable ores (Tsikos and 

others, 2003).

The Urucum deposit in Brazil may have a somewhat 

different origin. The manganese beds are interlayered with 

hematitic jasper, similar to the Kalahari and other deposits, but 

the enclosing rocks bear strong evidence of glacial conditions 

when this deposit formed. Floating glacial ice or a persistent 

ice pack has been proposed, which cut off the atmospheric 

oxygen supply to the ocean water. The decay of organic matter 

consumed available oxygen and resulted in the anoxic conditions, which, in turn, allowed high solubility of manganese 

and iron that resulted in the formation of banded iron and 

manganese deposits (Urban and others, 1992). Manganese 

grades range from about 27 to 45 percent, and associated 

iron ranges from 13 to 30 percent, which indicates that 

both iron and manganese were enriched in seawater during 

manganese deposition.

Secondary Enrichment (Supergene) Deposits

Many of the currently mined manganese ores have 

been enriched in manganese by processes at the present land 

surface, although they commonly formed in rocks that were 

already manganese rich. These secondary enrichment types 

of deposits, or supergene deposits, form where chemical 

reactions taking place within tens of meters of the surface 

redistribute manganese at a local scale and also leach out 

nonmanganese components, which results in residual enrichment of the manganese. Details of many of these deposits 

and their origin are presented in Varentsov (1996). Significant 

occurrences of these types of deposits are the Moanda deposit 

in Gabon, the Azul and the Serra do Navio deposits in Brazil, 

and numerous deposits in India. By their nature, the resources 

of these deposits are smaller than are those of the major 

sedimentary-type deposits, but the deposits are still substantial 

in terms of both their resources and current production. These 

deposits include some of the highest grade ores presently 

being mined. The process of supergene manganese enrichment 

is described in general terms by Lelong and others (1976), and 

details of the supergene enrichment at the Moanda deposit, 

which is the largest known deposit of this type, is outlined by 

Nahon and Parc (1990).

Supergene deposits are most abundant and commonly of 

the highest grade in tropical regions where humid conditions 

and abundant vegetation are key factors in forming deep 

tropical soils. Supergene manganese ores can be thought of 

as a specialized product of such tropical soil formation. The 

generation of humic acids by decomposition of abundant 

organic matter results in the dissolution of manganese-bearing 

minerals and transportation of manganese downward in 

the soil profile. It is then reprecipitated where soil water 

encounters less acidic conditions. The same acidic soil 

solutions dissolve and remove other components as well, 

resulting in a residual enrichment of manganese minerals. 

Supergene manganese deposits tend to have developed most 

frequently in areas of low topographic relief, commonly 

on plateaus, where slow erosion allows time for these oreforming processes to take place.

Seabed Deposits

The world’s ocean floors include tens of millions of 

square kilometers (km2) covered with manganese-rich deposits 

of ferromanganese nodules and crusts. No manganese or other 

metals have been produced from any of these deposits in spite 

of several decades of research and exploration. Numerous 

exploration and development programs are underway that 

could bring part of this vast resource online in the near future. 
L10 Critical Mineral Resources of the United States—Manganese

If this possibility materializes, seabed deposits could have 

a significant effect on the global production and supply of 

manganese and lead to a diversification of supply and a 

reduction in U.S. import reliance for manganese. The areas of 

greatest interest and potential for mining of seabed deposits 

are shown in figure L3.

Ferromanganese Nodules

The widespread occurrence of seabed nodules composed 

mostly of manganese and iron oxide and hydroxide minerals 

has been known since the research expedition of the 

H.M.S. Challenger in the 1870s; these nodules have come 

to be referred to as ferromanganese nodules. The locations 

of these nodules are shown in figure L3. In the area of prime 

economic interest in the Pacific Ocean, nodules range in size 

from 1 to 12 centimeters (cm) and median dimensions are 

8 cm, 6 cm, and 4 cm for the maximum, intermediate, and 

minimum dimensions, respectively (fig. L6; Morgan, 2000). 

The abundance of nodules within known nodule fields ranges 

from sparse to a nearly continuous blanket covering the seabed. 

These nodules also contain substantial amounts of other valuable metals, such as nickel, copper, and cobalt, and potentially 

valuable concentrations of rare-earth elements (REEs). The 

amount of manganese in known nodule fields dwarfs that 

found in traditional continental deposits, but its availability as 

a source of ore is uncertain in the near term. Considerable technological issues of mining at abyssal depths in the oceans, the 

economic competitiveness of seabed mining versus traditional 

mining, and legal issues of ownership and control of resources 

in international waters are still being addressed.

Recent summaries of the origin, distribution, and 

potential value of manganese nodules are provided by Hein 

(2012), and the International Seabed Authority (ISA) (2010), 

both of which form the basis of the following discussion. 

Manganese nodules occur widely in global oceans and their 

distribution has commonly been mapped, for both scientific 

and potential commercial purposes. Vast areas of ocean 

floor contain blankets of nodules in varying abundance at 

the sediment surface, and an equal amount of nodules is 

believed to be buried within the uppermost meter or two of 

underlying sediment. Nodule fields range in depth below the 

ocean surface of from 3,500 to 6,500 m. The nodule fields 

of the greatest current commercial interest are those located 

in the northern equatorial Pacific Ocean and the central 

Indian Ocean.

Ferromanganese nodules are geologically young features 

and many continue to grow today. They form on the seabed 

at the sediment-water interface, mostly by precipitation of 

metals that are mobilized by redox reactions in the underlying 

sediments. An expert panel convened by the ISA (an organization established by the 1982 United Nations Convention on 

the Law of the Sea) developed a biogeochemical model for 

nodule formation in the northern equatorial Pacific Ocean, 

generally referred to as the Clarion-Clipperton zone (CCZ), 

and subsequently used the model to make resource estimates. 

According to the ISA model, the principal source of copper, 

manganese, and nickel is believed to be continental runoff 

from North America; volcanogenic sources from the East 

Pacific rise, including submarine hydrothermal vents; and 

atmospheric dust. The metals from continental sources occur 

in dissolved form and are also adsorbed onto fine-grained sediment particles that are carried westward by the North Pacific 

current. Some metals can be delivered to the sea floor through 

slow settling of these particles, but biological activity is also 

an essential part of the deposition. In the near-surface photic 

zone, metals are scavenged by planktonic organisms and can 

then be delivered to the sea floor by settling of those organisms after their death and also by incorporating metals into 

fecal pellets, which likewise settle to the sea floor. Thus, areas 

of highest metal enrichment on the sea floor correspond to 

areas of high surface biological productivity, such as the CCZ.

Once deposited in sea-floor sediments, metals are available to be metabolized by benthic organisms and mobilized 

into dissolved form by reducing reactions within the upper 

meter of sediments. Metal-rich pore waters so produced are the 

source fluids for most components of nodules. Metals dissolved 

Figure L6. Photograph showing a dense carpet of ferromanganese nodules on the seabed off Johnston Island within 

the United States Exclusive Economic Zone near Hawaii. 

The field of view is about 4 meters square. The inset shows 

individual nodules, each of which is about 3 centimeters in 

diameter, from the equatorial Pacific Ocean. Photographs

by James Hein, U.S. Geological Survey.
Geology L11

in seawater may add additional components. As pore waters 

migrate to the sediment surface, they reenter oxidizing conditions where manganese is precipitated as oxide and hydroxide 

minerals that effectively scavenge other metals, such as nickel 

and copper. The precipitated minerals commonly nucleate 

around small solid particles and sequentially add growth layers 

to form the present nodules.

Ferromanganese Crusts

Ferromanganese crusts are layers of manganese and 

iron oxides and hydroxides, ranging from a thin veneer to 

260 millimeters (mm) thick, that have precipitated on hard 

rock substrates on the seabed. The geology and resource 

status of ferromanganese crusts are outlined by Hein and 

others (2009, 2010, 2013). The principal commercial interest 

in ferromanganese crusts has been their concentrations of 

cobalt (0.3 to 0.67 percent) and nickel (0.23 to 0.46 percent). 

Elevated concentrations of REEs might also add to their 

potential economic value. Manganese grades range from 

20 to 25 percent in Pacific Ocean and Indian Ocean crusts 

and are somewhat lower elsewhere. A detailed summary of 

the chemical composition of ferromanganese crusts from the 

world’s oceans was compiled by Manheim and Lane-Bostwick 

(1988) and provides much more detail than this chapter. 

The locations of these crusts are shown in figure L3, and an 

example of such crustal material is shown in figure L7.

Ferromanganese crusts are less extensive than nodule 

fields and are restricted to seabed topographic highs, such as 

volcanic edifices, seamounts, and ridges. Crusts are found at 

depths of 400 to 7,000 m; the thickest and most metal-rich 

crusts are found at depths of between 800 and 2,500 m. 

Ferromanganese crusts form in areas that were kept sedimentfree for millions of years, principally by bottom currents 

with sufficient velocity to sweep the seabed free of sediment 

accumulations. Crusts grow extremely slowly by precipitating 

manganese and other elements that occur in low concentrations in seawater. It is this very slow growth that allows the 

ferromanganese minerals to scavenge economically interesting 

concentrations of metals.

Figure L7. Photograph showing ferromanganese crust (black material above) on carbonate rock from 

the Blake Plateau off the southeastern coast of the United States. The surface below the black material 

was the seabed on which the manganese minerals were precipitated. Photograph by William F. Cannon, 

U.S. Geological Survey.
L12 Critical Mineral Resources of the United States—Manganese

Resources and Production

Manganese resources are very large relative to current 

and anticipated demand on a global scale, but they are 

unevenly distributed. Many industrialized countries, including 

the United States, are largely or wholly reliant on imports 

for their manganese supply. The United States has relatively 

large tonnages of resources in several districts, but the larger 

deposits are low grade and well below economic viability in 

current economic and technological scenarios.

The nature of manganese deposits presents problems 

for precise quantitative resource estimates. Many deposits 

occur as laterally extensive manganese-rich layers that do 

not have sharp lateral boundaries, but rather, grade gradually 

from higher to lower grades, in some cases over long 

distances. Very different resource estimates can be made 

for the same deposit based on what criteria are used for 

acceptable manganese content to be classed as a resource, 

as well as assumptions of how grades change in poorly 

explored areas. Additional variability in published resource 

values is introduced by a variety of resource classification 

schemes and definitions of resource categories used by various 

Governments around the world. There is no simple way to 

convert resource estimates between such diverse classifications 

into a uniform resource compilation. Global resources of 

manganese were estimated by the USGS and published in 

1973 (Dorr and others, 1973). A more-detailed study was 

conducted by a panel of industry, Government, and academic 

experts through the U.S. National Materials Advisory Board 

(1981). The following year, a summary of global manganese 

resources was conducted through the International Strategic 

Minerals Inventory (ISMI), a cooperative effort of Federal 

geological and resource agencies from the United States, 

Australia, Canada, the Germany, South Africa, and the 

United Kingdom (DeYoung and others, 1984). Estimated 

resources by country and district differ substantially between 

the two older reports, in large part because of somewhat 

different definitions of resource and reserve categories used in 

them, as well as the criteria used to calculate resource figures. 

The ISMI study produced consensus estimates based both on 

older reports and resource values compiled by the participating 

Government agencies. For this summary, we have not revised 

older resource estimates although we recognize that depletion 

of resources through continued mining and addition to 

resources through continued exploration has modified older 

resource estimates to some degree. The discussion below is 

based on the ISMI resource values, although the same general 

picture of global resources could be formed from any of the 

three cited studies.

Manganese Reserves

The United States possesses no reserves of manganese 

ore containing 35 percent or more manganese or from which 

concentrates of such grade could be commercially produced. 

As a result, manganese ore is not mined in the United States, 

with the exception of ultra-low-grade manganese schists that 

contain less than 5 percent manganese at two mines in South 

Carolina—the Grover and the Martin Mines. Manganese 

recovered from these mines is used as a brick colorant. 

The estimated manganese content of world manganese ore 

reserves in 2012 totaled 630 million metric tons (table L1). 

The countries with the most manganese ore reserves are, 

in descending order of reserves, South Africa, Ukraine, 

and Brazil.

Other Identified Resources

Manganese resources can be divided into (a) land-based 

deposits and districts, which are the traditional sources of 

manganese ore, and (b) seabed resources, which are deposits 

on the ocean floor that have been identified and characterized 

to varying degrees of detail. Seabed resources are enormous 

and dwarf traditional resources, but their technological and 

economic availability is unproven.

Land-Based Resources

The locations of identified manganese resources in 

conventional land-based deposits and mining districts are 

shown in figure L3. Some details of the resource amounts are 

in table L2.

Total global identified resources are in excess of 

17 billion metric tons of manganiferous material. Global 

resources are dominated by the Kalahari manganese district 

of South Africa, which contains more than 70 percent of 

global resources (fig. L3). Kalahari ores are among the 

highest grade, typically about 40 percent manganese, and 

are widely used in industrial processes in many parts of 

the world. The Kalahari district clearly has the potential to 

provide international markets with high-quality manganese 

ore for a very long time. One of the larger mines in the 

Kalahari district, the Mamatwan Mine, is shown in figure L8. 

The second-ranked district in terms of total resources is the 

Molango district in Mexico, which has about 9 percent of the 

world’s total resources. The resource consists of an extensive 

layer of manganese carbonate, with grades of approximately 

28 percent manganese. Although this resource is large, 

economic extraction is impaired by the steep mountainous 

terrain of the district, which requires that most, or all, of 

the mining be conducted from underground operations 

rather than more efficient surface mines. The carbonate 

character of the ore also is a detriment in that it requires an 

additional processing step to convert the carbonate minerals 

to manganese oxide. Molango is an important source of 

manganese in Mexico, but its future significance in global 

manganese supply is uncertain. The third largest resource 

is in the Bolshe Tokmak mining district in Ukraine, which 

constitutes about 6 percent of global resources. The Bolshe 

Tokmak ore is low grade by global standards, averaging about 
Resources and Production L13

Table L2. Identified resources estimated for major land-based manganese deposits of the world.

[Data are from DeYoung and others (1984). WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the 

deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. Deposits 

are shown on the map in figure L3. Resource values are the sum of all manganese resource categories. —, not applicable]

Country Deposit

(fig. L3)

Latitude Longitude

Type

Resource

(thousand

metric tons) Decimal degrees WGS 84

Australia Groote Eylandt –14.00 136.50 Marine sedimentary 436,000

Brazil Azul – 6.13 –50.35 Secondary enrichment 65,000

Brazil Morro da Mina –20.58 –43.83 Secondary enrichment 4,500

Brazil Serra do Navio 1.00 –52.00 Secondary enrichment 21,000

Brazil Urucum –19.13 –57.55 Marine sedimentary 59,000

Bulgaria Varna 43.22 27.92 Marine sedimentary 28,000

Burkina Faso Tambao 14.78 –0.07 Secondary enrichment 17,000

China Hsiang-tan 27.83 112.92 Marine sedimentary >10,000

China Leiping 23.33 109.50 Marine sedimentary >10,000

China Tsun-i 27.50 106.50 Marine sedimentary >10,000

China Wafangtzu 39.58 122.00 Marine sedimentary 4,000

Gabon Moanda 1.57 13.28 Secondary enrichment 220,000

Georgia Tchiatura 42.32 43.30 Marine sedimentary 155,000

Ghana Nsuta 5.28 –1.95 Secondary enrichment 20,000

Hungary Urkut 47.08 17.67 Marine sedimentary 215,000

India Andhra Pradesh 18.37 83.45 Secondary enrichment 1,700

India Goa 15.50 73.92 Secondary enrichment 8,550

India Gujarat 23.00 72.50 Marine sedimentary 2,900

India Karnataka 13.93 75.52 Marine sedimentary 15,000

India Madhya Pradesh 21.83 80.30 Marine sedimentary 49,000

India Odisha 21.63 85.67 Secondary enrichment 43,600

Mexico Molango 20.83 98.73 Marine sedimentary 1,526,000

Morocco Imini 30.73 6.90 Marine sedimentary 1,500

South Africa Kalahari –27.18 22.95 Marine sedimentary 12,565,800

South Africa Postmasburg –28.05 23.05 Secondary enrichment 15,000

South Africa Western Transvaal –26.25 27.02 Secondary enrichment 15,000

Ukraine Bolshe Tokmak 47.25 35.70 Marine sedimentary 1,100,000

Ukraine Nikopol 47.57 34.42 Marine sedimentary 655,000

World total — — — — >17,273,000

Figure L8. Photograph of the Mamatwan 

open pit mine in South Africa. The black 

manganese layer is about 45 meters thick 

and consists of a 20-meter-thick ore bed 

and two lower grade subeconomic beds. 

The slightly folded layering is shown by the 

shapes of the various mine benches. Lighter 

material in the background is waste that 

must be stripped to expose the ore bed; this 

waste material is used as backfill in mined 

areas (foreground). Photograph by 

William F. Cannon, U.S. Geological Survey.
L14 Critical Mineral Resources of the United States—Manganese

18 percent manganese. The mines at Bolshe Tokmak were 

significant sources of manganese to the Soviet Union, but 

their importance in market economies is impaired by their 

low grade. The Obrochishte deposit in the Varna district in 

Bulgaria, the Nikopol mining district in Ukraine, and the 

Tchiatura (also known as Chiat’ura) district in Georgia are 

similar to the Bolshe Tokmak district and together constitute 

an additional large but low-grade resource. All the deposits in 

these districts are of the marine sedimentary type.

A compilation of tonnage and grade for sedimentary-type 

deposits (Mosier, 1986) shows that their median tonnage is 

7.3 million metric tons and that the largest 10 percent of the 

deposits (made up of three deposits) have a median tonnage 

of greater than 280 million metric tons (fig. L9A). The median 

grade is 31 percent Mn and the grade of the richest 10 percent 

of deposits is greater than 49 percent Mn (fig. L9B).

Although they constitute a relatively small percentage 

of total global manganese resources, other types of deposits 

are also important sources of high-grade manganese ore to the 

United States and other countries. An example is Moanda in 

Gabon, which was formed from secondary enrichment.

China has become a major manganese ore producer in 

recent decades. In 2011, it was the second-ranked producer in 

the world in terms of gross weight of ore and the third-ranked 

producer in terms of contained manganese. China’s reserves 

are also large, consisting of about 44 million metric tons of 

manganese ore distributed over many deposits, according 

to figures compiled by the USGS (Corathers, 2013b). Most 

Chinese deposits consist of sedimentary accumulations of 

manganese carbonate that average about 20 percent contained 

manganese and thus are of relatively low grade compared 

with many other deposits globally. Most ores are upgraded, 

many by magnetic separation of manganese-rich material from 

waste material, before use in manufacturing (International 

Manganese Institute, 2012). The total manganese resource 

in China is poorly known. The ISMI summary prepared by 

DeYoung and others (1984) lists China’s proven resource at 

greater than 34 million metric tons, but the total resource, 

although not quantified, is likely much larger.

Seabed Resources

A vast resource of manganese lies on the seabed, mostly 

in international waters and at great depths. A strong interest 

in exploitation of these deposits in the late 1970s and early 

1980s, with hopes of developing mines by the late 1980s, 

vanished in a morass of technological, economic (low metal 

prices), and legal issues. More recently, with the establishment 

of a legal framework for exploring and mining in international 

waters, and a rebound in metal prices, a strong interest in 

exploiting these deposits has reemerged. Nevertheless, the 

technological and economic viability of doing so had yet 

15 31 49

0 10 20 30 40 50 60 70 80 90 100

Manganese grade, in percent

Proportion of deposits

Million metric tons

Manganese Fig 9

A B

1.0

0.9

0.8

0.7

0.6

0.5

0.4

0.3

0.2

0.1

0

0.0 0.025 0.1 0.4 1.6 6.3 25 100 400 1,600 6,300

0.19 7.3 280

1.0

0.9

0.8

0.7

0.6

0.5

0.4

0.3

0.2

0.1

0

n=39 n=39

Figure L9. Graphs showing the cumulative frequency of tonnages and grades of 39 marine sedimentary manganese 

deposits. A, The median tonnage of the deposits is 7.3 million metric tons, and the largest 10 percent of the deposits 

has a tonnage of greater than 280 million metric tons. B, The median grade is 31 percent manganese, and the richest 

10 percent of the deposits has a grade of greater than 49 percent manganese (Mosier, 1986).
Resources and Production L15

to be proven in 2013. Importantly, the economic interest 

in ferromanganese nodules is driven by such metals as, in 

order of economic importance, nickel, copper, and cobalt, 

rather than by manganese. Manganese would, at best, be a 

byproduct of seabed mining, and perhaps not a metal that 

would be economically recovered at all in some scenarios. 

Consequently, the availability of manganese from seabed 

deposits will be closely tied to the demand for those other 

metals. Two types of deposits of current interest—ferromanganese nodules and ferromanganese crusts—are similar in 

that they are composed largely of manganese and iron oxides 

and hydroxides, but they differ substantially in their mode 

of origin, chemical composition, location on the seabed, and 

economic potential. They are discussed separately below.

Ferromanganese Nodules

The possible commercial recovery of ferromanganese 

nodules, which were of little economic interest for several 

decades after initial investigations and conditions showed 

them unfavorable, has received significantly increased 

activity in recent years because of higher metal prices 

and the establishment of a legal framework for exploiting 

resources on the international seabed through the ISA. By 

2013, 13 organizations or Governments had signed contracts 

with the ISA to explore and potentially mine ferromanganese 

nodules, mostly in the CCZ. One of the companies that signed 

a contract with the ISA is a U.S. company, Lockheed Martin 

Corp., through its British subsidiary Lockheed Martin UK, 

which is evaluating the feasibility of nodule mining in two 

areas within the CCZ (Schrope, 2013).

The ISA assessed mineral resources within the CCZ using 

a biogeochemical model for nodule formation and several 

statistical methods (International Seabed Authority, 2010, 

2012). The various methods produced somewhat different 

results that ranged from 20 to 30 billion metric tons of nodules 

containing from 6 to 8.7 billion metric tons of manganese at 

a median grade of about 30 percent manganese. The resource 

covers an area of seabed of roughly 4 million km2. The central 

Indian Ocean, which is a 150,000-km2

 area in which exclusive 

exploration rights are held by India, is estimated to contain 

a nodule resource of 700 million metric tons containing 

20 to 30 percent manganese (Banakar, 2010). No comparable 

estimates have been done for other nodule fields, but it is 

reasonable to assume that their combined resource would be 

several times that of the CCZ and the central Indian Ocean.

As previously mentioned, an important factor in the 

potential of ferromanganese nodules as a manganese source 

is that most economic interest in nodules is driven by metals 

other than manganese, including cobalt, copper, nickel, 

possibly REEs, and other metals needed for high-technology 

applications. (See discussion in Hein and others [2013].) 

Without the value provided by the presence of these metals, 

it is unlikely that nodules would be an economically viable 

source of manganese based solely on their contained manganese value. It is not clear to what extent future seabed mining 

and processing would recover manganese rather than dispose 

of it as a component of limited economic value if manganese 

prices remain relatively low. In any case, the processing of 

manganese nodules would at least result in an inventory of 

manganese-rich material at processing facilities that could be 

accessed when economically or strategically appropriate.

Ferromanganese Crusts

The area of greatest interest for ferromanganese crusts 

in the north Pacific was outlined by Hein and others (2009) 

and is referred to as the Pacific prime crust zone (fig. L3). 

This large area contains 155 volcanic features that could host 

a ferromanganese crust resource. In 2013, three nations (China, 

Japan, and Russia) had contracts with the ISA to explore tracts 

totaling 9,000 km2. A model based on geologic parameters (and 

not on economic considerations) predicted that about eighteen 

100-km2 blocks of seabed might be included in an initial mining 

program under the auspices of the ISA (Hein and others, 2009).

An additional area in the northeast Atlantic Ocean that 

lies partly within the Portuguese Exclusive Economic Zone 

(EEZ) has been described by Muiños and others (2013). 

Crusts with resources of cobalt, copper, nickel, and other trace 

elements that may be comparable to those in the central Pacific 

region have been identified on 10 seamounts. The crusts 

average about 15 percent manganese, so manganese would be 

a byproduct, at best, of mining of these deposits.

Many other areas with potential for ferromanganese 

crusts are known and many lie within the EEZs of various 

nations rather than in international waters (fig. L3). Areas 

within the United States EEZ lie off both the Atlantic and 

Pacific Coasts. The area of greatest previous economic 

interest is the Blake Plateau off the southeastern coast of the 

United States. Geologic knowledge of this area is summarized 

by Manheim and others (1982). Median manganese grades are 

about 20 percent. Median cobalt and nickel grades are roughly 

0.5 percent and 0.4 percent, respectively. An estimate of the 

total tonnage of crusts in the area has not been made.

The future of mining of seabed ferromanganese crusts is 

more uncertain than for nodules. Whereas nodules are discrete 

masses lying unattached to the seabed, ready for the taking, 

crusts are firmly attached to bedrock substrate. Successful 

mining requires development of methods and equipment that 

can effectively break crusts free from their substrate with 

minimal incorporation of substrate into the recovered material 

to minimize dilution. As with nodules, the metals of principal 

economic interest do not include manganese, but, rather, are 

cobalt and nickel. Thus, future production of crusts will be 

driven by consumption needs and prices for those metals 

rather than by demand for manganese.
L16 Critical Mineral Resources of the United States—Manganese

Resources in the United States

Land-Based Resources

Manganese ore was last produced in the United States 

in 1973. Significant past production has been limited to times 

of war or Government subsidies and purchases for strategic 

stockpiles. Several deposits and districts in the United States 

have traditionally been included in identified resources, 

although all are well below manganese grades of material 

readily available on international markets. All would require 

substantial beneficiation to make usable manganese products 

or discovery of higher grade zones within them, which would 

likely be small relative to national consumption. Table L3 lists 

these domestic resources and some of their characteristics 

based on a manganese availability study conducted by the 

U.S. Bureau of Mines (Kilgore and Thomas, 1982). In that 

study, economic and technological parameters were applied 

to each deposit, and cost estimates were made for producing 

a usable manganese product. All deposits were shown to be 

far short of economic viability. Of these identified resources, 

those of Aroostook County, Maine (Maple Mountain-Hovey 

Mountain area and North Aroostook district) and the Cuyuna 

Range in Minnesota (fig. L3) are the only two that are of sufficient tonnage to be significant on a national scale of manganese 

demand. Others may be capable of providing some smaller 

amount of manganese or specialized manganese products, but 

the discussion below is limited to these two largest resources.

Aroostook County, Maine

The manganese deposits of Aroostook County lie in a 

discontinuous north-south belt about 50 kilometers (km) long. 

The bulk of the manganese resource lies in the middle part of 

this belt in the Maple Mountain-Hovey Mountain area. These 

deposits and their geologic setting were comprehensively 

studied in the early 1960s (Pavlides, 1962). The manganeserich rocks occur in three beds of Silurian age and are of two 

lithologic types: (a) hematitic shale and slate and hematitic 

ironstone, and (b) siliceous carbonate rocks. These iron-rich 

layers contain manganese in both braunite and rhodochrosite. 

Manganese grade varies between beds and ranges from 

7 to 11 percent. Iron grades range from 13 to 24 percent in the 

same beds. The grades are relatively uniform within individual 

beds. Individual manganese beds are as much as about 20 m 

thick. The deposits, although now folded and recrystallized, 

formed as discontinuous marine sedimentary layers in shallow, 

nearshore settings. The bulk of the resource as estimated by 

Pavlides (1962) is the hematitic variety that contains about 

300 million metric tons of material averaging 7 percent 

manganese and 17 percent iron. The carbonate variety is much 

smaller and contains only 3 million metric tons of material. 

These figures were revised downward by a more recent study 

(Kilgore and Thomas, 1982), and are shown in table L3. 

Deposits farther north in the North Aroostook district are 

similar to those of the Maple Mountain-Hovey Mountain area 

but are significantly smaller. Together, those northern deposits 

contain an additional 60 million metric tons.

Although the deposits in Aroostook County contain 

sufficient manganese to satisfy U.S. consumption for several 

decades, the production of a manganese product that would 

be viable in current markets would require substantial beneficiation of these low-grade resources. An economic analysis 

of the cost of mining, beneficiation, and transport showed 

that these deposits are far from being economically viable 

(Kilgore and Thomas, 1982). Although that analysis was done 

30 years ago, more recent developments have not substantially 

increased the economic value of these deposits. Thus, they 

remain a relatively large but subeconomic resource that can 

Table L3. Grade, tonnage, and quantity of contained manganese for eight manganese deposits and districts 

in the United States.

[Data are as reported in Kilgore and Thomas (1982). Numbers for contained manganese (Mn) have been rounded to three significant figures. %, percent; —, not applicable]

Deposit/district State Grade

(% Mn)

Resource

(metric tons)

Contained Mn

(thousand 

metric tons)

Hardshell Mine Arizona 15.0 5,895,500 884

Maggie Mine Arizona 8.75 8,441,000 739

Sunnyside Mine Colorado 10.0 24,909,000 2,490

Maple Mountain-Hovey Mountain Maine 8.87 260,000,000 23,100

North Aroostook district Maine 9.54 63,100,000 6,110

Cuyuna Range (southwest portion) Minnesota 7.84 48,960,000 3,840

Butte district (Emma Mine) Montana 18.0 1,232,000 222

Three Kids Mine Nevada 13.2 7,230,000 954

Total — — 419,767,500 38,300
Resources and Production L17

become a significant manganese supplier for the United States 

only with major advancements in processing and beneficiation 

technology or long-term disruption of international 

manganese trade.

Cuyuna Range, Minnesota

The Cuyuna Range is one of several iron ranges of the 

Lake Superior region that have supplied a large proportion 

of the iron needed to meet U.S. iron consumption. The 

deposits include Precambrian banded iron formations and 

secondary iron concentrations within them. The Cuyuna 

Range, which is now inactive as an iron producer, is distinct 

from other ranges in its relatively high content of manganese. 

The range produced substantial quantities of manganiferous 

iron ores, as distinct from manganese ore, for more than 

70 years. Production ceased in 1984, at which time more 

than 100 million metric tons of ore had been produced, 

largely from open pit operations. The produced ores averaged 

between 13 percent and 14 percent manganese. A relatively 

recent summary of the geology and manganese resources of 

the range was provided by Morey (1990) from which most of 

the following summary is extracted.

The Cuyuna Range ores are supergene concentrations 

of iron and manganese that formed in Precambrian strata 

that are themselves enriched in both iron and manganese. 

The original strata contain 20 to 35 percent iron and from 

less than 1 percent to 16 percent manganese. The supergene 

enrichments that constitute the manganiferous iron ores range 

from 30 to 40 percent iron and 5 to 15 percent manganese. 

Several estimates of manganese resources have been made 

and summarized by Morey (1990). An estimate of measured 

resources at 39 existing (abandoned) mines is 5.5 million 

metric tons that average about 11.7 percent manganese 

(Beltrame and others, 1981). A much larger tonnage of 

manganiferous material has been estimated within the ferruginous strata of the district. A resource of 80 million metric 

tons averaging about 10.5 percent manganese and within 120 m 

of the surface was estimated by Beltrame and others (1981). 

The same authors estimated that an additional inferred resource 

of 90 million metric tons within 120 m of the surface exists 

within less well characterized parts of the district. In 1982, a 

summary of U.S. resources (Kilgore and Thomas, 1982) estimated a resource of about 49 million metric tons with a grade 

of 7.84 percent manganese. These are included in table L3.

The production of a commercial manganese product 

from Cuyuna Range manganiferous iron ores requires either 

physical processing and concentration of manganese minerals 

or hydrometallurgical concentration. Because a great majority 

of the manganese minerals are fine-grained and intergrown 

with other minerals, physical separation and beneficiation 

would be difficult. Some renewed interest in manganese 

production is indicated by pilot-scale testing of an innovative hydraulic mining method in which high-pressure jets 

of water are used to mobilize these physically soft ores into 

slurry within large-diameter drill holes (Crow Wing Power, 

2013). The economic feasibility of this method was still being 

evaluated in 2013.

Seabed Resources

A large but incompletely characterized and quantified 

manganese resource lies within the vast area of the 

United States EEZ. Extensive areas of both ferromanganese 

crusts and, to a lesser extent, nodules, are known offshore 

of both the Atlantic Coast and the Pacific Coast of the 

U.S. mainland, including Alaska, as well as the Hawaiian 

Islands, and other island possessions. Most economic interest 

and exploration for seabed deposits has been focused on the 

prime nodule zone of the CCZ in international water where 

seabed mining of nodules is most likely to take place first. 

The United States EEZ, however, holds some promise to 

provide manganese and related metals in the future if the 

technology and economics of seabed mining lead to viable 

mining operations in the areas of current highest interest. 

Because most deposits within the EEZ are crusts, which are 

considerably more challenging to mine than nodules, the 

potential mining of these deposits may lag development of 

nodule mining operations. So, although the EEZ undoubtedly 

contains a very large tonnage of manganese resources, it is 

unlikely to have an effect on domestic and global manganese 

production and consumption in the short term.

Undiscovered Resources

Marine sedimentary manganese deposits of economic 

significance are large geologic features. They generally have 

been found by accidental discovery in surface rock exposures 

rather than by purposeful exploration or exploration for 

subsurface deposits. It has been many decades since the last 

such major discovery in spite of an increasing pace of geologic 

studies and mineral exploration. Currently, exploration for 

new deposits is mostly confined to known districts rather than 

exploration intended to find wholly new districts. With time, 

the likelihood of new discoveries of major deposits is diminished. That said, there is still some possibility of important 

new discoveries, especially in light of the rather unremarkable 

outward appearance of some manganiferous sediments, 

especially manganese carbonate deposits. In some cases, the 

manganiferous character of these rocks was realized only after 

discovery of supergene deposits formed from them.

Supergene deposits offer a greater potential for undiscovered deposits because of their smaller size and propensity to 

occur in tropical regions with poor bedrock exposure, limited 

access, and incomplete geologic studies. Exploration focused 

in suitable areas with known manganiferous sediments is 

likely to result in new discoveries.
L18 Critical Mineral Resources of the United States—Manganese

Exploration for New Deposits

There are currently 10 new land-based manganese 

mines under development in various parts of the world that 

are scheduled for completion by the end of 2014. The total 

annual production capacity of new production will be nearly 

5.42 million metric tons of ore, more than one-half of which 

will be from the Kalahari manganese district in South Africa 

(Corathers, 2013a). These new mines attest to the continued 

need for manganese exploration throughout the world. Most, 

if not all, of these new mines are in areas of previously 

known manganese mineralization and do not represent 

newly discovered manganese districts.

The exact methods of exploration and discovery are not 

known for all of these deposits, but likely included surface 

geologic studies and exploration drilling. In some cases, 

geophysical methods, particularly electromagnetic surveys, 

aided in defining targets for drilling.

It is reasonable to assume that, for the near future, 

land-based exploration will continue to focus on areas where 

there are known manganiferous rocks and will be designed to 

outline areas of sufficiently high grade and tonnage to constitute an economic orebody. More-wide-ranging exploration 

aimed at discovery of new frontier districts or new manganiferous sedimentary units may also take place in areas where 

geologic conditions suggest favorable settings for manganese 

deposits to have formed. Significant sedimentary manganese 

deposits cover large areas and are at least partly exposed at 

the surface. As such, they represent large exploration targets. 

Remote sensing methods, including satellite imagery, airborne 

geophysical surveys, and hyperspectral mapping, may be effective exploration tools. Satellite imagery may help identify and 

refine favorable exploration tracts and outcropping manganiferous rocks. Airborne geophysical surveys, especially electromagnetic surveys, may be useful in mapping shallowly buried 

layers rich in manganese because of the electrically conductive 

nature of manganese oxides. Hyperspectral mapping has the 

potential to specifically identify manganese oxide minerals 

because they have a low reflectance signature relative to 

other common minerals. Once a manganese deposit has been 

identified by these methods, more-detailed ground surveys are 

needed to refine the size and grade of deposits to evaluate their 

economic viability fully. Early stages of such exploration are 

commonly conducted at least partly confidentially by exploration and mining companies so that the degree to which such 

exploration might currently be taking place is not well known.

Exploration of the seabed is also active; 13 individual 

entities were exploring for ferromanganese nodules in 2013. 

Identification of nodule-rich seabed regions is relatively 

simple and can be done by remote visual examination to 

identify areas with the greatest abundance of nodules. Physical 

sampling of nodules is also required, followed by chemical 

assays to map the distribution of nodules with the most favorable grades of manganese and other economically important 

metals, such as cobalt, copper, and nickel. Exploration for 

ferromanganese crusts is much less advanced, although 

in 2013, the ISA received two plans for exploration in the 

western Pacific Ocean.

Environmental Considerations

Manganese is ubiquitous on Earth’s surface. Humans 

and virtually all other life forms interact with it every day. 

Manganese is an essential nutrient for most life, including 

humans, but it can produce a variety of toxic effects at 

anomalously high concentrations. As such, the potential for 

manganese mining and processing to produce such elevated 

concentrations must be addressed throughout the life cycle 

of manganese production and use. Manganese deposits 

and the types of operations that mine them are very diverse 

around the globe. Some examples of possible concerns 

regarding increased environmental manganese around mines 

and processing plants are discussed below, along with some 

general aspects of manganese geochemistry that could make it 

possible to predict potential environmental and health impacts 

of manganese mining.

Sources and Fate in the Environment

Examples of natural concentrations of manganese in 

rocks, soil, water, and air are given in table L4. Manganese 

contents in soils range from 7 to 9,200 milligrams per kilogram (mg/kg) (or parts per million), with higher manganese 

concentrations commonly occurring in soils developed from 

mafic rocks, soils rich in iron and (or) organic matter, and 

soils developed in arid or semiarid regions (Kabata-Pendias 

and Mukherjee, 2007). Dissolved manganese concentrations 

in large rivers around the world range from 0.4 to 113 μg/L 

(or parts per billion) (Gaillardet and others, 2003, and references

therein), whereas the average concentration of manganese in 

the suspended sediment of world rivers has been reported as 

1,679 mg/kg (Viers and others, 2009). Occasionally, dissolved 

manganese concentrations can be comparable to particulate 

manganese concentrations. For example, during flood events 

in Magela Creek in the Northern Territory, Australia, dissolved 

(<0.02 micrometer [μm]-size) manganese concentrations 

(0.8 to 9.6 μg/L) are similar to particulate (>0.02 μm-size) 

manganese concentrations (0.3 to 7.9 μg/L); the relatively 

high dissolved manganese concentrations are attributed to the 

acidic soils, which create favorable conditions for manganese 

dissolution from minerals (Hart and others, 1992). Manganese 

occurs naturally in the atmosphere as part of mineral dust 

particles. Nriagu (1989) estimated manganese emission to the 

atmosphere from natural sources to be 317 million kilograms 

per year (Mkg/yr). Air manganese concentrations in remote 

regions range from 0.004 to 0.02 nanograms per cubic meter 

(ng/m3

) over the South Pole and from 2.8 to 4.5 ng/m3 over 

Greenland (Kabata-Pendias and Pendias, 2001).
Environmental Considerations L19

Table L4. Background and above-background concentrations of manganese in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency. Units of measure: cm, centimeter; mg/kg, milligram per kilogram (or part 

per million); mg/L, milligram per liter; mg/m3, milligram per cubic meter; ng/m3, nanogram per cubic meter; μg/L, microgram per liter; μg/m3, microgram per 

cubic meter; μm, micrometer]

Environment

and (or) location

Manganese 

concentration Unit Notes Reference(s)

Rocks

Upper continental crust 600 mg/kg Average Taylor and McLennan (1995)

Bulk continental crust 1,400 mg/kg Average Taylor and McLennan (1995)

Lower continental crust 1,700 mg/kg Average Taylor and McLennan (1995)

Soils

Western United States 380 mg/kg Mean for 20 cm depth Shacklette and Boerngen (1984)

Eastern United States 260 mg/kg Mean for 20 cm depth Shacklette and Boerngen (1984)

Central Pennsylvania, 

United States

900 to 14,400 mg/kg Surface soil in an undisturbed 

watershed

Herndon and others (2011)

Central Pennsylvania, 

United States

340 to 1,300 mg/kg Surface soil proximal to historic 

steel manufacturing site

Herndon and others (2011)

Poland 143±5 mg/kg Agricultural soil; nonpolluted Połedniok and others (2012)

Poland 658±7 mg/kg Agricultural soil; nonpolluted Połedniok and others (2012)

Poland 130±4 mg/kg Agricultural soil; nonpolluted Połedniok and others (2012)

Poland 292 to 571 mg/kg Proximal to mining and 

industrial sources

Połedniok and others (2012)

Sweden 76 to 206 mg/kg Range of profile developed on 

quartzite and gneiss

Tyler (2004)

Proposed DOE benchmark 

(recommended

upper limit in soil)

500 mg/kg Contaminant screening benchmark 

for terrestrial plants

Efroymson and others (1997)

Waters

Bathurst Harbour, Australia 0.6 to 2 μg/L Dissolved and colloidal (<0.45 μm); 

nonpolluted

Mackey and others (1996)

Macquarie Harbour, Australia 1 to 10 μg/L Dissolved and colloidal (<0.45 μm); 

polluted

Mackey and others (1996)

African rivers 0.4 to 30 μg/L Dissolved load (<0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

Asian rivers 0.5 to 1 μg/L Dissolved load (<0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

European rivers 3.8 to 48 μg/L Dissolved load (<0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

North American rivers 0.4 to 15 μg/L Dissolved load (<0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

South American rivers 0.46 to 113 μg/L Dissolved load (<0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

River water, Republic 

of Georgia

0.52 to 6.65 mg/L Dissolved and colloidal (<0.45 μm); 

near Mn mining operation

Caruso and others (2012)

Magela Creek, Australia 0.8 to 9.6 μg/L Dissolved (<0.02 μm); wet season 

flooding

Hart and others (1992)

Magela Creek, Australia 0.6 to 7.9 μg/L Particulate (>0.02 μm); wet season 

flooding

Hart and others (1992)

Tap water, southern Quebec, 

Canada

98 μg/L Mean; range is 1 to 2,700 μg/L Bouchard and others (2011)
L20 Critical Mineral Resources of the United States—Manganese

Table L4. Background and above-background concentrations of manganese in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency. Units of measure: cm, centimeter; mg/kg, milligram per kilogram (or part 

per million); mg/L, milligram per liter; mg/m3, milligram per cubic meter; ng/m3, nanogram per cubic meter; μg/L, microgram per liter; μg/m3, microgram per 

cubic meter; μm, micrometer]

Environment

and (or) location

Manganese 

concentration Unit Notes Reference(s)

Waters—Continued

Sediment, world rivers

Proposed EPA benchmark

Proposed EPA benchmark 

United States recommended 

upper limit in secondary 

drinking water 

1,679

2,300

120

50

mg/kg

μg/L

μg/L

μg/L

Average suspended sediment

Tier II secondary acute value 

for aquatic life

Tier II secondary chronic value 

for aquatic life

 None

Viers and others (2009)

Suter and Tsao (1996)

Suter and Tsao (1996)

U.S. Environmental Protection 

Agency (2013)

Air 

Greenland

Molango mining district, 

Mexico

South Pole

Urban/industrial

Minimum risk level 

chronic inhalation

Workplace limit

2.8 to 4.5

420

0.004 to 0.02

50 to 900

0.3

5

ng/m3

ng/m3

ng/m3

ng/m3

μg/m3

mg/m3

None

Mean; range is 3 to 5,860 ng/m3

None

None

Chronic duration greater than 

or equal to 1 year

Not-to-exceed limit over an 

8-hour workday

Kabata-Pendias and Pendias 

(2001)

Rodriguez-Agudelo and 

others (2006)

Kabata-Pendias and Pendias 

(2001)

Kabata-Pendias and Pendias 

(2001)

Agency for Toxic Substances 

and Disease Registry (2012)

Occupational Safety and Health 

Administration (2013)

In weathering environments at Earth’s surface, manganese commonly occurs in three active oxidation states. 

Manganese(II) occurs as dissolved Mn2+ in solution, especially 

under acidic pH values, as insoluble carbonate minerals, and 

as a minor component in other minerals. Manganese(III) is 

thermodynamically unstable and tends to be a temporary phase 

during redox reactions or a minor component in manganesebearing minerals. Manganese(IV) tends to form insoluble oxide 

and hydroxide minerals. The manganese(III) and manganese(IV) 

phases are some of the strongest oxidants in the environment, 

so the reactivity of manganese is dominated by its tendency to 

participate in redox reactions (Tebo and others, 2004). In surface 

water, manganese tends to attach to particulate matter that is 

either transported downstream or deposited as sediment.

Mining and industrial activities can lead to abovebackground concentrations of manganese in the environment 

(table L4). For example, manganese concentrations in soils 

near mining and industrial sources in Poland range from 

292 to 571 mg/kg (Połedniok and others, 2012). Comparable 

manganese concentrations in surface soils collected in central 

Pennsylvania from both an undisturbed watershed site and 

a nearby historic steel plant site (table L4) led Herndon and 

others (2011) to a comprehensive analysis that showed that 

more than one-half of the soils surveyed in North America and 

Europe are enriched in manganese as a result of atmospheric 

deposition from industrial sources. Indeed, the atmospheric 

concentrations of manganese in urban and industrial areas 

(50 to 900 ng/m3) are far greater than those measured at 

the poles (0.004 to 4.5 ng/m3). Likewise, the dissolved and 

colloidal (<0.45 μm-sized) fraction of manganese in river 

water near a manganese mining site in the Republic of Georgia 

contains 0.5 to 6.7 milligrams per liter (mg/L) (Caruso and 

others, 2012), which is roughly an order of magnitude greater 

than dissolved manganese concentrations in some world rivers 

(table L4). Mackey and others (1996) measured dissolved and 

colloidal manganese concentrations of seawater in nonpolluted 

and polluted harbors of Tasmania and found ranges of 

0.6 to 2 μg/L and 1 to 10 μg/L, respectively.
Environmental Considerations L21

The primary anthropogenic inputs of manganese to the 

atmosphere are from mining and ore processing, as well as 

from iron and steel production. The median anthropogenic 

emission of manganese to the atmosphere in 1983 was 

estimated to be 38.3 Mkg/yr (Nriagu and Pacyna, 1988), 

and the median manganese concentration reported for 

“polluted air” is 90 ng/m3 (Reimann and de Caritat, 1998, 

and references therein).

Because manganese is so critical to steel manufacturing, 

society’s need for manganese continues to increase. Unlike 

many metals, manganese recycling is relatively efficient. 

The recycled content (that is, the proportion of scrap) used 

in manganese production is between 25 and 50 percent, and 

the fraction of manganese in discarded products that get 

recycled is greater than 50 percent (Graedel and others, 2011). 

The factors that contribute to the economic and technologic 

feasibility of manganese recycling are that (a) manganese is 

used in large quantities, and (b) it occurs in alloys in relatively 

pure form, making it easier to remelt and recover (Reck and 

Graedel, 2012).

Mine Waste Characteristics

Manganese deposits are mined by open pit and underground methods. The mine waste generated is generally 

considered to be the material that originates and accumulates 

at a mine site that has no current economic value (Lottermoser, 

2010), and includes both solid and liquid waste. Because sedimentary manganese deposits tend to be laterally extensive, the 

expanse of land disrupted during mining is potentially great. 

The largest deposits on land include the Kalahari manganese 

field, which covers approximately 400 km2 (Jupiter Mines 

Ltd., 2013), and the Molango mining district, which covers 

180 km2 (Rivera-Becerril and others, 2013). Although many 

mines in sedimentary manganese deposits can be viewed 

from satellite imagery, values for the volume or tonnage of 

solid mine waste generated during mining are difficult to 

obtain. Supergene manganese deposits tend to be smaller and 

higher grade, so the amount of solid mine waste generated is 

expected to be less.

Solid mine waste includes overburden (the waste rock 

that overlies an orebody, which must be removed during 

mining) and gangue (the economically valueless material 

that surrounds or is mixed with the ore). The mineralogy of 

overburden varies depending upon the deposit, whereas the 

mineralogy of gangue from manganese mining can be similar 

among deposits. In general, gangue minerals in sedimentary 

manganese deposits include biogenic silica, carbonates, chert, 

clays, glauconite, pyrite, and quartz (Force and others, 1995). 

Gangue minerals in the Groote Eylandt supergene manganese 

deposit include gibbsite, goethite, kaolinite, and quartz, and 

the trace metals associated with these minerals include As, Ce, 

Cr, Gd, P, Pb, Sc, Th, Ti, and Zr (Pracejus and Bolton, 1992). 

Clarke and others (2010) have characterized mine tailings 

in the Kalahari manganese field as being predominantly 

sand-sized and containing the minerals birnessite, bixbyite, 

braunite, calcite, hausmannite, hematite, manganite, and 

todorokite. Trace metals detected in these tailings include 

(in milligrams per kilogram): Ag (<0.8), As (20), Be (0.9), 

Cd (0.6), Cr (42), Cu (41), Hg (0.1), Mo (5.9), Pb (110), Se 

(<0.2), and Th (<0.1). Most of these trace metals are expected 

to be immobile given the tailings pore water pH of 8.8.

The enrichment of trace metals (for example, Co, Cu, 

Mo, Ni, and REEs) in manganese oxide minerals is why 

manganese nodules and crusts on the seabed have become 

an attractive metal resource. In weathering environments, 

trace metals contained in manganese oxides are expected 

to remain immobile under oxic, near-neutral pH conditions 

(5<pH<8), especially in the presence of iron-rich solids 

(Smith and Huyck, 1999). Under anoxic and (or) acidic 

conditions, however, both manganese oxides and associated 

trace metals are expected to be mobile. In some manganese 

deposits, generation of acidity is possible if pyrite dissolution 

is not balanced by the acid-neutralizing capacity of carbonate 

minerals, but this scenario is expected to be rare. Dissolution 

of manganese oxides and associated trace elements is more 

likely to occur in soil environments, where organic acids 

can be highly concentrated in pore water, creating acidic 

conditions. Manganese oxide minerals are also expected to 

be less stable under anoxic conditions, which can develop 

upon burial. In anoxic subsurface environments, manganesereducing microorganisms are known to catalyze manganese 

oxide dissolution, thereby contributing to increased dissolved 

manganese and trace metal concentrations in groundwater 

(Myers and Nealson, 1988; Lovley, 1991).

Human Health Concerns

A wealth of information exists on the effects of 

manganese on human health because it is both an essential 

nutrient and a potential neurotoxin in cases of overexposure. 

The Agency for Toxic Substances and Disease Registry 

(ATSDR) provides a useful summary of the human toxicology 

of manganese (Agency for Toxic Substances and Disease 

Registry, 2012). The general public is exposed to manganese 

through consumption of food and water; inhalation of air; 

and skin contact with air, water, soil, and consumer products 

that contain manganese. The U.S. Recommended Dietary 

Allowance of manganese is 2 to 5 milligrams per day (mg/d) 

(National Research Council, 1989). Extended oral exposure 

above these levels can lead to adverse neurological effects. 

Workers at manganese mining and processing facilities 

are most likely to inhale manganese-rich dust. Inhaled 

manganese can lead to a permanent neurological disorder 

called manganism, which causes such symptoms as tremors, 

difficulty walking, and facial muscle spasms. Additionally, 

occupational exposure to manganese dust can lead to lung 

inflammation and impaired lung function. The mean air 

concentration within the Molango mining district in Mexico is 

420 micrograms of manganese per cubic meter; in this region, 
L22 Critical Mineral Resources of the United States—Manganese

surveyed adults exhibited decreased motor abilities with 

increasing air manganese concentrations (Rodríguez-Agudelo 

and others, 2006).

Children exposed to high levels of manganese show 

neurodevelopmental effects, including changes in behavior, 

memory, and learning ability. For example, surveyed 

school-age children in the Molango mining district contained 

significantly higher manganese levels in blood and hair 

relative to children in a control site; children also showed 

decreased intellectual function with increased manganese 

content in hair (Riojas-Rodríguez and others, 2010). In a 

similar study of school-aged children exposed to relatively 

high manganese levels (mean=98 micrograms of manganese 

per liter) in drinking water supplied by groundwater, higher 

manganese concentrations in hair were significantly correlated 

with higher manganese concentrations in drinking water and 

with decreased IQ scores (Bouchard and others, 2011).

The U.S. national secondary drinking water regulation for 

manganese, which is a nonenforceable guideline, is 50 μg/L 

(U.S Environmental Protection Agency, 2013). The World 

Health Organization’s (WHO) “Guidelines for Drinking Water 

Quality” included manganese at 400 μg/L until 2011, at which 

point the inclusion of manganese was discontinued on the 

grounds that the recommended value was well above normal 

concentrations of manganese in drinking water (World Health 

Organization, 2011). Frisbie and others (2012) argue that a 

substantial number of surveyed countries have drinking water 

manganese concentrations greater than 400 μg/L, and that 

given the possible negative human health effects of manganese, the earlier WHO guideline was probably too high to 

protect human health.

Manganese levels in soils are not regulated because 

of the ubiquity of manganese in soil and groundwater. The 

U.S. Environmental Protection Agency (EPA) does, however, 

require the reporting of manganese concentrations in waste 

streams from factories that use or produce manganese. The 

chronic duration (equal to or greater than 1 year) minimal risk 

level of 0.3 microgram per cubic meter (μg/m3) Mn has been 

set for manganese inhalation by the U.S. Department of Health 

and Human Services (Agency for Toxic Substances and Disease 

Registry, 2012). Likewise, the U.S. Occupational Safety and 

Health Administration (OSHA) has set a limit of 5 milligrams 

per cubic meter (mg/m3

) for the average allowable manganese 

concentration in workplace air over an 8-hour workday 

(Occupational Safety and Health Administration, 2013).

Ecological Health Concerns

Many studies focus on the ecological effects of manganese in the environment. Because fish tend to be sensitive to 

low concentrations of dissolved metals, they are commonly 

useful indicators of contamination in aquatic systems. One of 

several useful endpoints used in toxicity tests is the concentration that leads to 50 percent mortality (lethal concentration 50

[LC50]) after exposure to a substance for a certain amount of 

time. Chronic toxicity tests with dissolved manganese and 

developing rainbow trout (Oncorhynchus mykiss) revealed a 

mean LC50 value of 2.9 mg/L after 28 days of exposure (Birge 

and others, 1980). Tests on the chronic toxicity of manganese 

to developing brown trout (Salmo trutta Linnaeus) showed 

that at 30 mg/L hardness (high dissolved mineral content), 

decreased survival after 62 days correlated significantly with 

increased manganese concentrations, which ranged from 

0.5 to 15.5 mg/L (Stubblefield and others, 1997). The toxic 

effects of manganese on this species decreased with increasing 

hardness because less toxic Ca2+ and Mg2+ were able to 

outcompete with Mn2+ for biologic sites. Decreased manganese toxicity with increasing hardness was also observed for 

the water flea (Ceriodaphnia dubia) (Peters and others, 2011). 

Some regions of the United States have adopted secondary 

acute and chronic screening benchmarks for manganese 

concentrations of 2,300 μg/L and 120 μg/L, respectively, for 

aquatic freshwater life, as recommended by the Oakridge 

National Laboratory (Suter and Tsao, 1996).

Manganese is considered to be an essential nutrient for 

healthy plant growth, partially because of its role in supporting 

photosynthesis (Kabata-Pendias and Mukherjee, 2007). Land 

plants exhibit a range of 15 to 100 micrograms of manganese 

per gram dry weight (Misra and Mani, 1991). Plants vary in 

their ability to withstand high manganese concentrations, and 

those plants with higher manganese tolerance may accumulate 

manganese in roots and (or) precipitate manganese oxide 

within the epidermis (Kabata-Pendias and Mukherjee, 2007). 

A survey of plant roots growing in the Molango mining 

district in Mexico revealed that native ragweed growing in 

soil proximal to manganese mine waste was more resistant 

to manganese incorporation than ragweed growing directly 

in the mine waste; higher manganese resistance in the former 

specimen was attributed to a symbiotic relationship with 

mycorrhizal fungi in plant roots (Rivera-Becerril and others, 

2013). As summarized by Nagajyoti and others (2010), plants 

that are sensitive to excessive manganese tend to develop 

diseases that affect their leaves. Some regions of the United 

States have adopted a soil-screening benchmark of 500 milligrams of manganese per kilogram of soil for terrestrial 

plants, as recommended by the Oakridge National Laboratory 

(Efroymson and others, 1997).

Carbon Footprint

Use of manganese contributes to greenhouse gas 

emissions. During the production of ferromanganese, which 

is an important component in steel, manganese oxides 

are mixed with iron ore and carbon and then reduced in 

either a blast furnace or an electric arc furnace. The energy 

requirements for this process can be great. Additionally, 

when rhodochrosite (MnCO3 ) is the starting manganese ore, 

either for ferromanganese or nonmetallurgical applications, 

an initial processing step involves converting MnCO3 to 

pyrolusite (MnO2) in a process known as calcination. This 

results in additional process-related carbon dioxide (CO2 ) 

emissions. Using a basic emission calculation and assuming 
Problems and Future Research L23

100 percent calcination, an emission factor of 383 kilograms 

of carbon dioxide per kilogram of rhodochrosite results 

(Hanle and others, 2006). The EPA requires facilities that 

emit 25,000 metric tons or more of greenhouse gases and that 

consume at least 2,000 metric tons per year of carbonates, 

such as rhodochrosite, to report their process-based CO2

emissions in addition to their energy-related greenhouse gas 

emissions. Certain industries, such as manganese alloy and 

iron and steel production, are excluded from this requirement 

(U.S. Environmental Protection Agency, 2011).

Mine Closure

Most recently opened and new mining operations include 

closure plans that address issues related to the mine footprint. 

A mine’s footprint includes the waste left on site and locally 

affected soil and water, as well as ecological effects, such as 

habitat destruction and loss of biodiversity. The environmental 

impact assessment for the proposed Kalahari manganese 

mine in Northern Cape Province, South Africa, which is part 

of the Kalahari manganese field, provides examples of some 

mine footprint issues and possible control measures (Strategic 

Environmental Focus (Pty) Ltd., 2007). The manganese 

ore, mainly rhodochrosite, would be mined by underground 

methods, and both ore beneficiation and ferromanganese 

production would take place on site. The mine waste generated 

on site would include tailings (projected to be 250,000 metric 

tons per year) and slimes from ore treatment (projected to 

be 45,000 metric tons per year). The tailings pile would 

increase in size for about 10 years before being returned to 

the underground workings as backfill. Any remaining piles 

would be graded and covered with vegetation. Potential loss 

of native vegetation and disruption of faunal migrations would 

be addressed by a 1.5-km no-development buffer zone along 

the major river that flows through the proposed mine site and 

by revegetation of disturbed soils with native species. Surface 

and groundwater quality would be monitored frequently, 

and potential seepage from tailings and slime dams would 

be prevented through engineering means. The potential for 

dust emissions from mining and sinter operations would be 

mitigated through various methods, such as wet suppression 

and early revegetation.

Long-term plans for closed manganese mining operations will depend on the country, State, and local regulations. 

Potential long-term environmental impacts that could be of 

concern are dust generation and seepage of metal-rich water 

from waste material. An additional long-term issue relates 

to revegetation of manganese-rich waste, which applies to 

both closure of current mines and attempted reclamation of 

abandoned mines. Some plant species may contribute to soil 

acidification, which could lead to solubilization of manganese from minerals and transport of dissolved manganese 

downstream. Attention to the type of plant species used for 

revegetation and close monitoring of soil parameters during 

soil development might prevent this from happening.

Problems and Future Research

Identified global manganese resources, including 

currently subeconomic deposits, are essentially inexhaustible 

by any scenario of future consumption. Thus, there are no 

geologic problems related to the adequacy of manganese 

resources from a global perspective. Substantial increases 

in the price of manganese and increases in the efficiency of 

mining and processing will be necessary to convert much 

of this vast resource to economically workable deposits, 

however. The unequal distribution of the global manganese 

resources has long been recognized as a matter of concern 

for the United States and other industrialized countries that 

have inadequate domestic resources to meet their own needs. 

For the United States, discovery of manganese deposits of 

better quality than the few known large low-grade domestic 

deposits would be advantageous, whether they were immediately economic or could provide a more accessible source of 

manganese in the event of future supply disruptions.

Perhaps the most significant area available for domestic 

manganese exploration is the seabed within the United States 

EEZ. The vast size of this area alone provides incentive for 

research and exploration opportunities, many of which have 

been only superficially addressed to date. The United States 

EEZ contains more than 11 million km2 of seabed, an area larger 

than all of the land areas of the Nation. There are substantial 

known deposits of ferromanganese crusts and nodules in the 

EEZ where the United States has sole authority for regulation 

and mining. A more definitive characterization of the EEZ’s 

potential as a source of manganese could help to focus future 

research and exploration efforts on areas that have the highest 

possibility for development and production. An expanded 

information base regarding the occurrence, abundance, and 

metal grades of these deposits would be a significant step 

toward eventual metal production. Beyond geologic research, 

prime areas for research include mining and processing 

technology and advances in understanding and mitigating the 

potential impact of seabed mining on marine ecosystems.

With regard to seabed mineral resources, development 

of this resource would disturb potentially millions of square 

kilometers of the abyssal seabed, an unprecedentedly large 

area of mine disturbance that would constitute a new type of 

human impact on the environment. Significant new research 

would be needed to assess the effects on marine ecosystems 

and understand other environmental impacts of this completely 

new human activity and to minimize its adverse impacts. 

The ISA has developed guidelines for the environmental 

impact statement (EIS) that must be submitted by companies 

interested in applying for seabed mineral exploration licenses. 

In an EIS, developers must address all environmental and 

social issues and relevant mitigation measures that could result 

from the mining activity (International Seabed Authority, 

2012). More information on these issues has been discussed 

in the various technical reports published online by the 

ISA (2014).
L24 Critical Mineral Resources of the United States—Manganese

Processing of manganese nodules and crusts mined 

on the seabed will most likely take place on land, so the 

same mine waste issues mentioned for terrestrial manganese 

ores will also apply to seabed resources. Seabed mining 

in the CCZ would involve many additional issues. In 

addition to overcoming the technological challenges 

presented by mining as deep as 6,500 m below sea level, 

perhaps one of the most controversial issues is the potential 

impact on benthic, midwater, and pelagic ecosystems. 

Benthic ecosystems developed on the abyssal plain are 

the ecosystems that are the most threatened by mining in 

the CCZ. Understanding of abyssal, benthic ecosystems 

has grown rapidly in the past several decades. Even so, 

an international effort to evaluate the biodiversity in the 

CCZ revealed that “high, unanticipated, and still poorly 

sampled levels of species diversity for all three sedimentdwelling faunal components (foraminifera, nematodes 

and polychaetes) [exist] at [the] individual study sites” 

(International Seabed Authority, 2008). These findings led 

to the recommendation that “marine protected areas” be 

maintained throughout the CCZ in order to preserve abyssal 

biodiversity (International Seabed Authority, 2008). Another 

major environmental concern relates to wastewater being 

discharged from mining ships. This waste could contain 

crushed nodules and trace metals and has the potential to 

disrupt phytoplankton and, because the wastewater would 

be colder and denser than the surface water, to disrupt the 

natural circulation patterns in the water column (Markussen, 

1994). As a result, current conceptual engineering designs 

separate nodules from waste material (such as clay, 

sediments, and muddy water) and return waste to the seabed 

at depths of nearly 5,000 m below sea level so as not to 

disturb the midwater and pelagic regions (for example, 

Agarwal and others, 2012). As interest in mining seabed 

resources continues, additional studies would be needed to 

document its impacts on ocean life.

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U.S. Environmental Protection Agency, 2011, Miscellaneous uses 

of carbonate: Washington, D.C., U.S. Environmental Protection Agency Information Sheet EPA–430–F–09–007R, 3 p.

U.S. Environmental Protection Agency, 2013, National

primary drinking water regulations: U.S. Environmental

Protection Agency Web page, accessed February 15, 2013, 

at http://water.epa.gov/drink/contaminants/index.cfm.

U.S. Geological Survey, 1996, Global 30 arc-second elevation

(GTOPO30): Reston, Va., U.S. Geological Survey dataset 

(digital elevation model), accessed March 13, 2014, at 

https://lta.cr.usgs.gov/GTOPO30.

Varentsov, I.M., 1996, Manganese ores of supergene zone—

Geochemistry of formation: Dordrecht, Netherlands, 

Kluwer Academic Publishers, 343 p.

Viers, Jérôme, Dupré, Bernard, and Gaillardet, Jérôme, 2009, 

Chemical composition of suspended sediments in world 

rivers—New insights from a new database: Science of the 

Total Environment, v. 407, no. 2, p. 853–868. [Also available at http://dx.doi.org/10.1016/j.scitotenv.2008.09.053.]

World Health Organization, 2011, Guidelines for drinkingwater quality (4th ed.): Geneva, Switzerland, World Health 

Organization, 541 p. [Also available at http://www.who.int/

water_sanitation_health/publications/2011/dwq_guidelines/en/.]
Niobium and Tantalum

By Klaus J. Schulz, Nadine M. Piatak, and John F. Papp

Chapter M of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–M

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter M

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract.......................................................................................................................................................M1

Introduction.................................................................................................................................................M2

Uses, Applications, and Consumption ...........................................................................................M2

Niobium......................................................................................................................................M2

Tantalum ....................................................................................................................................M2

Strategic and Critical Resource Issues ........................................................................................M4

Geology........................................................................................................................................................M5

Geochemistry.....................................................................................................................................M5

Mineralogy.........................................................................................................................................M6

Deposit Types ....................................................................................................................................M6

Carbonatites and Associated Rocks ..................................................................................M14

Alkaline to Peralkaline Granites and Syenites ..................................................................M16

Rare-Metal Granites and Lithium-Cesium-Tantalum–Type Pegmatites .......................M17

Resources and Production .....................................................................................................................M20

Niobium Reserves and Other Identified Resources .................................................................M20

Tantalum Reserves and Other Identified Resources ...............................................................M21

Undiscovered Niobium and Tantalum Resources .....................................................................M21

Niobium and Tantalum Production ...............................................................................................M22

Niobium....................................................................................................................................M22

Tantalum ..................................................................................................................................M22

Exploration for New Deposits ................................................................................................................M24

Environmental Considerations ...............................................................................................................M24

Sources and Fate in the Environment .........................................................................................M24

Mine Waste Characteristics ........................................................................................................M25

Human Health Concerns ...............................................................................................................M26

Ecological Health Concerns .........................................................................................................M26

Carbon Footprint..............................................................................................................................M27

Mine Closure ...................................................................................................................................M27

Problems and Future Research .............................................................................................................M28

References Cited......................................................................................................................................M28

Suggested citation:

Schulz, K.J., Piatak, N.M., and Papp, J.F., 2017, Niobium and Tantalum, chap. M of Schulz, K.J., DeYoung, J.H., Jr., 

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. M1– M34, http://dx.doi.org/

10.3133/pp1802M.
Niobium and Tantalum

By Klaus J. Schulz, Nadine M. Piatak, and John F. Papp

Abstract

Niobium and tantalum are transition metals that are 

almost always found together in nature because they have 

very similar physical and chemical properties. Their properties 

of hardness, conductivity, and resistance to corrosion largely 

determine their primary uses today. The leading use of 

niobium (about 75 percent) is in the production of highstrength steel alloys used in pipelines, transportation infrastructure, and structural applications. Electronic capacitors 

are the leading use of tantalum for high-end applications, 

including cell phones, computer hard drives, and such 

implantable medical devices as pacemakers. Niobium and 

tantalum are considered critical and strategic metals based on 

the potential risks to their supply (because current production 

is restricted to only a few countries) and the significant effects 

that a restriction in supply would have on the defense, energy, 

high-tech industrial, and medical sectors.

The average abundance of niobium and tantalum in 

bulk continental crust is relatively low—8.0 parts per million 

(ppm) niobium and 0.7 ppm tantalum. Their chemical 

characteristics, such as small ionic size and high electronic 

field strength, significantly reduce the potential for these 

elements to substitute for more common elements in rockforming minerals and make niobium and tantalum essentially 

immobile in most aqueous solutions. Niobium and tantalum 

do not occur naturally as pure metals but are concentrated in a 

variety of relatively rare oxide and hydroxide minerals, as well 

as in a few rare silicate minerals. Niobium is primarily derived 

from the complex oxide minerals of the pyrochlore group 

((Na,Ca,Ce)2

 (Nb,Ti,Ta)2 (O,OH,F) 7 ), which are found in some 

alkaline granite-syenite complexes (that is, igneous rocks 

containing sodium- or potassium-rich minerals and little or no 

quartz) and carbonatites (that is, igneous rocks that are more 

than 50 percent composed of primary carbonate minerals, by 

volume). Tantalum is derived mostly from the mineral tantalite 

((Fe,Mn)(Ta,Nb) 2

O6 ), which is found as an accessory mineral 

in rare-metal granites and pegmatites that are also enriched in 

lithium and cesium (termed lithium-cesium-tantalum (LCT)-

type pegmatites).

Brazil and Canada are the leading nations that produce 

niobium mineral concentrates, but Brazil is by far the leading 

producer, accounting for about 90 percent of production, 

which comes mostly from weathered material derived from 

carbonatites. Brazil and Canada also have the largest identified 

niobium resources; additional resources, although they are less 

well reported, occur in Angola, Australia, China, Greenland, 

Malawi, Russia, and South Africa. Australia and Brazil have 

been the leading producers of tantalum mineral concentrates, 

although recently Ethiopia and Mozambique have also 

been significant suppliers of tantalum. Artisanal mining of 

columbite-tantalite (also called coltan) is practiced in many 

countries, particularly Burundi, the Democratic Republic of 

the Congo (Congo [Kinshasa]), Nigeria, Rwanda, and Uganda. 

Brazil has about 40 percent of the identified tantalum 

resources; other countries and regions with identified tantalum 

resources include, in decreasing order of resources, Australia, 

Asia, Russia and the Middle East, Africa, North America, 

and Europe. Identified niobium and tantalum resources in the 

United States are small, low grade, and difficult to recover and 

process, and are thus not commercially recoverable at current 

prices. Consequently, the United States meets its current and 

expected future needs for niobium and tantalum through 

imports of mineral concentrates and alloys and through 

recovery from foreign and domestic alloy scrap that contain 

the metals.

 Environmentally, the main issues related to niobium and 

tantalum mining are land disruptions, the volume of waste 

materials and their disposal, and the radioactivity of some 

tailings and waste materials that contain thorium and uranium. 

Because of the relative biological inertness of niobium and 

tantalum, human and ecological health concerns are generally 

minimal under most natural conditions.

Demand for both niobium and tantalum is expected to 

increase as the world economy continues to recover from the 

downturn that began in 2008. Increased demand for niobium is 

linked to increased consumption of microalloyed steel, which 

is used in the manufacture of cars, buildings, ships, and refinery 

equipment. Demand for these steels will likely increase with 

continued economic development in such countries as Brazil, 

China, and India. In addition, increased global demand for 

cars, cell phones, computers, superconducting magnets, and 

other high-tech devices will likely spur increased demand for 

both niobium and tantalum. The estimated global reserves and 
M2 Critical Mineral Resources of the United States—Niobium and Tantalum

resources of niobium and tantalum are large and appear 

more than sufficient to meet global demand for the foreseeable 

future, possibly the next 500 years. The sale of “conflict 

coltan” attributed to rebel forces waging a civil war in 

Congo (Kinshasa) has been of recent concern and has highlighted the need for a transparent and traceable global supply 

chain that can exclude illegal columbite-tantalite from the 

conventional market while discerning legitimate artisanal mine 

production in central Africa.

Introduction

Niobium (Nb) and tantalum (Ta) are transition metals 

that are very similar in their physical and chemical properties 

(table M1). They are almost always found together in nature. 

Both are named after tragic figures from Greek mythology—

Niobium is named after Niobe, and tantalum is named after 

Niobe’s father, Tantalus. Niobium was first discovered in 1801 

by English chemist Charles Hatchett, who named it columbium after Columbia, the poetic name for North America. 

Columbium was used interchangeably with niobium until 

1949 when niobium was officially accepted as the designated 

name by the International Union of Pure and Applied 

Chemistry. Tantalum was first discovered by Swedish scientist 

Anders Ekeberg in 1802, but because it was difficult to 

separate from niobium, tantalum was confused with niobium 

until 1864, when it was confirmed to be a separate element. 

Relatively pure tantalum metal was not produced until 1903.

Niobium is a lustrous, gray, ductile metal with a high 

melting point, relatively low density, and superconductor 

properties. Tantalum is a dark blue-gray, dense, ductile, very 

hard, and easily fabricated metal. It is highly conductive to heat 

and electricity and is renowned for its resistance to corrosion 

by acids. It is these special properties, especially hardness, 

conductivity, and resistance to corrosion, that determine the 

primary uses of niobium and tantalum today (table M1).

Uses, Applications, and Consumption

Niobium

The leading use of niobium is in ferrous metallurgy. 

About 75 percent of world niobium production is used by the 

steel industry for the production of a variety of steel alloys 

that contain small amounts of other metals (microalloy and 

low-alloy steels) to improve corrosion resistance, strength, 

toughness, and other properties (Papp, 2013b). Niobium acts 

as a grain refiner and precipitation hardener in high-strength 

low-alloy and microalloyed steel, simultaneously improving 

mechanical strength, high-temperature strength, and toughness 

and improving corrosion resistance. These steels are used in 

pipelines, transportation, and structural applications (Roskill 

Information Services Ltd., 2009). Appreciable amounts of 

niobium are contained in nickel-, cobalt-, and iron-base 

superalloys used for high-temperature applications, such as 

jet engine components, gas turbines, rocket subassemblies, 

turbocharger systems, and heat-resisting and combustion 

equipment. Niobium alloys are also used in the manufacture 

of superconducting magnets used in magnetic resonance 

imaging (MRI) and nuclear magnetic resonance instruments 

(NMRI), as well as in particle accelerators, such as the Large 

Hadron Collider in Europe, which contains the world’s 

largest superconducting magnets (fig. M1). A new use of 

niobium is in a solid niobic acid that acts as a catalyst in the 

conversion of palm oil to biodiesel (Schwela, 2010). With 

increased production of biofuels in Brazil and elsewhere, this 

use of niobium could constitute a significant future market. 

Niobium’s uses are specialized, and substitution is generally 

likely to result in reduced performance and higher cost.

World consumption of niobium, by material produced, 

is shown in figure M2A (Schwela, 2011). The greatest 

consumption is of ferroniobium, which is an iron-niobium 

alloy that contains 60 to 70 percent niobium. Leading uses 

of niobium are for steelmaking (87 percent), followed by 

niobium chemicals (5.2 percent), vacuum-grade niobium 

(2.7 percent), other niobium alloys (2.5 percent), and 

niobium metal (0.9 percent). In 2011, U.S. consumption of 

niobium (measured as elemental niobium) was 9,160 metric 

tons, which was an increase of more than 12 percent from 

consumption in 2010 (Papp, 2013b).

Tantalum

The electronics industry accounts for about one-half 

of tantalum consumption, mainly as powder and wire 

(Schwela, 2010). Electronic capacitors are the leading end 

use of tantalum owing to tantalum’s particular ability to store 

and release energy. Because of this ability, components can 

be exceptionally small and are favored in space-sensitive, 

high-end applications, such as telecommunications (for cell 

phones), data storage (for hard drives), and implantable 

medical devices (for hearing aids and pacemakers). To date, 

Table M1. Selected properties of niobium and tantalum.

[Source: Winter (2014). Å, angstrom; °C, degree Celsius; g/cm3, gram per 

cubic centimeter; nΩ-m, nano ohm-meter]

Property Niobium (Nb) Tantalum (Ta)

Atomic number 41 73

Atomic weight 92.90638 180.94788

Atomic radius (Å) 1.46 1.46

Density (g/cm3

) 8.57 16.65

Melting point (°C) 2,477 3,017

Boiling point (°C) 4,744 5,458

Hardness (Mohs scale) 6.0 6.5

Electrical resistivity 

(nΩ-m)

152 at 0 °C 131 at 20 °C

Crystal structure Body-centered

cubic

Body-centered

cubic
Introduction M3

Toroids

Figure M1.

Figure M1. Photograph (central 

view) of the ATLAS detector in the 

Large Hadron Collider showing its 

eight superconducting barrel toroid 

magnets around the calorimeter. 

(For scale, note the person 

[circled] standing in the lower 

center portion of the photograph.) 

Together, these toroids make up the 

world’s largest superconducting 

magnet. Photograph of the ATLAS 

Experiment Copyright© 2014 CERN.

no effective substitute has been found for tantalum 

in electronic devices without loss in performance. 

Tantalum’s low mechanical strength and high biocompatibility allows it to be used as a coating on stronger 

substrates, such as stainless steel, for such medical 

applications as stents to support blood vessels, plates, 

bone replacements, and suture clips and wire. In addition, 

tantalum is used to impart strength and high temperature 

resistance to cracking in the manufacture of superalloys 

for use in aerospace applications and energy generation. 

Its resistance to corrosion makes tantalum useful in the 

chemical industry, generally as a lining to pipes, tanks, 

and vessels. Tantalum oxide is used to increase the 

refractive index of lens glass, whereas the hardness of 

tantalum carbide makes it ideal for cutting tools.

World consumption of tantalum, by material 

produced, is shown in figure M2B (Schwela, 2011). 

The leading uses are capacitor-grade tantalum powder 

(24 percent) and tantalum mill products (22 percent) 

followed by tantalum chemicals (18 percent); 

metallurgical-grade powder, unwrought metal, scrap, and 

other uses (17 percent); tantalum ingot (12 percent); and 

tantalum carbide (7 percent). In 2011, U.S. consumption 

of tantalum (measured as elemental tantalum) was 

1,210 metric tons, which was an increase of 4 percent 

from consumption in 2010 (Papp, 2013c).

B. Tantalum consumption

Capacitor-grade

Ta powder

24%

Ta mill

products

22%

Ta chemicals

18%

Metallurgicalgrade powder,

unwrought metal,

scrap, and other

17% 

Ta ingot

12%

Ta carbide

7%

Steel-grade FeNb

87%

Nb chemicals 5.2%

Vacuum-grade Nb 2.7%

Nb alloys 2.5%

Nb metal 0.9%

A. Niobium consumption

Fig M2.

Figure M2. Pie charts showing percentage of reported 

world consumption of A, niobium, and B, tantalum, by 

material produced (Schwela, 2011). Percentages have been 

rounded to two significant figures and do not add to 100. 

FeNb, ferroniobium; Nb, niobium; Ta, tantalum
M4 Critical Mineral Resources of the United States—Niobium and Tantalum

Strategic and Critical Resource Issues

Niobium and tantalum are considered critical and strategic 

metals owing to their specialized and important applications in 

the defense, energy, high-tech industrial, and medical sectors; 

because they are essential in many products and applications 

but generally have no effective substitutes; and because the 

mine production of these minerals takes place in only a few 

countries. The geographic concentration of niobium resources 

and production, and, to a lesser extent, those of tantalum, make 

their supply vulnerable to potential influence and disruption 

by such factors as civil unrest, environmental issues, market 

manipulation, natural disasters, and political changes. In their 

evaluation of the criticality of selected mineral commodities, 

the National Research Council (2008) ranked niobium as one 

of the most critical minerals, along with manganese, platinumgroup metals (especially rhodium and platinum), and rare-earth 

elements (REEs), and ranked tantalum as moderately critical 

based on risks to supply and the potential impact of supply 

restrictions (fig. M3). In its analysis of critical raw materials 

for the European Union, the European Commission reached a 

similar conclusion on the criticality of niobium and tantalum 

(European Commission, 2010).

The estimated global reserves and resources of both 

niobium and tantalum are large, but they are unevenly distributed 

geographically. The United States has had no reported mine 

production of niobium and tantalum since the late 1950s, and 

identified resources are mostly of low grade and are considered 

uneconomic at current commodity prices (Papp, 2013b, c). 

As a result, the United States has for many years been totally 

dependent on imports and recycling to meet its niobium and 

tantalum needs. This chapter gives a geologic perspective of 

niobium and tantalum resources with the purpose of providing 

an overview of how, why, and where sources of niobium and 

tantalum have formed and the geographic distribution of these 

sources around the world. Also included is a discussion of 

environmental issues that may derive from current and future 

niobium and tantalum mining.

1 (low) 2 3 4 (high)

Supply risk

1 (low)

2

3

4 (high)

Impact of supply restriction

Cu

Ta

Li

Nb

Rh

Pt

Pd

REE

Mn

Figure M3.

Ti

V Ga

In

Figure M3. Criticality matrix for niobium, tantalum, 

and selected other mineral commodities. Of all these 

commodities, indium, manganese, niobium, platinumgroup metals (palladium, platinum, and rhodium), and 

rare-earth elements fall into the most critical zone of 

high supply risk and high impact of supply restriction 

(after National Research Council, 2008, p. 165). 

Cu, copper; Ga, gallium; In, indium; Li, lithium; 

Mn, manganese; Nb, niobium; Pd, palladium; Pt, 

platinum; REE, rare-earth element(s); Rh, rhodium; 

Ta, tantalum; Ti, titanium; and V, vanadium
Geology M5

Geology

Geochemistry

Niobium and tantalum generally show strongly coherent 

geochemical behavior because they are identical in charge 

(5+ under most geologically relevant oxidation conditions) 

and nearly identical in effective ionic radius (table M1). As a 

result, they are closely associated and found together in most 

rocks and minerals in which they occur. Both elements are 

lithophile, in that they show a strong affinity for oxygen, and 

they are high-field-strength elements (HFSEs), meaning that 

their ions are relatively small and have intense electrostatic 

fields. Their HFSE characteristics significantly reduce their 

potential to substitute for more common elements in most 

rock-forming minerals and make them essentially immobile 

under most natural conditions (Wood, 2005); as a result, their 

concentrations in the surface environment are generally low. 

The average abundance of niobium and tantalum in bulk 

continental crust is 8.0 parts per million (ppm) niobium and 

0.7 ppm tantalum (Rudnick and Gao, 2003). In surface waters, 

concentrations are generally in the low parts per trillion.

Niobium and tantalum are depleted in continental crust 

relative to other highly incompatible elements (elements that 

do not readily substitute for more common elements in major 

rock-forming minerals), such as cesium, rubidium, thorium, 

and uranium, and the light REEs, including cerium and 

lanthanum. As a result, continental crustal rocks are typically 

characterized by high lanthanum to niobium (La:Nb) ratios—

the average crustal ratio is about 2.5 (Rudnick and Gao, 2003). 

The depletion of niobium and tantalum in continental crust 

is attributed to the formation of crustal rocks at convergent 

margins above subduction zones where titanium-rich minerals 

that host niobium and tantalum may remain as residual 

phases in the source region during generation of the magmas 

(Kelemen and others, 2003).

Although niobium and tantalum generally show coherent 

geochemical behavior, some chemical processes are able to 

separate them, which results in preferential enrichment or 

depletion of one or the other. These processes are still poorly 

understood, although crystal fractionation during magma 

evolution is the most commonly invoked mechanism. Niobium 

and tantalum show limited substitution for tin, tungsten, and 

zirconium in some rock-forming minerals, but particularly for 

titanium in such minerals as ilmenite (FeTiO3 ), rutile (TiO2), 

and titanite (sphene) (CaTiSiO5). As niobium and tantalum 

become more highly enriched in residual igneous melts, a 

variety of chiefly oxide and hydroxide niobium and tantalum 

minerals may form (see Mineralogy section and table M2), 

depending on melt composition, temperature, pressure, and 

fluid composition. In addition, some evidence suggests that 

later alteration by hydrothermal (hot) fluids—in particular, 

concentrated fluoride solutions (Wood, 2005)—may play a 

role in mobilizing and enriching niobium and tantalum.

Table M2. Selected niobium and tantalum oxide minerals and their end-member Nb2O5 and Ta2O5 contents or compositional range.

[Data are from Mackay and Simandl (2014) and Parker and Fleischer (1968). Nb2O5, niobium pentoxide; Ta2O5, tantalum pentoxide; NP, not present]

Mineral name Mineral group Chemical formula Nb2

O5

(percent)

Ta2O5

(percent)

Columbite Columbite-tantalite (Fe,Mn)(Nb,Ta)2

O6 46.8 to 81.2 5.3 to 31.2

Tantalite Columbite-tantalite (Fe,Mn)(Ta,)2

O6 0.3 to 26.8 56.5 to 86.2

Pyrochlore Pyrochlore (Na,Ca,Ce)2

(Nb,Ti,Ta)2(O,OH,F)7 34.3 to 86.8 Trace to 4.3

Microlite Pyrochlore (Na,Ca)2

(Ta,Nb,Ti)2(O,OH,F)7 Trace to 3.4 73.7 to 86.0

Tapiolite Tapiolite (Fe,Mn)(Ta,Nb)2

O6 0 to 35 Up to 86

Ixiolite Ixiolite (Ta,Fe,Sn,Nb,Mn)4

O8 8.3 68.96

Wodginite Wodginite (Ta,Sn,Mn,Nb,Fe,Ti)O2 1.35 to 7.1 67.5 to 70.1

Loparite Perovskite (Ce,La,Na,Ca,Sr)(Ti,Nb)O3 9.7 to 15.1 0.8 to 1.0

Lueshite Perovskite NaNbO3 81.1 Trace

Latrappite Perovskite (Ca,Na)(Nb, Ti, Fe)O3 43.9 NP

Euxenite Euxenite (Y,Ca,Ce,U,Th)(Nb,Ti,Ta)2

O6 21 to 34 1 to 22

Strüverite Rutile (Ti,Ta,Fe)O2 6.5 to 9.3 33.1 to 35.5

Ilmenorutile Rutile Fex

(Nb,Ta) 2x4Ti1– xO2 Up to 33 Up to 14
M6 Critical Mineral Resources of the United States—Niobium and Tantalum

Mineralogy

Niobium and tantalum do not occur naturally as pure 

metals, but they are essential components in a variety of oxide 

and hydroxide minerals, as well as in a few rare silicates and 

one borate (Parker and Fleischer, 1968). The economically 

important mineral species are all oxides (table M2); pyrochlore is the principal ore mineral for niobium, and tantalite 

is the principal ore mineral for tantalum (fig. M4). Until the 

discovery of pyrochlore-rich deposits in the 1950s, niobium 

was produced as a byproduct of mining columbite-tantalitebearing pegmatites for tantalum.

The columbite-tantalite mineral group is characterized 

by a wide range in niobium and tantalum content as well as 

variable iron and manganese content. Columbite (a niobiumrich mineral) and tantalite (a tantalum-rich mineral) have very 

similar properties because they have the same structure and 

similar chemistries, but tantalite has a much higher specific 

gravity (8.0 +) than does columbite (5.2). Tantalite also has 

a dimorphic relation (same chemistry but different crystal 

structure) to the mineral tapiolite. Columbite-tantalite minerals 

are found as accessory phases in rare-metal granites and 

pegmatites (Černý, 1991a, b).

The pyrochlore mineral group also has a wide compositional range, but generally has niobium greatly enriched 

over tantalum (a ratio of about 200:1). In addition, pyrochlore 

can contain significant amounts of barium, light REEs, and 

uranium. Because it can contain uranium, pyrochlore tends 

to be radioactive and can show damage to its crystal structure 

from radiation. Pyrochlore is typically found as a primary 

mineral in alkaline igneous rocks, particularly carbonatites 

(for example, the Araxá carbonatite in Brazil).

Other oxide minerals, such as loparite, ixiolite, tapiolite, 

and the perovskite group (table M2) are less common sources 

of niobium and tantalum. Loparite concentrate from the 

Lovozero intrusion in the Kola Peninsula has been a source 

of niobium in Russia but is not economically competitive 

with pyrochlore. The calcium- and niobium-dominant 

perovskite-group mineral latrappite is the main ore mineral 

at the Oka carbonatite complex in Quebec, Canada. Microlite 

is the tantalum-rich end member of the pyrochlore mineral 

group and is generally found in pegmatites in association with 

columbite-tantalite. The titanium-bearing mineral strüverite 

has been a low-grade source of tantalum recovered from 

tin-mining waste in Southeast Asia. Niobium- and tantalumbearing silicate minerals are relatively rare and found mostly 

in alkaline igneous rocks.

Deposit Types

Niobium and tantalum minerals are found in a variety 

of igneous rocks around the world but only rarely in concentrations great enough to be of economic interest. Primary 

niobium and tantalum mineral deposits are found in three main 

types of igneous intrusive rocks (table M3; Küster, 2009):

1. Carbonatites and associated alkaline rocks 

(Nb dominant),

2. Alkaline to peralkaline granites and syenites 

(Nb dominant), and

3. Rare-metal granites and pegmatites of the lithiumcesium-tantalum (LCT) family (Ta dominant) 

(Černý and Ercit, 2005).

A

Figure M4.

B

Figure M4. Photographs showing A, centimeter-size pyrochlore crystals from Uganda, and B, a tantalite crystal. Photograph A is by 

Gunnar Ries/CC-BY-SA-3.0; photograph B is by Roger Weller, Cochise College.
Geology M7

In addition, some secondary concentrations have been 

formed by weathering of primary deposits (laterites) and 

by sedimentary processes (placers). The secondary deposits 

are of particular interest because they can be less expensive 

to mine and can have higher grades than primary hard-rock 

deposits. Because secondary deposits typically occur in close 

proximity to their primary sources, they are not described 

separately below.

Deposits of niobium and tantalum are found around the 

world (fig. M5, table M4), but major production is currently 

restricted to only a few countries. Grade and tonnage figures 

for these deposits are shown in figure M6. Generally, grades 

for niobium are higher (from about 0.1 to 3.0 percent 

niobium pentoxide [Nb2O5]) than those of tantalum (generally < 0.1 percent tantalum pentoxide [Ta2

O5]). Carbonatites 

generally have the highest grades and tonnages of niobium, 

although they overlap with those of alkaline granite and 

syenite-hosted deposits. In addition, some alkaline graniteand syenite-hosted deposits have higher tonnages at similar 

grades of tantalum than do rare-metal granite and pegmatitehosted deposits (fig. M6A). The ratios of niobium to tantalum 

in the alkaline granite and syenite deposits are much higher 

(>10 to 20) because of the predominance of niobium-rich 

minerals, such as pyrochlore and columbite.

Table M3. Major types of niobium and tantalum deposits, with key characteristics and examples.

[Modified from British Geological Survey (2011). Grades and tonnages are highly variable among deposits, and the grades and tonnages given in this table are 

generalizations only. LCT, lithium-cesium-tantalum; Mt, million metric tons; Nb, niobium; Nb2O5, niobium pentoxide; Ta, tantalum; Ta2O5, tantalum pentoxide] 

Deposit type Brief description Typical grades and tonnage Major examples

Carbonatite-hosted 

primary deposits

Nb >Ta: Niobium deposits commonly

consisting of members of the 

perovskite and pyrochlore mineral 

groups found within carbonatite 

intrusions in alkaline igneous 

provinces

Deposits show a wide range in both 

grade and tonnage. Morro dos Seis 

Lagos is the largest reported deposit; 

it contains about 2,900 Mt at a grade 

of 2.85 percent Nb2

O5. More typical

is the deposit at Niobec, which

contains about 46 Mt at a grade

of 0.53 percent Nb2

O5

Niobec and Oka, Canada; 

Araxá, Catalão I and II,

and Morro dos Seis 

Lagos, Brazil

Alkaline granite 

and syenite 

Nb >Ta: Deposits containing niobium 

and lesser amounts of tantalum; 

the deposits are related to silicic 

alkaline granite and syenite igneous 

intrusions

Generally <1,000 Mt at grades of 

0.1 to 1 percent Nb2

O5 and 

< 0.05 percent Ta2

O5

Motzfeldt and Ilímaussaq, 

Greenland; Lovozero, 

Russia; Thor Lake and 

Strange Lake, Canada

Rare-metal granite Ta > Nb: Deposits containing tantalum 

and lesser amounts of niobium; the 

deposits are generally found in the 

uppermost parts of peraluminous 

and (commonly) hydrothermally 

altered late-stage granitic plutons

Generally <100 Mt at grades of 

< 0.05 Ta2

O5

Yichun, China; 

Abu Dabbab and 

Nuweibi, Egypt

LCT-type pegmatite Ta > Nb: Deposits containing tantalum 

and lesser amounts of niobium; the 

deposits are LCT-enriched-type 

pegmatites

Generally <100 Mt at grades of

< 0.05 Ta2

O5

Greenbushes and 

Wodgina, Australia; 

Tanco, Canada; Volta 

Grande, Brazil; 

Kenticha, Ethiopia

Secondary (regolith) 

deposits

Niobium and (or) tantalum ore minerals concentrated in zones of intense 

weathering above carbonatite and 

granite or pegmatite intrusions, or in 

sedimentary placer deposits derived 

from such intrusions

Lateritic deposits generally have 

<1,000 Mt at grades of up to 

3 percent Nb2O5. Placer deposits, 

such as the deposit at Tomtor, 

can have very high grades of

up to 12 percent Nb2

O5

Araxá and Catalão, Brazil; 

Tomtor, Russia; Greenbushes, Australia
M8 Critical Mineral Resources of the United States—Niobium and Tantalum 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10° 20° 30°

40°50°

80°

70°

70°

80°

60°

Blue River Tanco

Thor Lake

Strange Lake

Niobec

Crevier

Oka

Araxá

Catalão I, II

Volta Grande

Pitinga

Morro dos Seis Lagos

Ilimaussaq

Motzfeldt

Tamazeght

Ghurayyah Abu Dabbab

Nuweibi Umm Naggat

Bonga

Kenticha

Kanyika

MuianeMorrua

Marropino

Tantalite Valley

Three Aloes

Pilanesberg

Donsa

Eagle

Lovozero Tomtor

Katuginskoye

Etykinskoye

Vishnyalovskoye

Voznesenovskoye

Alakhinskoye

Halzan-Buregtei

Yichun Nanping

801

Sokli

Beauvoir-Echassieres

Greenbushes

Wodgina

Kougarok

Elk Creek

Iron Hill 

 (Powderhorn)

Bald Hill

Mount Cattlin

Mount Deans

Dubbo

Bayan Obo

Mutala

Jos Plateau

Kibara pegmatite belt

Central Nigerian pegmatite belt

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Niobium Figure M5 

Nb-Ta mines

CarbonatiteRare-metal graniteLCT pegmatiteAlkaline granite and syeniteAlkaline granite and syenite

Nb-Ta occurrences

CarbonatiteRare-metal granite

EXPLANATION

nes are currently active. World map showing locations of selected niobium and tantalum mines, deposits, and occurrences, by deposit type. Not all the mi

a, tantalum

Figure M5.LCT, lithium-cesium-tantalum; Nb, niobium; T
Geology M9

0.01 0.1 1.0 10 100 1,000 10,000

0.01

0.1

1.0

Blue River

Tanco

Dubbo

Crevier

Volta

Grande

Ilimaussaq

Abu Dabbab Ghurayyah

Nuweibi Umm Naggat

Kanyika

Marropino

Morrua

Muiane

Tantalite Valley

Three Aloes

Eagle

Donsa

Katuginskoye

Etykinskoye Voznesenovskoye

Yichun

Nanping

801

Greenbushes

Wodgina

Grade, in weight percent Ta O2 5

0.1 1.0 10 100 1,000 10,000

0.001

0.01

0.1

1.0

10

Blue River

Thor 

Lake Niobec

Oka

Araxá R

Araxá P

Volta Grande

Pitinga

Morro dos 

Seis Lagos

Ghurayyah

Abu Dabbab

Bonga

Kanyika

Voznesenovskoye

Elk Creek

Iron Hill

(Powderhorn)

Grade, in weight percent Nb O2 5

A

B

Motzfeldt

Thor Lake

Bald

Hill

Motzfeldt

Strange Crevier Lake Mount Deans

Mount Cattlin

Dubbo

Mount 

Deans

Kenticha

Bayan Obo

Carbonatite Rare-metal granite Lithium-cesium-

 tantalum (LCT) pegmatite

Alkaline granite

 and syenite

Catalão II

Catalão I

Resources, in million metric tons

Ilimaussaq

EXPLANATION

Alakhinskoye

Beauvoir-Echassieres

Figure M6. Figure M6. Log-log plots of deposit grades and tonnages of A, tantalum, and B, niobium, by 

deposit type. The data include different levels of probability, including measured, indicated, and 

inferred resources and (or) proven and probable reserves. Data and sources are given in table M4. 

P, primary deposit; R, residual deposit, Nb2

O5, niobium pentoxide; Ta2O5, tantalum pentoxide
M10 Critical Mineral Resources of the United States—Niobium and Tantalum

Table M4. Location, grade, tonnage, and other data for selected niobium-tantalum deposits of the world.—Continued

[Symbol size (small, large) corresponds to size of symbols shown in figure M5; deposit grade and tonnage are shown graphically in figure M6. Symbol size: 

S, small; L, large. WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative 

values for longitude indicate that the deposit is in the Western Hemisphere. Nb2O5, niobium pentoxide; Ta2O5, tantalum pentoxide. Resource categories: 

M+In+If, measured+indicated+inferred; In+If, indicated+inferred. LCT, lithium-cesium-tantalum; Ma, mega-annum; NA, not available]

Deposit Country Deposit type Symbol 

size

Latitude Longitude

(decimal degrees WGS 84)

Bonga Angola Carbonatite S –14.27 13.97 

Bald Hill Australia LCT pegmatite L –31.52 122.18 

Greenbushes Australia LCT pegmatite L –33.87 116.06 

Mount Cattlin Australia LCT pegmatite L –33.56 120.04 

Mount Deans Australia LCT pegmatite L –32.31 121.79 

Wodgina Australia LCT pegmatite L –21.18 118.68 

Dubbo Australia Alkaline granite

and syenite

L –32.42 148.60 

Araxá Brazil Carbonatite L –19.67 –46.95 

Araxá Brazil Carbonatite L –19.67 –46.95 

Catalão I Brazil Carbonatite L –18.13 –47.80 

Catalão I Brazil Carbonatite L –18.13 –47.80 

Catalão I Brazil Carbonatite L –18.13 –47.80 

Catalão II Brazil Carbonatite L –18.03 –47.87 

Volta Grande Brazil LCT pegmatite L –21.08 –44.58 

Pitinga Brazil Alkaline granite

and syenite

L 0.78 –60.08 

Morro dos Seis Lagos Brazil Carbonatite S 0.28 –66.68 

Blue River Canada Carbonatite S 52.32 –119.17 

Tanco Canada LCT pegmatite L 50.43 –95.45 

Thor Lake Canada Alkaline granite

and syenite

S 62.11 –112.60 

Strange Lake Canada Alkaline granite

and syenite

S 56.30 –64.12 

Niobec Canada Carbonatite L 48.53 –71.15 

Crevier Canada Carbonatite S 49.50 –72.82 

Oka Canada Carbonatite S 45.50 –74.02 

Yichun China Rare-metal granite L 27.65 114.52 

Nanping China LCT pegmatite L 26.67 118.10 

801 China Rare-metal granite L 44.62 120.97 

Bayan Obo China Carbonatite L 41.80 109.97 

Kibara pegmatite 

belt

Democratic Republic of the 

Congo and Rwanda

LCT pegmatite L –2.00 28.00 

Abu Dabbab Egypt Rare-metal granite S 25.34 34.54 

Nuweibi Egypt Rare-metal granite S 25.20 34.50 

Umm Naggat Egypt Rare-metal granite S 25.20 34.50 

Kenticha Ethiopia LCT pegmatite L 5.52 39.03 
Geology M11

Table M4. Location, grade, tonnage, and other data for selected niobium-tantalum deposits of the world.—Continued

[Symbol size (small, large) corresponds to size of symbols shown in figure M5; deposit grade and tonnage are shown graphically in figure M6. Symbol size: 

S, small; L, large. WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative 

values for longitude indicate that the deposit is in the Western Hemisphere. Nb O , niobium pentoxide; Ta O , tantalum pentoxide. Resource categories: 2 5 2 5

M+In+If, measured+indicated+inferred; In+If, indicated+inferred. LCT, lithium-cesium-tantalum; Ma, mega-annum; NA, not available]

Age

Resource 

(million 

metric tons)

Grade 

(% Nb O ) 2 5

Grade 

(% Ta O ) 2 5

Resource 

categories Reference(s) Deposit

130.8 ± 1.4 Ma

Archean?

2,527 ± 2 Ma

Archean?

Archean?

2,829 ± 11 Ma

Jurassic

NA

NA

85 ± 6.9 Ma 

85 ± 6.9 Ma 

85 ± 6.9 Ma 

NA

NA

About 1,820 Ma

NA

NA

NA

NA

NA

629 to 656 Ma 

(496 ± 17 Ma)

879 Ma

117 to 95 Ma

NA

NA

NA

NA

1,000 to 900 Ma

NA

NA

NA

Cambrian

824

2

135.1

17.155

9.1

86.5

73.2

462

940

19

29

55.2

3.4

6.32

164

2,897

39.75

2.1

312

229.9

419

59.79

13.85

0.068

0.0423

6.8

800

NA

44.5

98

25

116.4

0.48

NA

NA

NA

0.06

NA

0.46

2.48

1.6

1.08

1.22

1.27

1.67

0.0092

0.2

2.85

0.172

NA

0.36

0.172

0.42

0.146

0.66

NA

NA

NA

0.13

NA

0.012

NA

NA

NA

NA

0.038

0.022

0.0155

0.022

0.032

0.03

NA

NA

NA

NA

NA

NA

0.0375

NA

NA

0.0195

0.216

0.029

NA

NA

0.0175

NA

0.017 to 0.02

0.03

0.022

NA

NA

0.025

0.014

0.0151

0.017

NA

M+In+If

M+In+If

M+In+If

In+If

M+In+If

M+In+If

Residual deposit

Primary

Residual deposit

Rock total reserve

M+In+If

Reserves

Reserves

Disseminated ore 

with tin

NA

M+In+If

Pre-production, 1991

M+In+If

M+In+If

M+In+If

M+In+If

M+In+If

NA

NA

NA

NA

NA

M+In+If

NA

NA

Reserves

Cordeiro and others, 2011

Fetherston, 2004 

Fetherston, 2004

Fetherston, 2004; Galaxy 

Resources Ltd., 2012

Fetherston, 2004

Fetherston, 2004

Alkane Resources Ltd., 

2012

Cordeiro and others, 2011

Cordeiro and others, 2011

Cordeiro and others, 2011

Cordeiro and others, 2011

Anglo American plc, 

2012, p. 198

Cordeiro and others, 2011

Schwela, 2010

Bastos Neto and others, 

2009, p. 1329 –1357

Pollard, 1995

Chong and Postolski, 2011

Fetherston, 2004

Cox and others, 2011

Quest Rare Minerals Ltd., 

2012

IAMGOLD Corp., 2012, 

p. 26

MDN Inc., 2012

Niocan Inc., 2011 

Fetherston, 2004

Fetherston, 2004

Fetherston, 2004

Berger and others, 2009

Melcher and others, 2015

Fetherston, 2004; Gippsland 

Ltd., 2011

Papp, 2013a; Gippsland 

Ltd., 2011

Fetherston, 2004

Fetherston, 2004

Bonga

Bald Hill

Greenbushes

Mount Cattlin

Mount Deans

Wodgina

Dubbo

Araxá

Araxá

Catalão I

Catalão I

Catalão I

Catalão II

Volta Grande

Pitinga

Morro dos Seis Lagos

Blue River

Tanco

Thor Lake

Strange Lake

Niobec

Crevier

Oka

Yichun

Nanping

801

Bayan Obo

Kibara pegmatite 

belt

Abu Dabbab

Nuweibi

Umm Naggat

Kenticha
M12 Critical Mineral Resources of the United States—Niobium and Tantalum

Table M4. Location, grade, tonnage, and other data for selected niobium-tantalum deposits of the world.— Continued

[Symbol size (small, large) corresponds to size of symbols shown in figure M5; deposit grade and tonnage are shown graphically in figure M6. Symbol size: 

S, small; L, large. WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative 

values for longitude indicate that the deposit is in the Western Hemisphere. Nb O , niobium pentoxide; Ta O , tantalum pentoxide. Resource categories: 2 5 2 5

M+In+If, measured+indicated+inferred; In+If, indicated+inferred. LCT, lithium-cesium-tantalum; Ma, mega-annum; NA, not available]

Latitude Longitude Symbol Deposit Country Deposit type size (decimal degrees WGS 84)

Sokli Finland Carbonatite S 67.77 29.23

Beauvoir-Echassieres France Rare-metal granite S 46.18 2.93

Ilimaussaq Greenland Alkaline granite S 61.01 –45.71

and syenite

Motzfeldt Greenland Alkaline granite S 61.20 –45.00

and syenite

Kanyika Malawi Alkaline granite S –12.63 33.63

and syenite

Halzan-Buregtei Mongolia Alkaline granite S 48.42 91.97

and syenite

Tamazeght Morocco Alkaline granite S 32.61 –4.53

and syenite

Marropino Mozambique LCT pegmatite L –16.51 37.90

Morrua Mozambique LCT pegmatite L –16.27 37.87

Muiane Mozambique LCT pegmatite L –15.74 38.25

Mutala Mozambique LCT pegmatite L –15.92 37.92

Tantalite Valley Namibia LCT pegmatite L –28.72 18.80

Three Aloes Namibia LCT pegmatite L –21.22 14.87

Central Nigerian Nigeria LCT pegmatite S 8.00 6.00

pegmatite belt

Jos Plateau Nigeria Alkaline granite L 9.17 9.75

and syenite

Tomtor Russia Carbonatite S 71.05 116.57

Katuginskoye Russia Alkaline granite S 56.50 119.33

and syenite

Etykinskoye Russia Rare-metal granite L 47.57 134.72

Vishnyalovskoye Russia LCT pegmatite L 55.22 97.75

Voznesenovskoye Russia Rare-metal granite S 44.17 132.19

Alakhinskoye Russia Rare-metal granite S 49.44 86.96

Lovozero Russia Alkaline granite/ L 67.89 34.61

syentite

Ghurayyah Saudi Arabia Alkaline granite S 27.92 35.80

and syenite

Pilanesberg South Africa Alkaline granite S –24.23 27.07

and syenite

Kougarok United States Rare-metal granite S 65.71 –165.23

Elk Creek United States Carbonatite S 40.27 –96.19

Iron Hill (Powderhorn) United States Carbonatite S 38.23 –106.71

Eagle Zimbabwe LCT pegmatite L –18.50 32.43

Donsa Zimbabwe LCT pegmatite L –18.50 32.43
Geology M13

Table M4. Location, grade, tonnage, and other data for selected niobium-tantalum deposits of the world.—Continued

[Symbol size (small, large) corresponds to size of symbols shown in figure M5; deposit grade and tonnage are shown graphically in figure M6. Symbol size: 

S, small; L, large. WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative 

values for longitude indicate that the deposit is in the Western Hemisphere. Nb O , niobium pentoxide; Ta O , tantalum pentoxide. Resource categories: 2 5 2 5

M+In+If, measured+indicated+inferred; In+If, indicated+inferred. LCT, lithium-cesium-tantalum; Ma, mega-annum; NA, not available]

Age

Resource 

(million 

metric tons)

Grade 

(% Nb O ) 2 5

Grade 

(% Ta O ) 2 5

Resource 

categories Reference(s) Deposit

NA

308 ± 2 Ma

NA

NA

NA

NA

About 40 Ma

NA

NA

NA

NA

NA

NA

535 to 555 Ma

About 198 Ma

700 Ma and 

400 Ma (?)

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

110

0.02

54

340

60

NA

NA

7.396

7.77

1.375

NA

0.74

7.2

NA

NA

NA

774

0.0127

NA

17.2

0.015 to 0.020

NA

385

NA

NA

102.6

0.374

0.61

1.62

?

NA

0.11

0.46

0.29

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

NA

0.016

NA

NA

0.284

NA

NA

0.638

0.057

NA

NA

NA

0.012

0.04

0.185

0.014

NA

NA

0.0223

0.0248

0.025

NA

0.043

0.05

NA

NA

NA

0.025

0.013

NA

0.012

0.013 to 0.017

NA

0.0245

NA

NA

NA

NA

0.034

0.025

Hard rock ore

NA

Resources

M+In+If

M+In+If

NA

NA

M+In+If

NA

M+In+If

NA

NA

NA

NA

NA

Placer

Reserves

Reserves

NA

NA

NA

NA

NA

NA

NA

In+If

NA

NA

NA

Siirama, 2009

Raimbault and others, 1985

Salvi and Williams-Jones, 

2005; Fetherston, 2004

Ram Resources Ltd., 2012, 

p. 8

Papp, 2013a

Fetherston, 2004

Marks and others, 2008

Noventa Ltd., 2011, p. 91

Noventa Ltd., 2011, p. 91

Global Mining Services 

Pty. Ltd., 2009, p. 57

Papp, 2013a

Fetherston, 2004

Fetherston, 2004

Matheis, 1987; Melcher 

and others, 2015

Melcher and others, 2015

Pollard, 1995

Fetherston, 2004

Fetherston, 2004

Fetherston, 2004

Fetherston, 2004

Fetherston, 2004

Salvi and Williams-Jones, 

2005; Fetherston, 2004

Tertiary Minerals plc, 2013

Lurie, 1986

Fetherston, 2004

Daigle, 2012

Long and others, 2010

Fetherston, 2004

Fetherston, 2004

Sokli

Beauvoir-Echassieres

Ilimaussaq

Motzfeldt

Kanyika

Halzan-Buregtei

Tamazeght

Marropino

Morrua

Muiane

Mutala

Tantalite Valley

Three Aloes

Central Nigerian 

pegmatite belt

Jos Plateau

Tomtor

Katuginskoye

Etykinskoye

Vishnyalovskoye

Voznesenovskoye

Alakhinskoye

Lovozero

Ghurayyah

Pilanesberg

Kougarok

Elk Creek

Iron Hill (Powderhorn)

Eagle

Donsa
M14 Critical Mineral Resources of the United States—Niobium and Tantalum

Carbonatites and Associated Rocks

Carbonatites are igneous rocks that are more than 

50 percent composed of primary carbonate minerals, by 

volume, mainly calcite (CaCO3), dolomite (CaMg(CO3)2), 

and ankerite (Ca(Mg,Fe2+,Mn)(CO3)2). Associated minerals 

commonly include alkali-rich (sodium or potassium) 

pyroxenes and amphiboles, phlogopite (magnesium-rich 

mica), magnetite, apatite (calcium phosphate), and olivine 

(iron-magnesium silicate). In addition, a large number of rare 

or exotic minerals may be present. Most carbonatites occur 

as relatively small (up to a few tens of square kilometers) 

intrusions, such as dikes, sills, and small plug- or crescentshaped bodies (fig. M7A); more rarely, they can occur as 

larger intrusions (plutons) or as volcanic deposits (lava flows 

and fragmental rocks). In three dimensions, carbonatite 

intrusions typically form vertical pipe-like bodies (fig. M7B). 

Carbonatites are commonly spatially associated with alkaline 

silicate rocks characterized by a wide range of sodium- or 

potassium-rich minerals and little or no quartz (Barker, 1989). 

Many carbonatites are surrounded by an aureole of chemically 

altered country rocks (fenites) produced by the reaction of 

nearby rocks with hot sodium- and (or) potassium-rich fluids 

derived from the carbonatite (fig. M7).

A recent compilation by Woolley and Kjarsgaard 

(2008) lists 527 carbonatites worldwide. Most carbonatites 

(88 percent of all dated carbonatites) are located in cratons 

(the oldest, stable portions of continents) and were emplaced 

during periods of continental extension or rifting. In many 

carbonatite-bearing provinces, there have been repeated 

episodes of carbonatite emplacement, with as many as five 

episodes separated by hundreds of millions of years and 

extending from the late Archean to the present. Sixty-four 

percent of dated carbonatites are Paleozoic in age, however 

(Woolley and Bailey, 2012). Their source magmas are thought 

to be derived directly from the mantle by relatively small 

degrees of partial melting and emplaced along major faults or 

crustal boundaries reopened in response to major changes in 

global plate-tectonic movements (Woolley and Bailey, 2012).

N

0 1,000 METERS

0 3,000 FEET

VERTICAL EXAGGERATION ×9

1

2

3

3

2

1

A'

A'

A

A

?

300

450

600

1

2

3

2

1

Ordovician limestone and glacial till

Central core dolomitic carbonatite

1 Rare-earth-element-bearing zone

2 Rare-earth-element-deficient zone

Dolomitic carbonatite

1 Low-niobium hematized carbonatite

2 Pyrochlore-apatite±rare-earth- 

 element-bearing carbonatite

3 Low-niobium dolomitic carbonatite

Calcitic carbonatite—Locally enriched 

 in pyroxene, magnetite, and mica

Alkali syenite, feldspathoid syenite

Nepheline garnet syenite, nepheline syenite

Fenitized (altered) host rocks

Unaltered gneiss and other host rocks

Geologic contact

Niobec Mine headframe

3

A

B

EXPLANATION

0 1,000 METERS

0 3,000 FEET

Figure M7. Diagrams showing A , the subsurface geology of the Saint-Honoré carbonatite complex 

in southern Quebec, Canada, and B , a schematic north-south cross-section along line A– A’ in A. The 

location of the Niobec underground mine is also shown. Modified from Richardson and Birkett (1996).
Geology M15

Carbonatites are typically enriched in barium, fluorine, 

niobium, phosphorus, REEs, strontium, thorium, uranium, 

and zirconium. Not all carbonatites show similar enrichments 

in all elements, however; those related to carbohydrothermal 

fluids (hot fluids rich in carbon dioxide and water ± fluorine) 

typically have enriched REEs and strontium, but not niobium, 

whereas carbonatites associated with alkaline igneous rocks 

are typically enriched in niobium, phosphorus, and titanium, 

but not REEs (Mitchell, 2005). The most common niobiumbearing minerals in carbonatites include members of the 

pyrochlore and perovskite mineral groups, as well as niobiumrich silicates, such as titanite. The diversity of mineral types 

in carbonate-hosted deposits, their diverse compositions, and 

textural evidence for replacement and resorption of mineral 

phases suggest that the niobium-bearing minerals represent 

transported and (or) mixed mineral assemblages and are not 

in situ products of crystallization (Mitchell, 2005). Along with 

serving as the major source of niobium, carbonatites can also 

host deposits of barite, copper, fluorite, magnetite, phosphate, 

REEs, titanium, and vermiculite (Mariano, 1989).

Brazil is the world’s leading supplier of niobium (about 

90 percent); its major deposits occur in Late Cretaceous 

carbonatite complexes. These complexes were emplaced along 

deep-seated faults located along the southwestern border of the 

ancient (Archean) São Francisco craton (Cordeiro and others, 

2011). The Araxá deposit is the largest operating deposit; it 

has more than 460 million metric tons of weathered ore with 

a mean grade of 2.48 percent Nb2O5 (fig. M5; Cordeiro and 

others, 2011). The Barreiro carbonatite complex that hosts 

the Araxá deposit is approximately circular in shape with 

a diameter of about 4.5 kilometers (km); it is composed of 

carbonatite, glimmerite (altered ultramafic rock composed 

almost entirely of biotite or phlogopite), and phoscorite (a 

rock composed of magnetite, apatite, and one of the silicate 

minerals clinopyroxene, olivine, or phylogopite) (Nasraoui and 

Waerenborgh, 2001). The complex is surrounded by a 2.5-kmwide aureole of metasomatized (fluid-altered), alkali-enriched 

quartzite and mica schist country rocks. Phoscorite, which 

forms thick masses in dolomitic carbonatite in the central part 

of the complex, shows the highest concentrations of pyrochlore, the main niobium-bearing ore mineral, either as individual grains or in veins associated with magnetite. Intensive 

tropical weathering over the central part of the carbonatite 

complex has produced a lateritic cover up to 230 meters (m) 

thick that is enriched in pyrochlore and comprises the residual 

ore at Araxá (Nasraoui and Waerenborgh, 2001). The lateritic 

residual ore is exploited by open pit mining.

At Catalão, Brazil, niobium deposits are associated 

with two alkaline-carbonatite complexes, Catalão I and 

Catalão II (Cordeiro and others, 2011). The Catalão I complex 

consists of a vertical pipe-like, zoned intrusion about 6 

km in diameter at the surface. Two pipe-like orebodies, the 

Mine II and the East Area, have been defined (Cordeiro 

and others, 2011). The Mine II orebody is hosted mainly by 

dolomitic carbonatite, and the East Area orebody is hosted by 

glimmerite. Both orebodies have been confirmed by drilling 

to extend to at least 800 m depth. The main niobium-bearing 

mineral in both deposits is pyrochlore, which in the East 

Area deposit is concentrated in late-stage dikes composed of 

nelsonite (magnetite-apatite-phlogopite rock). As at Araxá, 

current mining at Catalão II is in the weathered lateritic zone 

above the center of the complex.

The leading producer of niobium outside of Brazil is 

the Niobec Mine in Quebec, Canada. This mine is the only 

operating underground niobium mine in the world. The 

Niobec Mine is hosted by the Saint-Honoré carbonatite 

complex dated at 650 mega-annum (Ma) (Belzile, 2009), 

which is covered by Paleozoic limestone and glacial deposits. 

The ellipsoidal Saint-Honoré carbonatite complex is zoned 

with an outer ring of feldspathic (feldspar-bearing) and 

feldspathoidal (minerals that resemble feldspar but have a 

different structure and much lower silica content) alkaline 

rocks and an inner zone consisting of a series of carbonatite 

lenses (fig. M7). Like the Brazilian deposits, pyrochlore is 

the main niobium-bearing mineral, but it is very fine-grained 

(0.2 to 0.8 millimeters) and rarely visible. Although disseminated throughout the carbonatite, pyrochlore is especially 

concentrated in steeply dipping (>70 degrees) mineralized 

lenses that are rich in, in order of abundance, magnetite, 

apatite, and biotite (Belzile, 2009). Some lenses extend to 

a vertical depth of at least 730 m and have average grades 

of between 0.44 and 0.51 percent Nb2O5. Other carbonatite 

complexes in Canada, including the nearby Crevier syenitecarbonatite complex, the Oka carbonatite complex to the 

south, and the Blue River complex in British Columbia, are 

currently being evaluated for their niobium resources.

Other carbonatite complexes around the world are known 

to host niobium mineralization, but none are currently in 

production. Two complexes are claimed to contain particularly 

large resources. The Tomtor complex in northern Siberia, 

Russia, is claimed to contain more niobium than the Araxá 

complex in Brazil along with significant REE resources 

(Kravchenko and Pokrovsky, 1995); grade and tonnage 

figures for the deposit are not publicly available. Although the 

central carbonatite in the Tomtor complex is reported to be 

mineralized, the higher grade niobium (>12 percent niobium) 

and REE-rich ore is in weathered carbonatite and in a buried 

placer deposit interpreted to have formed in an ancient lake 

overlying the complex (Kravchenko and Pokrovsky, 1995). 

The Morro dos Seis Lagos carbonatite in northwestern 

Brazil is poorly known but is reported to contain the world’s 

largest single niobium deposit; its resources are said to 

be 2,900 million metric tons grading 2.85 percent Nb2O5

(Pollard, 1995).

Woolley and Kjarsgaard (2008) report 23 carbonatites in 

the United States. Of these, the Elk Creek carbonatite, which 

is located in the subsurface of Nebraska, is currently being 

evaluated. The inferred resources at the deposit are reported 

to be 102.6 million metric tons grading 0.638 percent Nb2O5

(Daigle, 2012).
M16 Critical Mineral Resources of the United States—Niobium and Tantalum

Alkaline to Peralkaline Granites and Syenites

The term “alkaline” encompasses a large variety of 

igneous rocks containing certain sodium- or potassium-rich 

minerals, such as feldspathoids and (or) alkali-rich pyroxenes 

and amphiboles and generally little or no quartz (Sørensen, 

1974). Peralkaline rocks are a subset of alkaline rocks and are 

characterized by having a molecular amount of Na2O +K2O 

that exceeds Al2O3 (termed the agpaitic index >1). Alkaline 

rocks can range from ultramafic to felsic and from silicaundersaturated (feldspathoid-bearing) to silica-oversaturated 

(quartz-bearing). They are most commonly found in intraplate 

settings, such as continental rift zones (for example, the 

East African rift zone) and oceanic islands (for example, the 

Azores, Hawaii, and Réunion), but they can also be found in 

some orogenic belts where they were emplaced after major 

tectonic deformation. An important feature of alkaline rocks, 

but particularly peralkaline rocks, is unusual and locally 

extreme enrichment in alkalis; HFSEs, including zirconium, 

niobium, titanium, uranium, yttrium, and REEs; and halogens, 

such as fluorine and chlorine (Salvi and Williams-Jones, 

2005). As a result, alkaline igneous rocks can contain mineral 

deposits with high contents of HFSEs, including niobium, but 

they less commonly have high contents of tantalum.

Alkaline magmas are the product of low degrees of 

partial melting in a deep, chemically undepleted mantle (Niu 

and O’Hara, 2003). As a result, alkaline magmas are enriched 

over many other magma types in incompatible elements, 

including the HFSEs. These incompatible elements become 

further enriched as the alkaline magmas cool and crystallize, 

becoming concentrated in the most evolved, volatile-rich 

granitic and syenitic (same general composition as granite 

but with potassium feldspar dominant and little or no quartz 

[< 5 percent]) melts and eventually forming disseminated ore 

minerals in the crystallized rocks. In addition, because the 

HFSEs form relatively dense minerals, they can accumulate 

into layers through crystal settling and be transported and 

mixed by currents or new injections of magma (Mitchell, 

2005). Many alkaline igneous complexes also show extensive 

evidence of hydrothermal alteration, commonly with the most 

altered rocks also having the highest HFSE and REE contents. 

Although the role of fluids in the transport of HFSEs and REEs 

is still a matter of debate, growing evidence suggests that these 

elements are mobile in fluids that are enriched in fluorine, 

chlorine, and (or) carbon dioxide (CO2

) (Salvi and WilliamsJones, 2005). Typical ore minerals in mineralized peralkaline 

granites are usually niobium-rich pyrochlore and columbite.

Alkaline intrusive complexes are not currently major 

sources of niobium and tantalum production, although 

exploration is ongoing in some areas (for example, Strange 

Lake in Canada and Ghurayyah in Saudi Arabia). Examples 

of alkaline igneous complexes include the Pitinga complex 

in Brazil, the Strange Lake and Thor Lake complexes in 

Canada, the Ilímaussaq and Motzfeldt complexes in southern 

Greenland, the very large Lovozero complex in Russia, and 

the Pilanesberg complex in South Africa (fig. M5). Alkaline 

intrusive complexes have been identified in the United States, 

including Bokan Mountain in Alaska and Magnet Cove in 

Arkansas, but they are not known to contain significant 

niobium and tantalum resources of current economic interest.

The Devonian Lovozero alkaline complex on the Kola 

Peninsula of Russia and the adjacent Khibina complex 

together make up the largest peralkaline igneous intrusions in the world, covering an area of about 2,000 square 

kilometers (km2 ). The Khibina complex is host to one of the 

largest phosphate resources in the world (2.7 billion metric 

tons averaging 17.5 percent phosphorus pentoxide [P2O5]), 

whereas the Lovozero complex hosts zirconium, niobium, 

REEs, yttrium, strontium, barium, and phosphorus, which 

occur as eudialyte (a zirconium-bearing silicate mineral), 

loparite, and apatite (Salvi and Williams-Jones, 2005). The 

Lovozero complex, which is exposed over an area of 650 km2, 

occurs as a trough-shaped intrusion composed of three main 

intrusive phases (fig. M8) (Kogarko and others, 2010). The 

oldest phase (phase 1), which accounts for about 5 percent of 

the total volume, consists of nepheline syenites with varied 

mineralogy, but it contains no loparite or eudialyte. The next 

phase, (phase II), which accounts for about 77 percent of 

the total volume, consists of layered nepheline syenites with 

varied mineralogy and locally accessory loparite or eudialyte. 

Loparite is concentrated up to 25 percent, by volume, in some 

layers (Kogarko and others, 2002). Loparite concentrate has 

been produced intermittently during the past 50 years; about 

30,000 metric tons grading 8 percent Nb2O5 and 0.7 percent 

Ta2O5 has reportedly been produced annually (Semenov, 

1997). The third intrusive phase (phase III), which is also 

known as the eudialyte complex, accounts for about 18 percent 

of the total volume and consists of nepheline syenites that, 

particularly in the upper part of the complex, can contain 

seams and lenses of almost monomineralic eudialyte (50 to 

85 percent) (Kogarko and others, 2010).

The Mesoproterozoic Ilímaussaq alkaline complex in 

southwestern Greenland is one of the most studied alkaline 

complexes in the world (Sørensen, 2001). It is 1 of 10 alkaline 

intrusions in the Gardar igneous province, which is a failed 

continental rift that was active between 1,350 and 1,140 Ma 

(Upton and others, 2003). The Ilímaussaq alkaline complex 

is the type locality for the rock-type peralkaline nepheline 

syenite and also 27 minerals. It contains large concentrations 

of a number of rare elements, including beryllium, lithium, 

niobium, REEs, uranium and thorium, yttrium, and zirconium, 

contained in about 220 different minerals, of which 9 are 

unique to this complex (Sørensen, 2001). The complex covers 

136 km2

 and has an exposed vertical thickness of 1,700 m; 

its total depth is not known. The main rare-metal deposits 

with potentially exploitable niobium, REEs, yttrium, and 

zirconium are present in 29 separate eudialyte-rich layers 

interpreted to have formed by crystal settling (Sørensen, 

2001). Cumulatively, these rocks are estimated to contain a 

resource of about 54 million metric tons grading 1.1 percent 

zirconium dioxide (ZrO2), 0.09 percent yttrium oxide (Y2O3), 

0.56 percent rare-earth oxides, and 0.11 percent Nb2

O5
Geology M17

Phase I—Nepheline syenite

Phase II—Layered nepheline syenites

 with loparite concentrations

Phase III—Nepheline syenites 

 with eudialyte concentrations

Loparite

Eudialyte

EXPLANATION

Figure M8

Layering

NOT TO SCALE

Figure M8. Schematic cross-section of the Lovozero alkaline intrusion, Kola Peninsula, Russia, showing the relation 

among the three intrusive phases and the niobium mineralization contained in eudialyte and loparite. After Kogarko 

and others (2002).

(Steenfelt, 1991). The Motzfeldt intrusion, which is another 

of the alkaline intrusions in the Gardar igneous province, has 

zones of hydrothermally altered syenite containing niobiumtantalum-REE-bearing pyrochlore mineralization (Steenfelt, 

1991), which is the focus of ongoing exploration.

Rare-Metal Granites and Lithium-CesiumTantalum–Type Pegmatites

All economically important tantalum mineralization 

is related to rare-metal granites (also called rare-element 

granites) and lithium-cesium-tantalum (LCT)-type pegmatites. 

The rare-metal granites are generally peraluminous (have 

molecular Al2O3>[CaO+Na2O+K2O]), muscovite- and albiterich granites that display high degrees of chemical fractionation, and represent the last stages of felsic magma evolution 

in upwardly differentiated granitic intrusions (Linnen and 

Cuney, 2005). The parental magmas are formed by partial 

melting of preexisting crustal rocks, particularly aluminous 

sediments, and are generally emplaced at shallow levels of the 

crust (in the upper few kilometers) during the late stages of or 

after major tectonic deformation and regional metamorphism 

in orogenic belts. They may show pervasive hydrothermal 

alteration and host disseminated tantalum and niobium 

mineralization, as well as tin and tungsten, in complex vein 

systems (stockworks) that developed from circulation of 

late-stage hydrothermal fluids. Typical mineralization consists 

of microlite, columbite-tantalite, tantalum-rich cassiterite (tin 

oxide), and lepidolite (lithium-rich mica).

LCT-type pegmatites are generally small (meters rather 

than kilometers in length and width) granitic intrusions 

characterized by extremely coarse but variable grain-size 

and enrichments in lithium, rubidium, cesium, beryllium, 

tantalum, and niobium (Ta>Nb) (Černý and Ercit, 2005). 

They are the products of highly fractionated and volatile-rich 

granitic magmas generally derived from rare-metal granites. 

They commonly occur in aureoles surrounding the roof of 

their parental granite intrusion, and the mineralized and most 

fractionated pegmatites are found the farthest away (fig. M9). 

Most LCT-type pegmatites are concentrically but irregularly 

zoned (fig. M10), and typically have a thin border zone, a 

wall zone, an intermediate zone, an albite zone, and a core 

zone (Cameron and others, 1949; Fetherston, 2004). Tantalum 

mineralization is mostly concentrated in the intermediate and 

albite zones and generally consists of columbite-tantalite, 

ixiolite, microlite, and (or) wodginite. Additional mineralization may be present, including beryllium (as beryl, a 

beryllium aluminum silicate), cesium (as pollucite, a cesium 
M18 Critical Mineral Resources of the United States—Niobium and Tantalum

Parental rare-metal granite

Barren

Be

Be, Nb, TaBe, Nb, Ta

Li, Be, Ta, Nb

Li, Cs, Be, Ta, Nb

Migration of mineralized pegmatite melt

Pegmatite aureoles

in country rock

showing increasing

fractionation, volatile

enrichment, zoning

complexity, and extent

of metasomatic replacement as mineralizing fluids

move outwards from the

host granite

Pegmatite bodies

Rare-metal elements

EXPLANATION

Figure M9.

NOT TO SCALE

Be Beryllium

Cs Cesium

Li Lithium

Nb Niobium

Ta Tantalum

Figure M9. Schematic representation of regional lithium-cesium-tantalum (LCT) rare-metal-bearing pegmatite zoning above a parental 

granite. After Černý (1991b).

Border zone—Commonly a very thin zone of fine-

 grained albite and quartz and sometimes muscovite

Wall zone—Niobium>tantalum. Coarse-grained

 quartz, K-feldspar, albite, and muscovite, with

 accessory beryl, columbite, muscovite, and tourmaline

Intermediate zone—May range from 0 to 10 zones.

 Tantalum>niobium. Large crystals of amblygonite,

 quartz, K-feldspar, muscovite, and spodumene are

 common. Concentrations of beryllium, cesium, 

 lithium, tantalum, niobium, tin, titanium, and 

 zirconium mineralization may be present

Albite zone—Massive cleavelandite, often replaced

 by sugary albite. Contains minor amounts of quartz

 and muscovite. Host for tantalum mineralization.

 In areas of fracture filling, may be replaced by

 quartz-muscovite-rich metasomatic griesen

Core zone—Massive quartz, occasionally with

 K-feldspar. Margin with large crystals of 

 beryl, spodumene, and tourmaline

NOT TO SCALE

EXPLANATION

Figure M10. Figure M10. Schematic cross-section of a concentrically zoned lithium-cesium-tantalum (LCT) rare-metal-bearing pegmatite. After 

Černý (1991a) and Fetherston (2004).
Geology M19

zeolite), lithium (as spodumene, a lithium pyroxene; petalite, 

a lithium aluminum silicate; and lepidolite), tin (as cassiterite, 

a tin oxide), and a number of gemstones. LCT-type pegmatites 

are also mined for albite, muscovite, potassium feldspar, 

and ultrapure quartz. Like their parental rare-metal granites, 

LCT-type pegmatites are widely distributed globally (fig. M5), 

and range in age from Archean to Mesozoic, but they are 

found to be concentrated particularly during times of continental collision and supercontinent assembly (Bradley and 

McCauley, 2013). By far the largest LCT-pegmatite-hosted 

mineral deposits are of Archean age, however (for example, 

the Greenbushes and Wodgina deposits in Australia and the 

Tanco deposit in Canada).

In recent years, pegmatites in the State of Western 

Australia, Australia, have been major world suppliers of 

tantalum, particularly the Greenbushes and the Wodgina Mines. 

The giant (>3 km long) Greenbushes pegmatite is located 

250 km south of Perth and has been mined for tin and tantalum 

since 1888, and more recently also for kaolin and lithium 

(Partington and others, 1995). It contains the world’s largest 

identified pegmatite-hosted tantalum resource and mining 

operation. The pegmatite is late Archean in age and appears to 

have been intruded along a crustal fault zone synchronously 

with deformation. Unlike many other pegmatite districts, 

there is no evidence for a nearby parental granite pluton for 

the Greenbushes pegmatite. Four major compositional zones 

have been identified; however, the most compositionally 

evolved zones (that is, the most lithium-rich) occur at the top 

and bottom of the pegmatite rather than at the center. Early 

formed tantalum minerals are mainly wodginite and ixiolite; 

later tantalum minerals (tantalite and tapiolite) occur within 

fractures, whereas microlite is related to later hydrothermal 

mineralization (Partington and others, 1995). The main 

ore zones generally occur within the more tourmaline-rich 

portions of albite-rich zones in the pegmatite. Most ore 

produced at Greenbushes since 1888 has been by open pit 

mining of weathered pegmatite and alluvial sources; recently, 

however, underground mining of hard-rock ore has also begun.

The Wodgina pegmatite district is one of more than 

27 pegmatite districts located in the northern part of Western 

Australia, Australia; it includes the Wodgina main lode and 

Mount Cassiterite tantalum-mineralized pegmatites (Sweetapple 

and Collins, 2002). The Wodgina main lode pegmatite is a 

north-south-trending, easterly dipping (20 to 50 degrees) dike 

with a total length of about 1 km and a thickness ranging from 

5 to 40 m; most mining has been in the northern, thicker end. 

Originally discovered in 1902, the Wodgina pegmatite has 

intermittently produced beryl, tantalum, and tin during the 

past 100 years. The main tantalum mineral is manganese-rich 

tantalite, which, together with some manganese-rich columbite 

and wodginite, is found mainly in marginal cleavelandite (a 

platy form of albite) (Sweetapple and Collins, 2002). In the 

Mount Cassiterite pegmatite, which lies just south of Wodgina, 

wodginite is the main tantalum mineral (Sweetapple and 

Collins, 2002). Tantalum from both pegmatites is recovered by 

open pit mining. Additional tantalum-bearing pegmatites occur 

in Western Australia, including the producing Mount Cattlin 

pegmatite, as well as the Bald Hill and the Mount Deans 

deposits, which are under development (Fetherston, 2004).

A number of pegmatite fields are present in Africa 

(fig. M5). Pegmatites in Ethiopia (Kenticha) and Mozambique 

(Marropino) were significant tantalum producers in the 

2000–13 time period (Bleiwas and others, 2015); others were 

undergoing exploration or development for tantalum and 

associated resources (for example, the Morrua and the Mutala 

pegmatites in Mozambique). Mining of columbite-tantalite 

(also called coltan) from weathered pegmatites and from 

secondary placer deposits derived from pegmatites in central 

Africa, particularly in Burundi, the Democratic Republic 

of the Congo (Congo [Kinshasa]), Nigeria, Rwanda, and 

Uganda, has been conducted mainly by artisanal family 

mining groups and prospectors. The Main Kenticha pegmatite 

in Ethiopia is a Cambrian-age intrusion exposed across a 

length of more than 2 km (Küster, 2009). The pegmatite 

displays asymmetric internal textural and mineral zonation 

with manganese-rich tantalite as well as ixiolite concentrated 

in the upper intermediate and core zones. The deposit also 

contains valuable lithium resources, although these are not 

currently mined. Tantalum production has been from deeply 

weathered regolith over the pegmatite; however, exploration 

drilling of the hard-rock pegmatite is being conducted to 

evaluate the size of the primary mineralization (Küster, 2009). 

To the south of Kenticha, the Alto Ligonha pegmatite province 

in Mozambique contains numerous mineralized pegmatites, 

which have been mined intermittently since 1926. Tantalum 

concentrate was produced from ore from the Marropino Mine, 

concessions were held for the Morrua and Mutala deposits, 

and exploration was ongoing in adjacent areas. In Egypt, 

rare-metal granites were the focus of ongoing exploration 

and development activities, including the Abu Dabbab and 

Nuweibi granites (Küster, 2009). These deposits consist 

of fine-grained columbite-tantalite minerals disseminated 

throughout the granite intrusions. In addition, these intrusions 

host tin mineralization and ceramic-grade feldspar (Bleiwas 

and others, 2015).

In China, the Yichun tantalum-niobium-lithium deposit 

is the leading tantalum producer, accounting for more than 

one-half of China’s tantalum output (Fetherston, 2004). 

The deposit is in a small, sheet-like rare-metal granite that 

represents the most fractionated and youngest phase of the 

Jurassic Yanshan granite intrusion (Yin and others, 1995). The 

main tantalum minerals in this deposit are columbite-tantalite 

and tantalum-rich cassiterite along with minor microlite. The 

801 Mine, which is located in the eastern part of the Inner 

Mongolia Autonomous Region about 640 km northeast of 

Beijing, is reported to have large resources of niobium, REEs, 

tantalum, and zirconium (Fetherston, 2004). Additionally, 

pegmatites in the Altai region of northwestern China and in 

the Nanping region of southeastern China have also produced 

columbite-tantalite.
M20 Critical Mineral Resources of the United States—Niobium and Tantalum

In South America, the Volta Grande pegmatite mine 

located near the city of Nazareno in the State of Minas 

Gerais, Brazil, is a significant supplier of lithium, niobium, 

and tantalum. The main deposit is in a large, about 1-kmlong, subhorizontal, zoned albite-spodumene pegmatite 

body of Paleoproterozoic age (Lagache and Quéméneur, 

1997). Cassiterite, microlite, and tantalite are the main 

tantalum-niobium-bearing minerals. A distinctive feature 

of pegmatites in the Volta Grande district is that they are 

exceptionally enriched in lithium and rubidium (Lagache 

and Quéméneur, 1997).

The only pegmatite mine in North America producing 

tantalum in the 2000–13 time period was the Tanco deposit 

at Bernic Lake in southeastern Manitoba, Canada (Bleiwas 

and others, 2015). Operated briefly as a tin mine beginning in 1929, tantalum production began only in 1969. 

Full-scale mining to produce lithium from spodumene (the 

underground mine’s major product), along with cesium, 

rubidium, and tantalum, began in 1984. Tanco is part of the 

rare-metal-bearing Bernic Lake pegmatite group, which 

intrudes metamorphosed Archean volcanic rocks. The Tanco 

pegmatite occurs as a shallowly dipping sheet that is up to 

about 100 m thick and can be traced for about 1,600 m along 

strike. The pegmatite consists of nine zones, each of which has 

a different mineralogy, texture, and location, as well as a halo 

of altered mafic host rock. The deposit has a very complex 

mineralogy—more than 80 minerals are present, including 

14 tantalum-bearing minerals (Černý, 2005). Wodginite is the 

principal ore mineral, and it occurs mainly in a zone lying 

below the quartz core (Grice and others, 1972). Numerous 

other tantalum-bearing pegmatites have been identified in 

Ontario and Manitoba (Selway and others, 2005); several 

were undergoing evaluation of their tantalum and associated 

resources. In the United States, niobium-tantalum minerals, 

chiefly columbite, were recovered intermittently from a 

number of pegmatites in the past (Barton, 1962). The Harding 

pegmatite in New Mexico, which is a complexly zoned 

lithium- and beryllium-bearing pegmatite, was an important 

source of tantalum briefly during World War II and supplied 

about 3 metric tons of microlite concentrate (Parker and 

Adams, 1973). There has been no reported tantalum mine 

production in the United States since 1959.

Resources and Production

Data on global reserves and resources of niobium 

and tantalum are fragmentary and incomplete (Crockett 

and Sutphin, 1993). Identified resources of niobium and 

tantalum are large, however, and are estimated to be more 

than sufficient to meet global demand for possibly the next 

500 years (British Geological Survey, 2011). The specificity 

and accuracy of worldwide data for measured, indicated, 

and inferred resources and (or) proven and probable reserves 

vary widely. In general, the most accurate data are available 

only for individual mines or deposits, particularly those in 

Australia, Brazil, and Canada. Even in these cases, however, it 

is uncertain how reliable the national resource estimate totals 

are. In addition, some world demand for tantalum, and to a 

lesser extent, niobium, is met by production of these elements 

as byproducts of tin mining and metallurgy. Accurately 

measuring and classifying niobium and tantalum resources 

within the tin sector is difficult and has not been attempted 

(Crockett and Sutphin, 1993).

Niobium Reserves and Other Identified Resources

The current identified resources of niobium are located 

predominantly in Brazil (about 95 percent) and Canada 

(about 3.5 percent) (fig. M11) (British Geological Survey, 

2011; Papp, 2013b). Additional (though less well reported) 

resources are located in Angola, Australia, China, Greenland, 

Malawi, Russia, South Africa, and the United States. The 

largest niobium resources are hosted by carbonatite complexes 

in Brazil. Resources in alkaline granite or syenite complexes 

are widespread and can be large, but they are generally lower 

in grade than carbonatite-hosted deposits (fig. M6). In addition, these deposits tend to have more complex geology and 

0

1,000

2,000

3,000

4,000

5,000

6,000

7,000

8,000

Brazil Canada

Country

United States

Inferred resources

Reserves

Other identified 

 resources

EXPLANATION

Niobium resources, in thousand metric tons of Nb

Figure M11. Figure M11. Bar chart showing niobium (Nb) 

resources and reserves in Brazil, Canada, and 

the United States, in thousand metric tons of Nb 

(British Geological Survey, 2011, p. 18).
Resources and Production M21

mineralogy than do carbonatites, which can affect the ease 

with which the niobium could be recovered.

The United States is estimated to have about 

150,000 metric tons of niobium resources; these resources 

are associated primarily with carbonatites and alkaline 

intrusive complexes and, to a lesser extent, with pegmatites 

(Parker and Adams, 1973). The resources in the United States 

are all considered to be subeconomic at current niobium prices 

(Papp, 2013b).

Tantalum Reserves and Other Identified Resources

Tantalum resources are more widespread than those of 

niobium. Brazil has the largest share of global resources—

about 106,000 metric tons of contained tantalum, which 

accounts for approximately 40 percent (fig. M12); the 

remaining 60 percent of identified tantalum resources occur in, 

in decreasing order of tonnage, Australia (about 54,000 metric 

tons), China and Southeast Asia (about 27,000 metric tons), 

Russia and the Middle East (about 26,000 metric tons), 

central Africa (about 23,000 metric tons), other Africa (about 

17,500 metric tons), North America (about 4,500 metric tons), 

and Europe (about 1,900 metric tons). Tantalum resources in 

Brazil are principally associated with large pyrochlore deposits 

in carbonatite complexes and in the Volta Grande pegmatite 

district, whereas those in Australia are principally hosted by 

LCT-type pegmatites.

The United States has about 1,500 metric tons of 

identified tantalum resources, all of which are considered 

subeconomic (Papp, 2013c). These resources are primarily 

hosted in numerous LCT-type pegmatite districts around the 

country (Bradley and McCauley, 2013).

Undiscovered Niobium and Tantalum Resources

Additional niobium and tantalum resources are contained 

in undiscovered deposits around the world. Some of these are 

likely to occur in more remote or covered areas, potentially 

making for more challenging and expensive exploration 

and mining. Many of the 527 known carbonatite complexes 

worldwide (Woolley and Kjarsgaard, 2008) have not been 

fully explored and may contain undiscovered mineralization. 

These complexes are particularly abundant in southern 

Africa, parts of Asia, Brazil, Canada, and Russia (Woolley 

and Kjarsgaard, 2008). These same areas also commonly host 

alkaline granite and syenite complexes, many of which also 

remain to be more fully explored. A number of carbonatites 

and alkaline intrusions are currently undergoing exploration 

activity because of recent efforts to find new sources of 

REE resources (Long, 2011).

Most of the identified carbonatites in the United States 

are currently known to have only low grades of niobium 

(Berger and others, 2009). An exception may be the Elk Creek 

carbonatite in Nebraska, which has inferred niobium resources 

of 102.6 million metric tons grading 0.638 percent Nb2O5

(Daigle, 2012). Alkaline intrusions are widespread and occur 

in the Appalachian Mountains and the Ouachita Mountains, 

in west Texas, throughout the Rocky Mountain region, 

and in parts of Alaska (Barker, 1974). Additional alkaline 

granite-syenite intrusions have been identified in central 

Wisconsin and the Black Hills of South Dakota. Although 

most of these intrusions are not known to have significant 

niobium or tantalum mineralization, they have not all been 

fully explored. An exception is the Magnet Cove complex 

in southern Arkansas, which has a central carbonatite with 

niobium-rich perovskite but overall low grades (Morris, 1987). 

The Bokan Mountain alkaline granite intrusion in southern 

Alaska is being explored for REEs and uranium (Long and 

others, 2010). In addition, the Kougarok tin-bearing granite 

in the Seward Peninsula of Alaska is being evaluated for 

its tantalum resources. Archean rocks that could host giant 

LCT-type pegmatite deposits have limited occurrence in the 

United States, and are found mostly in northern Michigan, 

Minnesota, and Wyoming. Major pegmatite fields are not 

known in these areas, although exposures of rocks are limited 

in Michigan and Minnesota by glacial cover. Exploration 

of LCT-type pegmatites, such as those in the Appalachian 

Mountain and the Black Hills regions, have resulted in only 

subeconomic tantalum resources.

100

120

60

80

20

0

40

Tantalum resources, in thousand

metric tons of Ta 

Brazil

North America

Europe

Other Africa

Central Africa

Russia and Middle East

China and Southeast Asia

Australia

Identified resources

Reserves

EXPLANATION

Figure M12. Bar chart showing global tantalum (Ta) 

resources and reserves, in thousand metric tons of Ta

(British Geological Survey, 2011, p. 19).
M22 Critical Mineral Resources of the United States—Niobium and Tantalum

Niobium and Tantalum Production

Niobium

World mine production of niobium rose steadily from 

approximately 25,000 metric tons of elemental niobium in 

2000 to more than 60,000 metric tons of niobium in 2008, 

and it has remained fairly steady since then (fig. M13A). The 

increase between 2000 and 2008 is largely linked to increased 

consumption of microalloyed steel, which is used in the 

manufacture of cars, buildings, ships, and refinery equipment. 

Increased demand for these steels reflects in large part 

industrial development in such countries as Brazil, China, and 

India in that time period. In addition, increased demand for 

mobile telephones, superconducting magnets, and other hightech devices has helped spur niobium production. Brazil and 

Canada have been the leading nations that produce niobium 

mineral concentrates, with Brazil by far the leading producer 

(fig. M14A). The leading companies that produced niobium ore 

and concentrates were Companhia Brasileira de Metalurgia e 

Mineração (CBMM) (Araxá Mine), Anglo American Brazil 

Ltd. (Catalão Mine), and Companhia Industrial Fluminense 

Mineração S.A. (Volta Grande Mine) in Brazil, and IAMGOLD 

Corp. (Niobec Mine) in Canada. Other countries, including 

Burundi, China (Yichun Mine), Ethiopia (Kenticha Mine), 

Nigeria, Russia (Lovozero Mine), Rwanda, and Uganda, also 

produced niobium mineral concentrates in 2011.

The leading niobium producers process the ore to make 

ferroniobium (FeNb) or niobium pentoxide (Nb2

O5), which is 

then shipped to consumers. Other niobium mining operations 

ship niobium mineral concentrates to processing companies. 

The ore is crushed and ground, beneficiated by flotation and 

magnetic separation (to remove iron minerals) to produce 

a pyrochlore mineral concentrate grading 55 to 60 percent 

Nb2O5. Columbite and tin slags account for a minor amount of 

niobium production. Artisanal mining of columbite-tantalite 

(coltan) is carried out in central Africa, particularly in 

Burundi, Congo (Kinshasa), Nigeria, Rwanda, and Uganda.

Tantalum

Global tantalum mine production peaked in 2004 at 

just greater than 1,400 metric tons of elemental tantalum 

and mostly declined since 2006 to 770 metric tons in 2011 

(fig. M13B). The decrease in tantalum production reflects the 

drop in tantalum prices that accompanied the global economic 

downturn and the increased volume of tantalum coming out 

of areas of armed conflict in Congo (Kinshasa). Australia and 

Brazil have been the leading national producers of tantalum 

mineral concentrates (fig. M14B), although since 2009, 

production from Australia has decreased, and production from 

Mozambique and some other countries has increased (Papp, 

2013a). The leading companies producing tantalum mineral 

concentrates in 2011 were Companhia Industrial Fluminense 

Mineração S.A. (Volta Grande Mine) and Mineração Taboca 

S/A (Pitinga Mine) in Brazil and Noventa Ltd. (Marropino 

Mine) in Mozambique. The leading tantalum producer in other 

years was Global Advanced Metals from its Greenbushes 

and Wodgina Mines in Western Australia. Other countries, 

including Burundi, Canada (Tanco Mine), China (Yichun 

Mine), Congo (Kinshasa), Ethiopia (Kenticha Mine), Nigeria, 

Russia (Lovozero Mine), Rwanda, and Uganda also produced 

tantalum mineral concentrates in 2011.

About one-half of all tantalum production consists of 

tantalite ore, and the remainder comes from tin slag, from such 

other minerals as strüverite and columbite-tantalite, and from 

recycling and synthetic concentrates. Tin slag is produced 

primarily in Southeast Asia, Australia, and Brazil. Tantalum is 

mined from both open pit and underground operations. Heavy 

minerals that contain tantalum are separated from the bulk 

ore by gravity methods and flotation. The concentrate is then 

subject to electrical (electrostatic and [or] electromagnetic) 

separation and other methods used to separate tantalum 

minerals from other heavy minerals. Because most mining 

of tantalum recovers minerals that also contain niobium, a 

chemical processing step involving the addition of potassium 

fluoride is necessary to separate niobium from tantalum, 

resulting in compounds of Nb2

O5, Ta2O5, potassium niobium 

fluoride (K2

Nb2F5), and potassium tantalum fluoride (K2Ta2F5). 

The tantalum compounds are then smelted to make tantalum 

metal products (Roskill Information Services Ltd., 2012). The 

global supply of tantalum historically has been 70 percent 

from concentrates, 10 percent from tin slag, and 20 percent 

from recycling and synthetic concentrates (Schwela, 2010). 

Conventional mining accounted for 51 percent of the world’s 

tantalum supply in 2011, artisanal mining accounted for 

29 percent, and tin slags and synthetic concentrates accounted 

for 10 percent (Roskill Information Services Ltd., 2012).
Resources and Production M23

A

B

Niobium production, in metric tons Tantalum production, in metric tons

0

10,000

20,000

30,000

40,000

50,000

60,000

70,000

2000 2001 2002 2003 2004 2005 2006

Year

Figure M13

2007 2008 2009 2010 2011

0

600

400

200

1,000

800

1,600

1,400

1,200

Figure M13. Bar charts showing global mine production of A, niobium, and B, tantalum, from 2000 to 

2011, in metric tons of contained niobium and tantalum (Papp, 2013a).

Brazil

90%

Canada

8%

other

1%

Brazil

90%

Canada

8%

Other 1%

Brazil

18%

Other

14%

A. Niobium production B. Tantalum production

Australia

42%

Rwanda

8%

Mozambique

9%

Ethiopia

6%

Canada

4%

Figure M14. Pie charts showing percentage of average annual world production of A, niobium, and 

B, tantalum, for the period 2007–11, by country (Papp, 2013a). Australia and Canada had no reported 

tantalum production in 2010. The data for Ethiopia are for 2007 only. Percentages have been rounded 

to no more than two significant figures and do not add to 100.
M24 Critical Mineral Resources of the United States—Niobium and Tantalum

Exploration for New Deposits

Carbonatites are an important exploration target because 

of their potential to host large, commonly bulk-minable 

niobium (with lesser amounts of tantalum) resources. About 

80 percent of carbonatites are associated with silicate rocks, 

the majority of which are highly alkaline (Woolley and 

Bailey, 2012). Not all alkaline complexes and provinces 

contain carbonatites, however. Carbonatites are preferentially 

concentrated in Precambrian terranes and are associated with 

intracontinental rifts, such as the East African rift system 

(Bailey, 1974). Their common circular plan and concentric 

arrangement of alkaline rocks can be identified and mapped 

by standard geologic-mapping techniques in areas with 

surface rock exposures. In remote areas and covered areas, 

geophysical surveys, including magnetic, gravity, and 

electrical methods, play a critical role in exploration. The 

natural association of carbonatites with uranium, thorium, and 

potassium also makes airborne radiometric surveys a useful 

exploration tool. Once a target has been identified, conventional exploration procedures are used in evaluating niobium 

and tantalum resources, including detailed ground geophysical 

surveys, and trenching, drilling, and chemical analysis. 

Niobium and tantalum minerals typically are not easily 

identified and can require laboratory analysis for confirmation. 

In addition, the insoluble and refractory character of niobium 

and tantalum can require special analytical methods.

LCT-type pegmatites generally occur along large 

regional-scale faults in low to moderately metamorphosed 

terranes with mafic volcanic and sedimentary host rocks 

(Selway and others, 2005). The largest deposits are in Archean 

crust. Initial exploration typically involves a search for 

rare-metal peraluminous granite intrusions that can serve as 

the source of pegmatites. Rare-metal granites are identified 

by (a) high rare-metal contents (which are at least three times 

the average crustal abundances); (b) magnesium to lithium 

(Mg:Li) ratios less than 10; and (c) niobium to tantalum 

(Nb:Ta) ratios less than 8 (Selway and others, 2005). These 

granites are also characterized by white potassium feldspar, 

green coarse-grained mica, and distinctive accessory minerals 

(such as cordierite, fluorapatite, garnet, and (or) tourmaline). 

Rare-metal pegmatites may be found in an aureole above 

the granite, with the most enriched pegmatites generally 

the farthest (up to 10 km) from the granite body. Altered 

country rocks can be an indication of rare-metal pegmatite 

bodies nearby and may contain elevated contents of boron, 

cesium, fluorine, lithium, and rubidium that are detectable 

in geochemical surveys. Pegmatites with potential to host 

a tantalum deposit generally have low Nb:Ta ratios (< 8) 

and tantalum to tin (Ta:Sn) ratios of >1 in bulk analyses, 

as well as abundant tantalum-rich oxide minerals. As with 

carbonatites, radiometric surveys can aid in exploration for 

LCT-type pegmatites.

Environmental Considerations

Sources and Fate in the Environment

The main sources of niobium and tantalum in the environment are geologic, mostly as a result of the weathering of 

rocks, but a potential anthropogenic source of both is from 

the combustion of coal (Divine and Goering, 2004; Goering 

and Ziegler, 2004). Tantalum compounds are less mobile than 

niobium compounds, which are slightly soluble under both 

acid and alkaline conditions and in the presence of organic 

complexing agents (Kabata-Pendias and Mukherjee, 2007). 

Gaillardet and others (2003) reported worldwide average 

dissolved concentrations in river waters of 2 parts per trillion 

(ppt) niobium and 1 ppt tantalum. Globally, the average 

concentration in suspended sediments in rivers is 13.5 ppm 

for niobium and 1.3 ppm for tantalum (Viers and others, 

2009). Also, surface (0 to 5 centimeters depth) soils from the 

conterminous United States contain an average of 9.3 ppm 

niobium and range from less than 1 ppm to 80.1 ppm; values 

for tantalum were not reported (Smith and others, 2013).

The concentrations of some elements may be naturally 

enriched in soils, sediments, and waters in the vicinity 

of niobium and tantalum deposits before mining begins; 

knowledge of these geochemical anomalies is used as an 

exploratory tool as well as to set realistic remediation goals 

after mining. The geochemical signatures around the deposits 

depend on the type of deposit and the compositions of 

associated rocks. Stream water and groundwater near deposits 

may not display the same geochemical signature as soils and 

stream sediments because some of the enriched elements 

occur in relatively stable insoluble minerals. In general, 

soils derived from various alkaline rock types commonly 

contain increased levels of niobium (up to 300 ppm) 

(Ure and others, 1979). The niobium in the soils can be 

taken up by plants, such as in the Gumi region of Russia 

where the niobium content (up to 10 ppm) of certain plant 

species has served as an exploration indicator for niobium 

deposits (Kabata-Pendias and Pendias, 2001). Soils and 

weathered rock associated with carbonatite deposits may be 

naturally enriched in Fe, Ba, Nb, P, Th, Ti, U, Zr and REEs. 

Additionally, stream sediments downstream from deposits 

commonly contain anomalous abundances of Ba, Nb, Th, 

and REEs (Modreski and others, 1995). LCT-type pegmatite 

deposits may exhibit enrichments in B, Cs, F, Li, and Rb in 

surrounding soils and waters because these elements tend to 

alter country rock during pegmatite emplacement (Bradley 

and McCauley, 2013). For example, the Tanco pegmatite 

in Canada has alteration halos of Li and Rb over the buried 

deposit (Trueman and Černý, 1982). In addition, Smith and 

others (1987) reported that zones of enriched As, Be, Sb, and 

Sn, and of B, Nb, and Ta are found in lateritic soils around 

the Greenbushes pegmatite.
Environmental Considerations M25

Globally, niobium metal is one of 18 elements with a 

greater than 50 percent end-of-life recycling rate, which is the 

proportion of the element in a discarded product that is able 

to be used to generate a new product. In contrast to niobium, 

less than 1 percent of discarded tantalum metal or alloy is 

used to generate a new product worldwide. Tantalum is used 

in small quantities in complex products, such as electronics, 

which makes recycling technically difficult. To recover more 

tantalum from consumer products, additional separation in 

the collection and presorting of the electronics would need to 

be done (Graedel and others, 2011). The recycled content of 

products containing niobium and tantalum is high, at greater 

than 50 percent for niobium and between 10 and 25 percent 

for tantalum (Graedel and others, 2011). According to 

Papp (2013a), scrap recovery specifically for niobium was 

negligible, but significant amounts of niobium were recycled 

from niobium-bearing steels and superalloys. Tantalum was 

recycled mostly from new scrap generated during the manufacturing of cemented carbide, electronics, and superalloys.

Mine Waste Characteristics

Mine waste is generally considered to be the material 

generated during ore extraction and processing at the mine site 

that has no economic value. The volume and characteristics of 

the waste materials depend on the type of deposit and its size, 

the mineral commodity or commodities recovered, and the 

mining and processing methods used. Niobium and tantalum 

ores are treated similarly and commonly are crushed and 

ground before mineral separation, which may involve dry and 

wet gravity separation, dry screening, magnetic separation, 

flotation, acid leaching, and, sometimes, smelting, to remove 

impurities and concentrate the niobium and (or) tantalum. 

The removal of overburden and ore extraction and processing 

operations generate waste rock, tailings, and slag, which are 

commonly disposed of on site in dumps or in tailings ponds. 

The amount of waste produced at the mine sites may be significant relative to the amount of niobium or tantalum recovered 

because of the low grades of the ores; however, the recovery 

of additional commodities can reduce the waste volume.

The mining and processing of niobium and tantalum 

deposits commonly generates various types of waste, some 

with unique compositions. One example of mine waste 

generated from an LCT-type pegmatite deposit is waste from 

the Greenbushes Mine in Australia where the various ore 

minerals are extracted and processed separately to recover 

either lithium or tin and tantalum. Processing the tin-tantalum 

ores generates tailings composed mainly of gangue minerals 

(that is, predominantly quartz and feldspars). Minor amounts 

of antimony and sulfides are removed by flotation and 

roasting, and smelting is done to separate the tantalum from 

the tin (Fetherston, 2004); these processes create tailings 

and slag waste. In addition, overburden rock consisting 

predominantly of quartz (35 to 45 percent) and kaolinite 

(35 to 60 percent) is used to cover waste rock dumps to aid 

in rehabilitation (Hearman and Hinz, 2004).

Another example of the variety of waste generated 

from processing LCT-type pegmatite ore is the waste from 

the underground Tanco Mine in Manitoba, Canada. At 

this site, tantalum is processed by gravity concentrations, 

spodumene is processed mainly by flotation, and pollucite is 

treated chemically. The wastes generated include tailings and 

waste rock that are generally composed of the predominant 

gangue minerals of feldspar and quartz. For comparison to 

the LCT-type pegmatite deposits, mine waste generated from 

exploiting niobium-enriched carbonatites is usually some 

combination of the following common gangue minerals: 

amphiboles, ankerite, apatite, calcite, dolomite, magnetite, 

micas, pyroxenes, and olivine, among others. Mine waste 

from one processing step at the Araxá Mine in Brazil is 

predominantly hematite and magnetite; trace amounts of 

anatase, apatite, barite, and sulfides are likely present (Silva 

and others, 2012). Similar to ore from the Araxá Mine, 

niobium ores from the Niobec Mine in Quebec, Canada, are 

extracted and processed in several steps, and each generates 

waste of a different composition. Flotation and desliming 

generates carbonate-rich tailings; magnetic separation generates magnetite waste; and additional flotation generates a 

sulfide waste containing pyrite (Fetherston, 2004).

In general, tailings and waste produced from niobium 

and tantalum deposits may contain significant concentrations 

of uranium and (or) thorium and their daughter radionuclides. 

For example, tailings from the Jos area of Nigeria contain 

up to 4 weight percent thorium and up to 0.2 weight percent 

uranium (Funtua, 2001). The smelting and some chemical 

processes (acid leaching) of niobium ores from two operations 

in Brazil significantly enrich the waste in radionuclides; 

however, the prior physical processing (that is, magnetic 

separation, desliming, and flotation) of these particular ores 

does not significantly alter the radionuclide content of the 

waste compared with that of the ore (Pires do Rio and others, 

2003). Also, the chemical treatment of phosphate ores from 

carbonatite deposits in Brazil (for example, Araxá, Catalão, and 

Tapia) generates waste with high radioactivity (concentrations 

of up to hundreds of kilobecquerels per kilogram of radium 

as 226Ra) (Paschoa, 1998). Although the activities of uranium 

and thorium in ore from the Greenbushes Mine in Australia are 

generally low (less than 60 becquerels per kilogram (Bq/kg) 

of uranium and less than 5 Bq/kg of thorium), processing the 

ore may generate enriched waste and concentrates, such that 

tantalum products contain from 7.5 to 75 kilobecquerels per 

kilogram (kBq/kg) of uranium and thorium (Cooper, 2005). 

The volume of radionuclide-enhanced waste generated at the 

mine sites can be significant, such as at the Pitinga tintantalum-niobium mine in Brazil, where at least 10,000 metric 

tons of naturally occurring radioactive wastes are disposed of 

every year (Paschoa, 1998).
M26 Critical Mineral Resources of the United States—Niobium and Tantalum

Human Health Concerns

Niobium and tantalum are generally considered 

nonessential for humans and have low toxicity (Reimann and 

de Caritat, 1998). Because of their relative biological inertness, both, but in particular tantalum, are used in a variety of 

medical devices and surgical implants (Divine and Goering, 

2004; Kabata-Pendias and Mukherjee, 2007). Tantalum is 

absorbed only minimally by the human body from oral or 

inhalation exposure; nearly all the tantalum that enters the 

human body is eliminated within a relatively short period of 

time. Humans exposed to large amounts of tantalum metal 

or to its oxide dusts may experience eye and skin irritation. 

Also, there is some evidence that tantalum can cause tumors 

(Divine and Goering, 2004). The permissible exposure limits 

in the United States for tantalum metal and oxide dust for an 

8-hour work day is 5 milligrams per cubic meter (mg/m3) and 

is based on lung irritation (Occupational Safety and Health 

Administration, 2013). Although data are sparse and studies on 

humans lacking, niobium is likely similar to tantalum in that 

its retention time in the human body is usually relatively short, 

and much of it is eliminated unabsorbed (Goering and Ziegler, 

2004). Studies have shown, however, that small mammals 

that ingested niobium compounds in drinking water had a 

shortened lifespan. Niobium has been shown to be associated 

with tumors, but is not mutagenic (Goering and Ziegler, 2004, 

and references therein). Niobium is considered hazardous by 

the Occupational Safety and Health Administration (OSHA), 

but OSHA has not established a permissible exposure 

limit for niobium dust (Occupational Safety and Health 

Administration, 2013).

Although niobium and tantalum may not be the cause 

of significant human health concerns, industrial exposure to 

naturally occurring radioactive materials from enhanced levels 

of thorium and uranium during the processing of niobium 

and tantalum ores may be a concern. Occupational exposure 

can take place from external radiation and from inhaling 

and ingesting radioactive particles generated during mining 

and milling. Smokers, who inhale twice or more the amount 

of radon than do nonsmokers, have an increased risk for 

exposure (Mangset and Sheyin, 2009). In Congo (Kinshasa), 

the grinding and sieving of columbite-tantalite ores during 

artisanal mining has been shown to expose workers to high 

doses of radioactivity (Mustapha and others, 2007).

Humans living near mining and processing operations 

may also be exposed to technically enhanced naturally 

occurring radioactive materials. Exposure may be a concern 

from waste materials used for building construction and from 

drinking water or eating foods contaminated with radiation. 

Slag and acid-leaching wastes (mostly barium sulfate) 

disposed of in alternating layers in waste dumps at a niobium 

processing facility in Brazil yielded higher radionuclide 

concentrations than did ores and other waste materials, and 

doses associated with them exceed international exposure 

guidelines. Also, the doses of radiation from treated effluent 

and drainage from waste ponds were lower than local 

limits but higher than some international exemption levels 

(Pires do Rio and others, 2003). In Nigeria, mining tin, 

tantalum, and niobium from placer deposits and weathered 

granites has resulted in large amounts of radioactive tailings 

dumped in and around mine sites (Funtua, 2001). The 

activities of radionuclides of radium and thorium in tin 

tailings, especially those containing significant monazite and 

zircon, and soils 500 m from the tailings were higher than 

background activities, and the estimated annual effective 

doses to the population were much higher than the recommended levels (Usikalu and others, 2011). In addition to being 

radioactive, tin-niobium-tantalum tailings from the Jos area 

of Nigeria contained concentrations of lead, manganese, and 

tin that exceeded soil screening levels based on residential 

and industrial use in the United States; these levels are used 

to determine whether contamination at hazardous waste 

sites warrant further investigation or site cleanup (Funtua, 

2001; U.S. Environmental Protection Agency, 2012). After 

operations have ended at niobium and tantalum mines and 

processing sites, the use of the land may need to be restricted 

to minimize human exposure to radiation and high concentrations of some trace elements.

Ecological Health Concerns

Ecological health concerns associated with niobium 

and tantalum are generally minimal under most naturally 

occurring conditions. Niobium may accumulate in plants 

and is generally less than 0.4 ppm dry weight in terrestrial 

plants from background locations and greater than 1 ppm in 

plants growing near niobium deposits (Goering and Ziegler, 

2004). According to one recent study, the toxicity of niobium 

and tantalum to freshwater amphipods (Hyalella azteca) was 

relatively low compared with that of other trace elements 

tested in laboratory experiments; 1-week lethal concentrations 

of hundreds of ppb niobium and tantalum resulted in significant mortality for amphipods in soft waters, whereas those 

in hard waters were not affected until concentrations reached 

thousands of ppb (Borgmann and others, 2005).

The mining and processing of niobium and tantalum 

ores may have a deleterious effect on terrestrial and aquatic 

plants and organisms from the release of trace elements and 

(or) radionuclides associated with the ore or host rock into 

the surrounding ecosystem. If sulfides are present, even in 

minor to trace quantities, they can oxidize, generate acid, and 

exacerbate the release of trace elements; however, rocks that 

contain carbonates can help mitigate some deleterious effects 

naturally. Metal-rich and radioactive mine waste can be 

physically transported into local waterways, and fine-grained 

tailings dust and emissions from ore processing sites can be 
Environmental Considerations M27

transported in the air for long distances; these processes can 

result in contaminated soils, waters, and stream sediments. The 

mining and processing of the ore may take place at the same 

location, and emissions from the processing and smelting may 

contain toxic elements and substances that can be released into 

the air and surrounding soils and waterways.

A few studies of niobium and tantalum mine sites have 

reported potential ecological health concerns. In Rwanda, the 

concentrations of arsenic in stream sediments downstream of 

Ruhanga and at the Nyatubindi tin-niobium-tantalum mine 

and lead in stream sediments at the Ruhanga site were above 

the threshold effect concentrations for aquatic organisms 

(that is, the concentrations below which effects are not likely 

to occur) but below probable effect concentrations (that is, the 

concentrations above which effects are likely to occur), which 

suggests potential impact to biota (MacDonald and others, 

2000; Haidula and others, 2011). Also, rare but extreme mine 

water compositions existed within the mine workings at the 

Lovozero Mine. Waters were highly alkaline (pH range of 

10 to 12), enriched in fluorine (1 to 1.5 percent) and silica 

(1 to 1.3 percent), and contained high amounts of aluminum 

(up to 1,000 ppm), copper (up to 900 parts per billion [ppb]), 

niobium (up to 400 ppb), and REEs (greater than 600 ppb); the 

concentrations of aluminum and copper exceeded guidelines 

for aquatic ecosystem health in the United States (Kraynov 

and others, 1969; U.S. Environmental Protection Agency, 

2013). In addition to these studies, Paschoa (1998) reported 

significant radioactivity (approximately 100 kBq/kg of 226Ra) 

for fine-grained tailings at the Araxá Mine in Brazil. Tailings 

are disposed of in waste ponds, thus contaminating the water 

(approximately 20 kilobecquerels per cubic meter of 226Ra) 

(Paschoa, 1998). Concentrations above 1 to 10 kBq/kg, 

depending on the radionuclide, warrant special consideration 

for disposal of solid material, which may need to be buried in 

low-level radioactive waste disposal facilities. Beside reports 

of radioactivity and some high concentrations of elements, 

mine waste and mine waters generated at some niobium and 

tantalum mine sites are not of ecological health concern. For 

example, processing water and tailings runoff and discharge 

released into the environment at the world’s largest pegmatitehosted tantalum resource, the Greenbushes Mine, is monitored 

to ensure compliance with internationally developed 

standards (Ingham and others, 2012). The mine waste generated from pegmatite deposits, although possibly not an issue 

chemically, may pose ecological health concerns because of 

the extensive amount of waste disposed of at the sites.

Carbon Footprint

The carbon footprint of niobium centers on its use in 

the production of biofuels. Niobium catalysts have recently 

proved to be successful in the conversion of palm oil to 

biodiesel. In Brazil, one large mining company has developed 

palm oil plantations to supply its own biodiesel needs 

(Schwela, 2010). The use of biofuels decreases greenhouse 

gas emissions compared with the use of petroleum fuels; 

however, whether first-generation biofuels have a less 

substantial carbon footprint than petroleum when other 

factors are taken into consideration remains controversial. 

Such other factors include the amount of energy consumed to 

convert plants to biofuel, the opportunity cost of using land 

to grow crop plants for biofuels that is then not available to 

grow food for humans and animals, and the potential increase 

in the amount of forest land that may be cleared to grow 

biofuel crops.

Mine Closure

The postmining landscape at niobium and tantalum 

mines depends on the type and size of the deposit, the mining 

and processing methods used, the mineral commodity or 

commodities recovered, the climate, and the relevant regulations regarding mine closure. Extraction from large igneous 

complexes by open pit methods, which is the most common 

mining method used, will leave a distinctly different signature 

than recovery from small artisanal mining of secondary 

placer deposits or weathered pegmatites. For example, the 

active Greenbushes Mine in Australia, which has been mined 

intermittently since 1888, consists of a complex landscape of 

several open pits, underground workings, waste rock piles, 

and tailings piles (Ingham and others, 2012). The largest 

open pit is the Cornwall Pit, which is more than 200 m deep. 

The closure cost for this site was estimated to be more than 

$17 million to rehabilitate nearly 2,000 hectares of disturbed 

land (Ingham and others, 2012).

Artisanal and small-scale mining, such as that conducted 

in central Africa, is typically unregulated, and environmental 

issues are commonly not considered. Many such mining 

operations do not have mine closure plans and are left as 

is after recoverable ore is extracted. For example, tin (and 

tantalum-niobium) mining in the Jos area of Nigeria has 

drastically changed the landscape, and little has been done 

in the way of remediation efforts. Some detrimental effects 

of the altered landscape include disturbances of the natural 

flora and fauna, low crop yields on agricultural lands, local 

land disputes from soil erosion problems, and neglected 

mine excavations, mine dumps, and pits. On the other 

hand, flooded mine pits are a source of water for irrigation 

as well as for domestic and livestock uses. They can also 

provide sites for potential fisheries (Abegboye, 2012). Future 

measures may focus on remediating the affected land back 

to productive farm land.
M28 Critical Mineral Resources of the United States—Niobium and Tantalum

Problems and Future Research

Consumption of niobium and tantalum is expected to 

increase in the coming years as industrial development in such 

countries as Brazil, China, and India continues, and the rapidly 

increasing global population demands more consumer goods, 

such as cars, cell phones, computers, and other high-tech 

products. Although the estimated reserves and resources of 

both niobium and tantalum are large and appear more than 

sufficient to meet global demand for the foreseeable future 

(British Geological Survey, 2011), the restricted concentration 

of niobium resources and production, and, to a lesser extent, 

those of tantalum, makes their stable supply vulnerable to 

potential influence and disruption. The United States currently 

lacks domestic self-sufficiency for both mineral commodities 

and is dependent on imports and recycling to meet its current 

and future needs. For the United States, the discovery of 

additional deposits of niobium and tantalum of better quality 

and quantity than those currently identified would be advantageous. New discoveries in the United States, particularly of 

niobium, may be facilitated by ongoing efforts to identify 

additional REE resources because alkaline and carbonatite 

complexes can host both deposit types.

An issue of international concern specific to tantalum 

is the mining of columbite-tantalite (also called coltan) in 

central Africa, particularly in Congo (Kinshasa). Militias 

implicated in human rights abuses are accused of using the 

sale of columbite-tantalite and other mineral commodities 

to help fund a civil war in Congo (Kinshasa). In 2010, 

the U.S. Government passed the Dodd-Frank Wall Street 

Reform and Consumer Protection Act. Section 1502 of that 

Act directs that companies that report to the U.S. Securities 

and Exchange Commission must reveal whether they buy 

minerals from Congo (Kinshasa) or any of its neighbors and, 

if so, whether conflict minerals exist in their supply chain. 

This has led to efforts to develop a system that can trace the 

origin of columbite-tantalite ore from the region (Burt, 2012). 

A program related to establishing a conflict-free supply chain 

for columbite-tantalite was launched by the International Tin 

Research Institute in 2011 in Rwanda and the Katanga Province 

of Congo (Kinshasa), with baseline studies of almost 500 mine 

sites completed (Burt, 2012). This program (the Tin Supply 

Chain Initiative, or iTSCi) is attempting to identify legitimate 

mines, with the overall goal being to provide a way to exclude 

illegal columbite-tantalite from the conventional market while 

discerning legitimate small-scale artisanal mine production in 

central Africa.

Geologically, there are still many unanswered questions 

about the origin of niobium and tantalum deposits, including 

(a) why some carbonatites are enriched in niobium and others 

are enriched in REEs; (b) what is the role of fluids in the mobilization and concentrations of niobium and tantalum; (c) why 

are the largest pegmatite-hosted tantalum deposits of Archean 

age; and (d) are there additional types of tantalum deposits, 

particularly low-grade large-tonnage commercial deposits?

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Platinum-Group Elements

By Michael L. Zientek, Patricia J. Loferski, Heather L. Parks, Ruth F. Schulte, 

and Robert R. Seal II

Chapter N of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–N

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter N

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................N1

Introduction..................................................................................................................................................N2

Uses, Applications, and Consumption ............................................................................................N2

Geology.........................................................................................................................................................N4

Geochemistry......................................................................................................................................N4

Mineralogy..........................................................................................................................................N5

Deposit Types .....................................................................................................................................N5

Magmatic Deposits ..................................................................................................................N5

Tectonic Setting of Magmatic Deposits (Large Igneous Provinces).......................N8

Types of Magmatic Deposits ..........................................................................................N8

Conduit-Type Deposits ...........................................................................................N8

Reef-Type and Contact-Type Deposits ..............................................................N12

Hydrothermal and Sedimentary Deposits ...........................................................................N18

Residual Deposits Formed By Weathering and Laterites ................................................N20

Platinum Placer Deposits .....................................................................................................N22

Resources and Production ......................................................................................................................N28

Production.........................................................................................................................................N28

Identified Resources .......................................................................................................................N29

Tonnage and Grade Relations for Magmatic Ore Deposits ......................................................N30

Undiscovered (Hypothetical and Speculative) Resources (United States and Global) .......N32

Extensions to Identified Deposits .........................................................................................N32

Undiscovered Reef-Type and Contact-Type Deposits ......................................................N34

Undiscovered Conduit-Type Deposits in the United States—Nikolai Greenstone 

and Eastern Wrangellia ............................................................................................N35

Undiscovered Platinum Placer Deposits in Alaska ...........................................................N35

Metals in Stored Waste Products .................................................................................................N35

Exploration for New Deposits .................................................................................................................N37

Environmental Considerations ................................................................................................................N37

Sources and Fate in the Environment ..........................................................................................N38

Platinum-Group-Element Mining and Mineral-Processing Methods .....................................N39

Mine Waste Characteristics ..........................................................................................................N40

Acid-Base Accounts ..............................................................................................................N40

Acid-Base Accounts for Reef-Type and Contact-Type Deposits ...........................N40

Acid-Base Accounts for Conduit-Type Deposits ......................................................N41

Air Quality Assessments ........................................................................................................N41

Human Health Concerns .................................................................................................................N41

Ecological Health Concerns ...........................................................................................................N41

Carbon Footprint...............................................................................................................................N42

Mine Closure.....................................................................................................................................N42

Problems and Future Research ..............................................................................................................N43

Future Demand and Possible Risks to Supply .............................................................................N43

Research Directions ........................................................................................................................N45

Acknowledgments....................................................................................................................................N45

References Cited.......................................................................................................................................N45

Suggested citation:

Zientek, M.L., Loferski, P.J., Parks, H.L., Schulte, R.F., and Seal, R.R., II, 2017, Platinum-group elements, chap. N of

Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—

Economic and environmental geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802,

p. N1– N91, http://dx.doi.org/10.3133/pp1802N.
Platinum-Group Elements

By Michael L. Zientek, Patricia J. Loferski, Heather L. Parks, Ruth F. Schulte, and Robert R. Seal II

Abstract

The platinum-group elements (PGEs)—platinum, palladium, rhodium, ruthenium, iridium, and osmium—are metals 

that have similar physical and chemical properties and tend 

to occur together in nature. PGEs are indispensable to many 

industrial applications but are mined in only a few places. 

The availability and accessibility of PGEs could be disrupted 

by economic, environmental, political, and social events. The 

United States net import reliance as a percentage of apparent 

consumption is about 90 percent.

PGEs have many industrial applications. They are used in 

catalytic converters to reduce carbon monoxide, hydrocarbon, 

and nitrous oxide emissions in automobile exhaust. The 

chemical industry requires platinum or platinum-rhodium 

alloys to manufacture nitric oxide, which is the raw material 

used to manufacture explosives, fertilizers, and nitric acid. In 

the petrochemical industry, platinum-supported catalysts are 

needed to refine crude oil and to produce aromatic compounds 

and high-octane gasoline. Alloys of PGEs are exceptionally 

hard and durable, making them the best known coating for 

industrial crucibles used in the manufacture of chemicals and 

synthetic materials. PGEs are used by the glass manufacturing 

industry in the production of fiberglass and flat-panel and 

liquid crystal displays. In the electronics industry, PGEs are 

used in computer hard disks, hybridized integrated circuits, 

and multilayer ceramic capacitors.

Aside from their industrial applications, PGEs are used 

in such other fields as health, consumer goods, and finance. 

Platinum, for example, is used in medical implants, such as pacemakers, and PGEs are used in cancer-fighting drugs. Platinum 

alloys are an ideal choice for jewelry because of their white 

color, strength, and resistance to tarnish. Platinum, palladium, 

and rhodium in the form of coins and bars are also used as 

investment commodities, and various financial instruments

based on the value of these PGEs are traded on major exchanges.

PGEs are among the rarest metals; Earth’s upper crust 

contains only about 0.0005 part per million (ppm) platinum. 

Today, the average grade of PGEs in ores that are mined 

primarily for their PGE concentrations varies from 5 to 15 ppm, 

although the concentration of PGEs in hand-picked ore specimens may range from tens to hundreds of parts per million.

More than 100 different minerals have one of the PGEs as 

an essential component. PGE minerals occur as native metals. 

They also occur as compounds with other transition metals 

(copper, iron, mercury, nickel, and silver), post-transition 

metals (bismuth, lead, and tin), metalloids (antimony, arsenic, 

and tellurium), and nonmetals (selenium and sulfur).

From 1900 to 2011, approximately 14,200 metric tons of 

PGEs was produced, and roughly 95 percent of that production 

(13,500 metric tons) took place between 1960 and 2011. The 

breakdown of production by country shows that, since 1900, 

about 90 percent of the production came from South Africa 

and Russia. The secondary supply of platinum, palladium, 

and rhodium is obtained through the recycling of catalytic 

converters from end-of-life vehicles, jewelry, and electronic 

equipment. Recycled platinum, palladium, and rhodium 

provide a significant proportion of the world’s total supply; 

these secondary sources are sufficient to close the gap between 

world mine production and consumption.

Exploration and mining companies report resources of 

about 104,000 metric tons of PGEs (including minor amounts 

of gold) in mineral deposits around the world that could be 

developed. For PGEs, almost all the reported production 

and identified resources are associated with deposits in three 

geologic features—the Bushveld Complex, which is a layered 

mafic to ultramafic intrusion in South Africa; the Great Dyke, 

which is a layered mafic to ultramafic intrusion in Zimbabwe; 

and sill-like intrusions associated with flood basalts in the 

Noril’sk-Talnakh area of Russia.

The metallic forms of PGEs are generally considered to 

be inert. PGEs pose a risk to human health only in cases where 

individuals are occupationally exposed to synthetic PGE 

compounds, especially workers in precious-metal refineries. 

In the natural environment, background PGE concentrations 

are low in water, sediment, soil, and plants. Anthropogenic 

sources of PGEs in the environment include catalytic 

converters used in modern automobiles, platinum-based 

chemotherapy drugs, and smelter emissions.

The abundance of sulfide minerals defines the environmental and geologic characteristics of PGE-enriched magmatic 

sulfide deposits; those deposits with the highest amount of 

sulfide minerals could have the highest environmental impact. 

Acid rock drainage from reef-type and contact-type deposits 
N2 Critical Mineral Resources of the United States—Platinum-Group Elements

is unlikely because the ores and their host rocks contain 

low proportions of sulfide minerals. For some conduit-type 

orebodies with massive ores, mineral-processing techniques 

separate and produce concentrates of copper-, iron-, and 

nickel-bearing sulfide minerals; those with copper and nickel 

are processed to extract metal, but the iron-sulfide minerals, 

mainly pyrrhotite, are discarded as waste. This results in waste 

material with a high acid-generating potential.

The most significant primary source of PGEs in the 

United States is a deposit in the Stillwater Complex, which 

is a layered igneous intrusion in Montana. Approximately 

305 metric tons of platinum and palladium have been 

mined from the Stillwater Complex deposit since 1986. 

Exploration and development drilling indicate that another 

2,200 metric tons are present. Mining has progressed to depths 

of 1,800 meters below the surface, but the bottom of the ore 

deposit has not been reached; geologic estimates suggest that 

another 1,000 to 6,200 metric tons of PGEs could be present at 

depth. In the future, PGEs may be mined from deposits found 

near the base of the Duluth Complex, which is a group of 

igneous intrusions in Minnesota.

Introduction

The platinum-group elements (PGEs)—platinum, 

palladium, rhodium, ruthenium, iridium, and osmium—are 

chemical elements that have similar physical and chemical 

properties and tend to occur together in nature. They are 

transition metals, lying in Group 8B and periods 5 and 6 of the 

periodic table of elements. Along with gold and silver, they are 

also precious metals that have high economic value determined not only by their practical uses but also by their role as 

investment instruments and as a store of value. The PGEs are 

also known as the platinum-group metals (PGMs).

Pre-Columbian peoples found naturally occurring 

platinum and platinum-rich alloys in stream deposits in 

Colombia and Ecuador. In the 1500s, Spanish colonists also 

found grains of these alloys mingled with gold nuggets they 

recovered from stream deposits in the same area; they called 

the metal “platina” (McDonald and Hunt, 1982). At that time, 

the metal had no known use and was considered worthless. 

The platinum-rich grains were a nuisance because they could 

not be separated from gold nuggets by panning. Small samples 

of platinum-enriched nuggets from South America reached 

Europe during the 1740s. Platinum was described as a new 

metal in 1750, followed by iridium and osmium in 1803, 

palladium and rhodium in 1804, and ruthenium in 1807. The 

unique properties of the metals were realized later in the 

19th century.

PGEs are essential for important industrial applications 

but are mined in only a limited number of places. Therefore, 

the availability and accessibility of PGEs could be disrupted 

by economic, environmental, political, and social events. 

This is not a new development. During World War I, the 

United States considered platinum a “vital war material” 

needed for the production of amplifiers, magnetos, munitions, 

shell primers, and sulfuric acid, and it was ranked second on 

the “list of strategic raw material” for which the United States 

was either entirely or partially dependent on foreign countries 

(Lael and Killen, 1982). During World War II, these metals 

had even greater strategic importance to science and industry, 

and the United States was still dependent on sources in 

foreign countries (DeMille, 1947; U.S. Bureau of Mines and 

U.S. Geological Survey, 1947). Security of supply was also 

a concern during the Cold War (U.S. Congress, Office of 

Technology Assessment, 1985; Sutphin and Page, 1986). With 

the dissolution of the Soviet Union and the peaceful transition 

of government in South Africa in the early 1990s, import 

vulnerability became less of an immediate concern. The 

United States net import reliance as a percentage of apparent 

consumption is still about 90 percent, however, and the main 

sources of these elements are still limited to a few mineral 

deposits in South Africa and Russia (Loferski, 2012a, b). 

The potential restrictions on supplies of critical raw minerals, 

such as the PGEs, have again become a topic for international 

concern and discussion (National Research Council of 

the National Academies, 2008; Buchert and others, 2009; 

European Commission, 2010; American Physical Society 

Panel on Public Affairs and Material Research Society, 2011; 

British Geological Survey, 2012).

This chapter summarizes how the PGEs are used, gives 

an overview of where PGEs are mined, and summarizes 

how much mineralized material has been found by mineral 

exploration companies. The geology of significant deposits is 

described, and the assessment for undiscovered resources is 

discussed, as are possible topics for future research. The report 

also discusses environmental aspects of the PGEs, including 

their sources and fate in the environment; mine waste characteristics; and human health, ecological, and carbon footprint 

concerns. Where indicated, data in tables 1 through 9 are 

taken from data prepared by the authors in table 10 (back of 

chapter).

Uses, Applications, and Consumption

The PGEs possess excellent catalytic and stable electrical 

properties. PGEs are also highly resistant to wear, tarnishing, 

and chemical attack, and can withstand high temperatures, 

making them indispensable to many industrial applications. 

Their leading use is as catalysts for oxidation and reduction 

reactions that decrease hydrocarbon, carbon monoxide, and 

nitrous oxide emissions in automobile exhaust (fig. N1). 

The chemical industry requires either platinum or platinumrhodium alloy to manufacture nitric oxide, which is a raw 

material used to manufacture explosives, fertilizers, and 

nitric acid. PGEs are also used in the production of specialty 

silicones, which, in turn, have many uses, such as in coatings 

on automobile airbags, water repellent coatings, and adhesives 

for sticky notes. In the petrochemical industry, platinumsupported catalysts are needed to refine crude oil and to 

produce high-octane gasoline and aromatic compounds.
Palladium consumption, in thousand troy ounces

0

2,000

4,000

6,000

8,000

10,000 10,000 Palladium consumption, in thousand troy ounces

0

1,000

2,000

3,000

4,000

Year

Palladium consumption, in thousand troy ounces

2000 2002 2004 2006 2008 2010 2012

0

500

1,000

1,500

2,000

A B

C D

E F

World

North America

China

0

500

1,000

1,500

2,000

Platinum consumption, in thousand troy ounces

Year

2000 2002 2004 2006 2008 2010 2012

0

500

1,000

1,500

2,000

2,500

Platinum consumption, in thousand troy ounces

0

2,000

4,000

6,000

8,000

Platinum consumption, in thousand troy ounces

Investment

Jewelry

Industrial

Autocatalyst

EXPLANATION

Introduction N3

Figure N1. Graphs showing platinum and palladium consumption, 

by category of use, from 2000 to 2012 for the world (A and B), 

North America (C and D), and China (E and F ). The decrease in platinum 

consumption for North America is directly related to the decreased 

production of cars and trucks during the same period (Organisation 

Internationale des Constructeurs d’Automobiles, 2013). The graphs 

also illustrate how consumption of platinum jewelry varies by region. 

In each graph, the layers are placed one above the other, forming a 

cumulative total. Created from data in Platinum Today (2013c).
N4 Critical Mineral Resources of the United States—Platinum-Group Elements

Alloys of PGEs are exceptionally hard and durable, 

making them the best known coating for industrial crucibles 

used in the manufacture of chemicals and synthetic 

materials, such as high-purity single crystals used in the 

production of light-emitting diodes (LEDs). PGEs are 

used by the glass manufacturing industry in the production 

of fiberglass and flat-panel and liquid crystal displays. In 

the electronics industry, PGEs are used in computer hard 

disks to increase storage capacity, and they are ubiquitous 

in electronic devices, hybridized integrated circuits, and 

multilayer ceramic capacitors. Platinum does not corrode 

inside the body, and allergic reactions to platinum are 

extremely rare; therefore, it is used in medical implants, such 

as pacemakers. PGEs are also used in cancer-fighting drugs. 

Other uses of PGEs include gas sensors in automobiles 

and homes, spark plug tips, and additives to superalloys 

(Impala Platinum Holdings Ltd., 2012; Loferski, 2013a; 

Platinum Today, 2013a).

Their white color, strength, and resistance to 

tarnish make platinum alloys an ideal choice for jewelry 

(Platinum Today, 2013b). In parts of Ecuador and Colombia 

before the Spanish conquest, platinum was used by indigenous peoples for the fabrication of small objects, as well 

as sintered with gold, and incorporated into bimetallic and 

platinum-clad articles (fig. N2; Scott and Bray, 1980; Meeks 

and others, 2002; Noguez and others, 2006). Their masterworks reflect a high degree of technical innovation because 

the high melting temperature of platinum (1,769 degrees 

Celsius [oC]) requires sophisticated manufacturing techniques and craftsmanship. The modern tradition of using 

platinum for jewelry began in Europe in the 18th century. 

The works produced by Cartier and Tiffany in the late 

19th century and early 20th century created interest in 

platinum jewelry, particularly in the United States. The Great 

Depression of the 1930s and the advent of World War II 

suppressed the platinum jewelry market in the United States. 

In the 1960s, demand for platinum jewelry in Japan surged, 

followed by a surge in demand in China in the mid-1990s.

Platinum, palladium, and rhodium are also used for 

investment in the form of physical or financial assets. 

Physical assets include platinum and palladium as collectible 

coins, or as bullion coins, bars, or wafers. Financial assets 

include stocks, mutual funds, and exchange-traded funds. 

Financial assets allow investors to own platinum, palladium, 

and rhodium without the difficulties associated with physically holding the metal.

Figure N2. Photograph of gold mask with platinum highlights, 

from the period of La Tolita culture, Ecuador. From Museo 

Nacional del Banco Central del Ecuador collection.

Geology

Geochemistry

All chemical elements heavier than oxygen, including 

the PGEs, are created by nuclear fusion and nucleosynthesis 

processes during supernova explosions. The matter derived 

from supernova explosions is dispersed into the interstellar 

medium and enriches molecular clouds, which are sites of star 

and planetary formation. Earth formed from the same cloud 

of matter as the sun, so the bulk PGE composition of Earth is 

thought to be similar to that of the solar system. Carbonaceous 

meteorites of the Ivuna type (carbonaceous chondrites 

[type 1], or C1 chondrites) are used to estimate solar system 

abundances (Lodders, 2010). The Orgueil meteorite, which is 

considered most representative of the C1 chondrites, contains 

0.947 part per million (ppm) platinum.

Early in its evolution, Earth differentiated into a metallic 

core, a silicate mantle, and a silicate crust. The segregation 

processes affected the distribution of the PGEs and resulted in 

concentrations progressively decreasing from core to mantle 

to upper crust. The range of platinum concentration in samples 

of iron meteorites, which are perhaps the best analogs for the 

composition of Earth’s core, is 2.4 to 16 ppm (Wasson and 

others, 1989). The average platinum concentrations of samples 

of the upper mantle vary from about 0.002 to 0.005 ppm 

(2 to 5 parts per billion [ppb]) (Maier and others, 2012). 

In contrast, the upper crust is estimated to contain only 

0.0005 ppm (0.5 ppb) platinum (Rudnick and Gao, 2003). 

Today, the average grade of PGEs in ores that are mined 

primarily for their PGE concentrations ranges from 5 to 15 ppm; 

however, the concentration of PGEs in hand-picked ore specimens may range from tens to hundreds of parts per million.
Geology N5

Mineralogy

The PGEs can occur as an essential component of the 

crystal structure of a mineral (a naturally occurring inorganic 

element or compound having a periodically repeating arrangement of atoms and characteristic chemical composition, 

resulting in distinctive physical properties). The Commission 

on New Minerals, Nomenclature and Classification of the 

International Mineralogical Association has recognized 

more than 100 different minerals in which at least one of 

the PGEs is an essential component (Cabri, 2002); they are 

called platinum-group minerals. Platinum-group minerals 

include native-metal-minerals and compounds with other 

transition metals (such as copper, iron, mercury, nickel, and 

silver), post-transition metals (such as bismuth, lead, and tin), 

metalloids (such as antimony, arsenic, and tellurium), and 

nonmetals (such as selenium and sulfur) (table N1). In most 

rocks, platinum-group minerals are fine-grained and range 

in size from less than a micron to a few hundred microns 

in diameter. Most geologists can spend a lifetime working 

on rocks enriched in PGEs and never see a platinum-group 

mineral in a hand specimen. Yet the platinum-group minerals 

are important to characterize because information about their 

mineralogy is needed to extract metals from their ore minerals 

effectively and to understand ore-forming processes.

PGEs also occur in solid solution in base-metal sulfides 

and sulfarsenides. In solid solution, different chemical elements 

can substitute in specific atomic sites without changing a 

mineral’s crystalline structure. For example, palladium can 

substitute for nickel in pentlandite (Cabri, 1992); the amount 

of substitution can be significant and is detectable by several 

microanalytical techniques. PGEs also occur in solid solution in 

nickeliferous cobaltite (nickel-iron sulfarsenide) in ores mined 

at the Copper Cliff North Mine in Sudbury, Ontario, Canada 

(Szentpéteri and others, 2002). Low-level concentrations of 

PGEs in solid solution in base-metal sulfide minerals can be 

assessed using laser-ablation inductively coupled plasma-mass 

spectrometry (ICP–MS) (for example, Cabri and others, 2003; 

Godel and others, 2007; Pagé and others, 2012). These studies 

document that pentlandite is the principal base-metal sulfide 

mineral that hosts PGEs. Pyrrhotite (Fe1-xS) can contain some 

iridium, osmium, rhodium, and ruthenium, but not palladium 

or platinum. Chalcopyrite (CuFeS2) does not contain significant amounts of PGEs in solid solution. Platinum and gold do 

not partition into base-metal sulfide minerals.

Deposit Types

Mineral deposits can be classified into groups or types 

based on common features and associations that ultimately 

relate to the underlying geologic processes that formed the ore. 

Each mineral deposit type has characteristic geometries, distributions of tonnage and grade, and rock and mineral properties 

that determine the potential value of the deposit. Each deposit 

type also has specific characteristics that determine how much 

sampling will be required to delimit the deposit’s mineral 

resources and in what manner the valuable material will be 

mined and processed. Furthermore, each deposit type exerts a 

specific impact on the environment, whether through natural 

weathering processes or the result of mining.

At a basic level, economic geologists recognize the 

following mineral deposit types: (a) magmatic—formed 

during the cooling and crystallization of magma, 

(b) hydrothermal—formed by the interaction of hot water 

and rock, (c) sedimentary—formed by the precipitation or 

settling of minerals directly from water in oceans or lakes, 

(d) residual—formed by the intense weathering of rocks 

under hot, humid conditions, and (e) placers—formed by the 

physical concentration of heavy particles of rocks or minerals 

by the action of moving water. PGE-enriched rocks can be 

found in each of the deposit types listed above. The following 

sections describe examples of PGE mineralization found in 

each of these deposit types. More information is given for 

magmatic deposits because they are the source of almost all 

the PGEs recovered by mining. The locations of the world’s 

major and minor PGE deposits are shown in figure N3.

Magmatic Deposits

PGEs are transferred from Earth’s mantle to its crust by 

magnesium-rich magmas that were formed by the melting of 

a high percentage of mantle material from which magma had 

not been previously extracted. A high degree of partial melting 

is needed to release PGEs into the melt from the small quantities of sulfide minerals or alloys that are present in the mantle. 

Upon emplacement in the crust, the magnesium-rich magmas 

cool to form mafic and ultramafic igneous rocks.

Magmatic deposits are concentrations of metallic 

oxide or sulfide minerals that formed during the cooling and 

crystallization of magma. Textural and experimental evidence 

indicate that PGE-enriched mineralization forms when mafic 

to ultramafic magma becomes saturated with sulfur and an 

immiscible sulfide liquid exsolves from the silicate magma 

(Barnes and others, 2008; Holwell and McDonald, 2010; 

Naldrett, 2010a). The solubility of sulfur in mafic magmas is 

affected by changes in the bulk composition of the magma, 

the fugacity of sulfur and oxygen, temperature, and pressure 

(Ripley, 1999). Processes that change the solubility of sulfur 

and may cause an exsolution event include (a) fractional 

crystallization of the silicate magma, (b) mixing of magmas, 

(c) assimilation of additional sulfur from sources external to 

the magma, and (d) modification of the magma composition 

by bulk contamination, such as changing the silica content of 

the magma (Mungall, 2005).

The silicate magma solidifies first, followed by the sulfide 

liquid, which solidifies at temperatures in excess of 900 °C. 

The textures and mineralogy of PGE ores record a prolonged 

and complex process of solid-state transformation and 
N6 Critical Mineral Resources of the United States—Platinum-Group Elements

Table N1. Chemical formulas for selected platinum-group minerals as well as other common rock-forming minerals mentioned 

in this chapter.

[Sources: Deer and others (1966); Vaughan and Craig (1978); and Cabri (2002). Elements: Al, aluminum; As, arsenic; Bi, bismuth; C, carbon; 

Ca, calcium; Cr, chromium; Cu, copper; Fe, iron; Hg, mercury; Ir, iridium; K, potassium; Mg, magnesium; Na, sodium; Ni, nickel; O, oxygen; 

Os, osmium; Rh, rhodium; Ru, ruthenium; Pb, lead; Pd, palladium; Pt, platinum; S, sulfur; Sb, antimony; Si, silicon; Sn, tin; Te, tellurium]

Mineral name Chemical formula

PLATINUM-GROUP MINERALS

Native metals

Iridium Ir

Osmium Os

Palladium Pd

Platinum Pt

Rhodium Rh

Ruthenium Ru

Compounds with transition metals

Isoferroplatinum Pt3

Fe

Potarite PdHg

Tulameenite PtFe0.5Cu0.5

Compounds with post-transition metals (Bi, Pb, and Sn)

Atokite Pd3

Sn

Froodite PdBi2

Insizwaite PtBi2

Rustenburgite Pt3Sn

Paolovite Pd2

Sn

Compounds with metalloids 

Genkinite (Pt,Pd)4

Sb3

Geversite PtSb2

Isomertieite Pd11Sb2As2

Kotulskite PdTe

Merenskyite PdTe2

Moncheite PtTe2

Sperrylite PtAs2

Stibiopalladinite Pd5+xSb2-x

Stillwaterite Pd8As3

Sudburyite PdSb

Compounds with nonmetals

Braggite (Pt,Pd)S

Cooperite PtS

Erlichmanite OsS2

Laurite RuS2

Vysotskite PdS

Other

Hollingworthite RhAsS

Maslovite PtBiTe

Michenerite PdBiTe

Ruarsite RuAsS

Mineral name Chemical formula

BASE-METAL SULFIDE MINERALS

Bornite Cu5

FeS4

Chalcopyrite CuFeS2

Covellite CuS

Cubanite CuFe2

S3

Marcasite FeS2

Mooihoekite Cu9Fe9S16

Nickeliferous cobaltite Co(Ni)AsS

Pentlandite (Fe,Ni)9

S8

Pyrite FeS2

Pyrrhotite Fe(1-x)S

Talnakhite Cu9

(Fe, Ni)8S16

Troilite FeS

OTHER ROCK-FORMING MINERALS

Calcite CaCO3

Chlorite (Mg,Fe)3(Si,Al)4O10(OH)2 • 

(Mg,Fe)3

(OH)6

Chromite (Fe,Mg)(Cr,Al)2

O4

Goethite FeO(OH)

Gypsum CaSO4

Jarosite KFe(SO4)2(OH)6

Magnetite Fe3O4

Native sulfur S

Olivine (Mg,Fe)2

SiO4

Plagioclase (Ca,Na)(Si,Al)4O8

Pyroxene Ca(Mg,Fe)Si2O6 to

Mg2

Si2O6

Serpentine (Mg,Fe)3Si2O5 (OH)4

Talc Mg3Si4O10(OH)2
Geology N7

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Goodnews Bay

Wellgreen

Intrusion

Skaergaard

Intrusion

Stillwater

Complex Duluth

Lac des Iles

Complex Sudbury Complex

Bushveld

Complex

Stella

Intrusion

Great

Dyke

Uitkomst Intrusion

Portimo

Igneous

Complex

Fedorov-Pana

Intrusion

Burakovskaya

Intrusion

Onega

Intrusion

Kytlym

Noril’sk Talnakh

Kondyor

Koryak Kingash and 

Upper Kingash

PGE figure N3.

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

EXPLANATION

Intrusion with greater than  200,000 kilograms PGEIntrusion with less than  200,000 kilograms PGESelected placer deposits

s platinum-group-element (PGE) deposits, as well 

o be major when they contain 200,000 kilograms (kg) 

World map showing locations of igneous intrusions and intrusive complexes that contain most of the world’ Figure N3.

as the placer deposits that are mentioned in the text. On the map, intrusions and intrusive complexes are considered tor more of PGEs, and minor when they contain less than 200,000 kg of PGEs.
N8 Critical Mineral Resources of the United States—Platinum-Group Elements

recrystallization after solidification (Barnes and others, 2008; 

Holwell and McDonald, 2010). At high temperature (about 

1,000 °C), osmium, iridium, and ruthenium preferentially 

partition from the immiscible sulfide liquid into a mineral 

called monosulfide solid solution (MSS). Platinum, palladium, 

and gold, on the other hand, behave as incompatible elements 

with respect to MSS and instead are concentrated in residual 

copper-rich sulfide liquids and associated with the minerals 

that crystallize from the residual immiscible sulfide liquid. 

Upon cooling, some of the PGEs held in sulfide crystal 

structures are expelled, forming discrete minerals. The final 

mineral assemblage consists of varied proportions of pyrrhotite, pentlandite, chalcopyrite, and bornite, with platinum and 

gold occurring primarily in platinum-group minerals.

Magmatic sulfide minerals in mafic to ultramafic igneous 

rocks do not always contain elevated concentrations of PGEs. 

The metal content of immiscible sulfide liquids is a function 

of the (a) amount of metal in the silicate magma; (b) relative 

affinity for metals to occur in the sulfide or silicate liquid; 

and (c) relative amounts of the two liquids (Campbell and 

Naldrett, 1979; Barnes and Maier, 1999). Such elements 

as copper, nickel, and PGEs originally in the silicate melt 

preferentially concentrate into the sulfide liquid when it 

exsolves. Concentration of these metals into the sulfide liquid 

can, under some circumstances, deplete their concentration in 

the silicate magma. If there is a large volume of sulfide liquid 

compared with that of the silicate magma, the sulfide liquids 

and the resulting ores will have lower concentrations of PGEs. 

If the sulfide liquid effectively interacts with a large volume of 

silicate magma, the sulfide liquids become enriched in PGEs, 

resulting in high-grade ores.

Economic geologists who study magmatic processes 

expend a lot of effort trying to understand how large mass 

ratios are achieved. Proposed answers have included mixing 

of magmas; the migration of interstitial melts (and “fluids”) 

upward through crystal mush; and (or) the streaming of 

magma over sulfide liquids in a channelized lava flow, sill, 

or feeder dike (Barnes and Maier, 1999; Mathez, 1999; 

Naldrett, 2010a).

Tectonic Setting of Magmatic Deposits 

(Large Igneous Provinces)

As the theory of plate tectonics gained acceptance in the 

1960s and its predictive power became evident, geologists, 

including those working on magmatic ore deposits, tried to 

understand magmatism and ore genesis in a plate-tectonic 

context (Hutchison, 1983; Sawkins, 1984). For some mafic 

and ultramafic rocks that are related to subduction zones and 

mid-oceanic ridges, this conceptual framework provided 

insight and clarity; however, for other types of mafic and 

ultramafic rocks found in intraplate settings, which include 

rocks that host most of the significant magmatic mineral 

deposit types, it did not.

Most magmatic copper-nickel-PGE deposits occur 

with continental flood basalts, aerially extensive mafic dike 

swarms, sill provinces, and large layered ultramafic-to-mafic 

intrusions, which are indicative of large igneous provinces 

(LIPs) (Coffin and Eldholm, 1994). Bryan and Ernst proposed 

that LIPs are magmatic provinces with areal extents greater 

than 1×105 square kilometers (km2

), igneous volumes greater 

than 1×105 cubic kilometers (km3) and maximum lifespans 

of about 50 million years (m.y.) that “have intraplate tectonic 

settings or geochemical affinities, and are characterized 

by igneous pulse(s) of short duration” (~1 to 5 m.y.), 

during which time “a large proportion (>75 percent) of the 

total igneous volume has been emplaced. [The LIPs] are 

dominantly mafic, but also can have significant ultramafic 

and silicic components, and some are dominated by silicic 

magmatism” (Bryan and Ernst, 2008, p. 175). Unlike the 

vast majority of igneous rocks that are associated with plate 

tectonic processes at convergent or divergent tectonic plate 

margins, LIP-related rocks usually occur in an intraplate 

tectonic setting; the association of some LIPs with hotspot 

tracks—for example, the North Atlantic Igneous Province and 

the Iceland hotspot (Storey and others, 2007) is one reason 

that LIPs are attributed to mantle plumes.

Types of Magmatic Deposits

Magmatic mined deposits associated with LIP-related 

igneous intrusions are divided into types based on such 

criteria as the lithology and form of associated igneous rocks, 

the depth of emplacement, the abundance of sulfide minerals, 

the relative proportion of metals, and the position of the ores 

within the intrusion. In this chapter, we distinguish between 

conduit-type deposits, which occur in intrusions that are part 

of sill complexes and dike swarms, and reef-type and contacttype deposits, which occur in layered intrusions (fig. N4). The 

most significant and representative examples for the world and 

the United States are the conduit-type deposits of the Noril’skTalnakh area (Russia); the reef-type deposits in South Africa 

(the Merensky Reef and the UG2 Chromitite), Zimbabwe 

(the Main Sulphide Zone), and Montana (the J–M Reef); and 

the contact-type deposits in South Africa (the Platreef) and 

Minnesota (the Duluth Complex).

Conduit-Type Deposits

The conduit-type deposits of the Noril’sk-Talnakh area of 

Russia are associated with an enormous outpouring of mafic 

magma that formed the Siberian Traps—the largest continental flood basalt province on Earth (fig. N5). The magmas 

erupted in a short period of time; the 3.5-kilometer (km)-thick 

basalt succession at Noril’sk-Talnakh was emplaced between 

248.7±0.6 and 250.3±1.1 mega-annum (Ma) (Reichow and 

others, 2009). The flood basalts crop out on the Siberian 

craton, covering an area of approximately 2.5 million km2

(Fedorenko and others, 1996). In addition, flood basalts 
Geology N9

Contact-type

NOT TO SCALE

mineralization

Reef-type

mineralization

Conduit-type

mineralization

Conduit-type

mineralization

Lava flow

Dike Sill

Contact-type

mineralization

Layered

intrusion

PGE fig N4 pluton_dike_sill_v2

Layered mafic to ultramafic intrusion

Platinum-group element mineralization

EXPLANATION

Magma or lava

Sill or dike

Sedimentary rock layers Figure N4. Schematic block diagram showing changes in the form of igneous intrusions with depth and 

the relative occurrence of conduit-type, contact-type, and reef-type magmatic ore deposits. Layered 

igneous rocks with reef-type and contact-type deposits occur at depth in large layered intrusions. Sills and 

dikes provide the conduits or pathways that magma follows as it rises upwards through the crust. Conduittype magmatic sulfide deposits can be localized in sills or dikes. Modified from Naus-Thijssen (2007).

are also found in the late Mesozoic fold and thrust belt in 

the Taimyr Peninsula and under the West Siberian Basin 

(Reichow and others, 2009).

A sill complex lies beneath the thick succession of 

volcanic flows and is exposed where the flood basalts have 

been removed by erosion (fig. N5). The sill complex consists 

of thousands of intrusions that delineate the pathways along 

which large volumes of magma were transported through 

the crust (Arndt, 2005). Sills with major nickel-copper-PGE 

deposits have been identified only in the Noril’sk-Talnakh 

area; however, copper-nickel sulfide occurrences in sills 

are distributed over a much larger area (fig. N5). All these 

occurrences of volcanic rocks and associated sills make up 

the Siberian Traps LIP (Ernst and Buchan, 2001).

In the Noril’sk-Talnakh area, plateaus covered by flood 

basalts have been dissected by erosion and exposed the 

underlying sill complex (fig. N6). More than 300 intrusions have 

been mapped, but only 33 contain elevated sulfide concentrations.

Of these 33 intrusions, 16 contain relatively rich disseminated 

ore and only 4 contain rich massive sulfide orebodies (Diakov 

and others, 2002). High sulfur contents coupled with increasing 

δ34S values is consistent with the addition of crustally derived 34S-enriched sulfur to the magma of the ore-bearing intrusions 

(Grinenko, 1985; Li and others, 2003).

The ore-bearing intrusions have an elongate, finger-like 

shape; they may be up to 1 km wide by 500 meters (m) thick 

and up to 15 km long. Most of the other sills form sheet-like 

bodies that are tens of meters thick. Rock textures of sills 

indicate crystallization and cooling at shallow depth. The orebearing intrusions are internally differentiated, with magnesium 

rocks and minerals concentrated towards the base of the sills; 

the laterally continuous sills show less internal differentiation.
N10 Critical Mineral Resources of the United States—Platinum-Group Elements

KAZAKHSTAN

RUSSIA

MONGOLIA

Taimyr Peninsula

Ural Mountains

60° E 90° E 120° E

70° N

60° N

50° N

Undeformed cover of sedimentary rocks

West Siberian Basin

Siberian flood basalts, exposed

Siberian flood basalts, covered

Orogenic belts

Archean and Proterozoic basement rocks

EXPLANATION

PGE fig N5 Norilsk_index_map

Political boundaries from U.S. Department of State (2009) 0 200 400 600 800 KILOMETERS

Asia North Lambert Conformal Conic Projection

Central meridian, 90° E., latitude of origin, 30° N. 0 200 400 MILES

Extent of sill complex associated with

 exposed Siberian flood basalts

Major rift boundary

Nickel–copper–platinum-group- 

 element occurrence 

Kara Sea

Lake Baikal

Barents Sea Laptev Sea

Extent of

figure N6

Figure N5. Map showing the geology of the Siberian flood basalt province in Russia, which is the largest flood basalt 

province in the world. The province includes a large area of exposed lavas and sills. It also includes an equally large 

area of lavas that are covered by younger deposits that make up the West Siberian Basin. The sill complex associated 

with the exposed lavas underlies and extends beyond the area of lavas. The deposits of the Noril’sk-Talnakh area, along 

with many occurrences of magmatic sulfide mineralization, have been found in the northwestern part of the province 

where the lavas and sills are exposed. Map includes information from Zonenshain and others (1988), Fugro Robertson, 

Ltd. (2008), and Reichow and others (2009).
Geology N11

! Noril`sk

Oktyabr’sk

Mine

Komsolmosk

Mine

Taimyrsky

Mine

Mayak

Mine

Skalisty

Mine

Medvezhy

Creek

open pit

Maslovskoe

deposit

Zopolyarny

Mine Noril’sk II

deposit

Chernogorskoye

deposit

Mineralized sill (ore-bearing intrusion),

 projected to the surface 

Mafic sills

Permo-Triassic Siberian province 

 flood basalt

Sedimentary rocks

Quaternary sediments

Nickel–copper–platinum-group-

 element deposit

EXPLANATION

0 2 4 6 8 KILOMETERS

0 2 4 MILES

Asia North Albers Equal Area Conic Projection

Central meridian, 88° E., latitude of origin, 30° N.

Fault

Area of fig N7A

88°00' E 88°30' E

69°30' N

69°10' N

PGE fig N6 Norilsk

80° E 100° E

75° N

65° N

RUSSIA

Kara Sea

Map area

Figure N6. Map showing the geology of the Noril’sk-Talnakh area and the location of nickel–copper–platinum-groupelement (Ni-Cu-PGE) deposits. Flood basalts crop out in the northeast and the southeast; in other areas, erosion has 

removed the lavas and exposed the underlying sills that intrude older sedimentary rocks. The mineralized sills largely 

underlie the lavas. Their subsurface extent has been projected to the surface. Geology simplified from Sherman and 

others (1991). 
N12 Critical Mineral Resources of the United States—Platinum-Group Elements

The amount of massive sulfide ores associated with the 

mineralized intrusions at Talnakh is astounding. The massive 

sulfide ores, which are up to 45 m thick, underlie most of the 

intrusions in the Talnakh area (fig. N7A; Kunilov, 1994). The 

ore-bearing sills contain, or are associated with, concentrations 

of sulfide minerals that exceed the amount of sulfur that could 

have been dissolved in the volume of magma now in the sills. 

The sulfide mineralization also contains high concentrations 

of PGEs, indicating a silicate-to-sulfide ratio that is larger than 

the observed proportion of silicate minerals to sulfide minerals 

in the sills. This suggests that at the time of ore formation, 

the sulfide liquids equilibrated with a much larger volume of 

magma than is now represented by the igneous rocks in the 

sill. Field relations indicate lateral movement of immiscible 

sulfide liquids during and after solidification of the silicate 

igneous rocks.

In the Talnakh area, fractional crystallization of the 

immiscible sulfide liquids formed mineralogically and 

compositionally zoned orebodies (fig. N7B). Massive ores 

formed by early crystallization of sulfides contain less copper 

and PGEs than ores that formed from sulfide liquid that are the 

end product of the fractional crystallization process (fig. N8). 

The copper- and PGE-rich late-stage ores are the likely source 

of most of the PGE production from the Talnakh deposits 

(fig. N7C). Between 1960 and 2011, Russia (including all 

the Soviet Union until 1991) produced about 5,000 metric 

tons of PGEs, most of them from the Noril’sk-Talnakh area 

(U.S. Bureau of Mines, 1933–96; U.S. Geological Survey, 

1997–2016). Mining depths in the Noril’sk-Talnakh area 

range from 300 to 1,500 m (Kunilov, 1994).

Reef-Type and Contact-Type Deposits

LIP-related layered mafic to ultramafic intrusions contain 

most of the world’s resources of PGEs. Layered intrusions 

form by crystallization of magma deep in Earth’s crust and 

have sheet-like to dike-like shapes. They consist primarily 

of cumulates, which are rocks that represent accumulations 

(concentrations) of minerals that crystallize from magmas. 

Cumulus rock textures are distinctive, as they record 

nucleation and growth of crystals from the melt, enlargement 

of crystals to form a touching framework (a crystal mesh), 

solidification, and subsolidus grain boundary adjustments that 

minimize grain boundary surface energies (Hunter, 1987). 

Cumulus rocks are also distinctive in that they do not have the 

composition of naturally occurring magmas but are instead a 

mixture of early formed crystals with variable proportions of 

trapped liquid.

As the magma crystallizes in these intrusions, layering 

features develop that are recognized by the variations in 

the modal proportions of the minerals, the rock textures, 

the grain size, and the mineral compositions. Individual 

layers range from laminations formed by the orientation of 

individual crystals that are millimeters in diameter to strata 

that are tens to hundreds of meters thick and may extend for 

hundreds of kilometers along strike. The layering features 

are so pronounced and consistent that stratigraphic principles 

developed for sedimentary rocks are applied to these igneous 

rocks. Stratigraphic columns are measured through the 

sequence of layered igneous rocks to define mappable units; 

the thickness of the layered rocks in intrusions ranges from 

hundreds of meters to as much as 10 km.

Layered intrusions have two significant styles of 

magmatic sulfide mineralization—reef type and contact 

type. The analogy with sedimentary rocks extends to the 

terminology used for magmatic deposits in layered intrusions. 

Some of the magmatic deposits are strata-bound, which means 

that the deposit is confined to a stratigraphic unit, not to a 

particular bed. Other magmatic deposits are stratiform, which 

means that the deposit constitutes one or more of the igneous 

layers; an example would be the chromium deposits in layered 

intrusions (which correspond to igneous rock layers that are 

composed primarily of the mineral chromite). Reef-type and 

contact-type deposits are strata-bound, as described below.

Reef-type PGE deposits are disseminated copper-, 

iron-, nickel-, and PGE-bearing minerals that are associated 

with one or more strata within a layered igneous intrusion 

(fig. N4). The mineralized rock unit almost always contains 

disseminated magmatic sulfide minerals, but the rock layers 

are dominated by silicate minerals or oxide minerals such 

as chromite or magnetite. The modal abundance of sulfide 

minerals is usually much less than a few percent. The term 

“reef” is an Australian and South African mining term for a 

relatively flat-lying, tabular orebody. Within a layered igneous 

intrusion, reef-type mineralization is laterally persistent, 

extending for the strike length of the intrusion, which is 

typically tens to hundreds of kilometers. The mineralized 

interval is thin (generally centimeters to meters thick), 

however, relative to the total stratigraphic thickness of the 

layers in the intrusion, which can vary from hundreds to 

thousands of meters.

Copper-nickel-PGE contact-type deposits consist of 

disseminated magmatic sulfide minerals found near the lower 

contact or margin of mafic to ultramafic layered intrusions 

(fig. N4). The host rocks for the disseminated sulfide minerals 

are the igneous rocks of the intrusion and adjacent contact 

metamorphosed country rocks. Some igneous lithologies are 

cumulates, but other mineralized mafic and ultramafic igneous 

rocks are texturally and lithologically heterogeneous, exhibiting changes in texture and mineral proportions at scales of 

from centimeters to meters. Textures indicative of chilling or 

unidirectional growth of silicate minerals may also be present. 

Inclusions of autoliths and xenoliths are common; they are 

surrounded by igneous rocks that have textural, mineralogical, 

and isotopic features suggestive of reaction with the inclusions 

(Iljina and Lee, 2005).
Geology N13

Chalcopyrite and pyrrhotite

Chalcopyrite, cubanite, 

 and troilite

Mooihoekite and cubanite

Pyrrhotite

Oktyabr’sk orebody mineral

 assemblage zoning (all ore

 types carry pentlandite)

0 500 1,000 1,500 FEET

0 250 500 METERS

Severnij

Mine

Skalisty

Mine

Glubokij

Mine

Taimyrsky

Mine

Oktyabr’sk

Mine

Mayak

Mine

0 1 2 MILES

0 1 2 3 4 KILOMETERS

Talnakh intrusion (includes

 intrusion of central graben)

Kharayelakh intrusion

Massive sulfide orebody

Mine boundary

EXPLANATION

Map A

Map C

Map B

Fault

Area of B

Area of C

0 500 1,000 1,500 FEET

0 250 500 METERS

Schedule for mining

Mined out as of 2000

2001 to 2010

2021 to 2030

2031 to 2045

No data

Intrusion not present

2011 to 2020

A B

C

PGE fig N7 Talnakh deposit with mined out areas

N

N

N

Figure N7. Maps showing nickel–copper–platinum-group-element (Ni-Cu-PGE) deposits in the Talnakh area, Russia. A, Extent of 

two elongate sill-like intrusions (Talnakh and Kharayelakh) and the distribution of massive sulfide ores that underlie them. B, Zoning 

of sulfide minerals in the Oktyabr’sk orebody. The concentric pattern of minerals formed by fractional crystallization of immiscible 

sulfide melt; the pyrrhotite (Fe1–xS) assemblage solidified first, and mooihoekite (Cu9

Fe9S16 ) and cubanite (CuFe2S3) solidified last. 

The copper-rich ore assemblages contain the most PGEs. C, Mined-out areas of the Oktyabr’sk orebody and future mine plans to 

2045. Most of the copper- and PGE-rich ores have been mined. Maps modified from Zientek and others (1994) and Naldrett (2006). 
N14 Critical Mineral Resources of the United States—Platinum-Group Elements

Figure N8. Photograph of copper-rich massive 

sulfide ore exposed in a stope in the Oktyabr’sk 

Mine in the Talnakh area, Russia. The ore consists 

of several iron-copper-sulfide minerals, such as 

chalcopyrite (CuFeS2

), mooihoekite (Cu9Fe9S16 ), 

and talnakhite (Cu9

(Fe, Ni)8S16 ). The minerals 

are brassy to golden yellow, but some develop 

iridescent blue and maroon colors when exposed 

to air. These ores represent the end product of 

the fractional crystallization of immiscible sulfide 

liquid and can contain tens to hundreds of parts per 

million platinum-group elements. Photograph by 

Michael L. Zientek, 1992.

Sulfide abundance in contact-type PGE deposits is 

typically about 3 to 5 volume percent, although some nettextured and (or) massive sulfide ores may be present. Erratic 

variation in the distribution of sulfide minerals is typical, 

although the concentration of sulfide minerals within the 

intrusion generally increases towards the contact with adjacent 

country rocks. The mineralization can be laterally persistent, 

commonly extending for the entire strike length of the layered 

igneous intrusion. The interval that can contain magmatic 

sulfide minerals is generally tens to hundreds of meters in 

thickness, however. The proportion of sulfide minerals varies 

within the rock layers that can host ore; during exploration 

and development, economic cutoff grades are used to define 

the volume of rock within the igneous rock unit that could 

be mined. Therefore, low-grade mineralization may occur 

outside the cutoff limit that defines minable parts of deposits. 

Examples of major reef-type (a–c) and contact-type (a and d) 

deposits are discussed below.

(a) Merensky Reef, UG2 Chromitite, and Platreef, 

Bushveld Complex, South Africa. The Paleoproterozoic 

Bushveld Complex (2,054.4±1.3 Ma, using the uraniumlead-zircon technique for determining age) is a large mass of 

igneous rock that underlies an area of approximately 69,000 km2

in South Africa (fig. N9; Hall, 1932; von Gruenewaldt, 1977; 

Scoates and Friedman, 2008). It is part of the Bushveld LIP 

(Ernst and Buchan, 2001), which also includes the Molopo 

Farms Complex (another large layered mafic to ultramafic 

intrusion in Botswana and South Africa (Prendergast, 2012); 

a large complex of sills related to the Bushveld Complex 

(Sharpe, 1981; Sharpe and Hulbert, 1985); and the Dullstroom 

basalts in South Africa (Buchanan and others, 1999).

The Bushveld Complex consists of several igneous suites. 

The most important economically is the Rustenburg Layered 

Suite, which is an approximately 8-km-thick layered sequence of 

mafic to ultramafic cumulates (Vermaak and von Gruenewaldt, 

1986; Walraven, 1986) that contain world-class deposits of 

chromium, PGEs, and titanium-vanadium. The cumulates of the 

Rustenburg Layered Suite are exposed intermittently around the 

periphery of the Bushveld Complex in areas referred to as limbs. 

Igneous layering dips gently towards the center of the Bushveld 

Complex. Seismic surveys trace igneous units exposed at the 

surface to depths exceeding 6 km (Sargeant, 2001; Campbell, 

2011). Gravity modeling indicates that the western and eastern 

limbs of the Bushveld Complex are connected at depth (fig. N9,

cross section A–A’; Webb and others, 2004).

Two PGE-enriched reefs, the UG2 Chromitite (fig. N10) 

and the Merensky Reef (fig. N11; Viljoen, 1999), each occur 

near the base of different repetitive rock sequences (cyclic 

units) and can be continuously traced onstrike for the full 

extent of the eastern and western limbs of the complex. The 

grade of the UG2 Chromitite deposit is about 5 grams per 

metric ton (g/t) PGEs and gold; the grade of the Merensky 

Reef is about 6 g/t in the western part of the complex and 

about 4.2 g/t in the eastern part.

In the northern limb, varitextured pyroxenite, norite, and 

gabbro are found near the lower contact of the complex with 

metasedimentary rocks of the Transvaal Supergroup (van der 

Merwe, 1976). These igneous rocks host the Bushveld 

Complex’s contact-type copper-nickel-PGE deposits (known 

as the Platreef). The combined PGE and gold grades for 

the Platreef deposits range from 0.55 to 3.7 g/t. Since the 

1920s, mining has recovered 7,200 metric tons of PGEs from 

the Bushveld Complex (U.S. Bureau of Mines, 1933–34; 

1933–96; U.S. Geological Survey, 1997–2016).

(b) Main Sulphide Zone, Great Dyke, Zimbabwe. The 

Neoarchean Great Dyke (2575.4±0.7 Ma, age determined from 

zircon using uranium-lead technique; Oberthür and others, 

2002) is a long (about 550 km) and narrow (about 11 km) 

layered igneous intrusion in Zimbabwe. This intrusion and 

some subparallel dikes are the geologic features associated 

with the Great Dyke of Zimbabwe LIP (Ernst and Buchan, 

2001). The rock types of the Great Dyke consist of layered 

mafic to ultramafic cumulates that dip inwards from the sides 

towards the center of the intrusion (fig. N12). Since the 1980s, 

mining has recovered 107 metric tons of PGEs from a reeftype PGE deposit within the Great Dyke—the Main Sulphide 

Zone—which occurs 10 to 50 m below the contact between 

the ultramafic and mafic sequences. The Main Sulphide Zone 

is typically 2 to 3 m thick, and the grade of the deposit varies 

from about 3.5 to 4 g/t PGEs and gold (U.S. Bureau of Mines, 

1933–96; Wilson, 1996; U.S. Geological Survey, 1997–2016; 

Wilson and Prendergast, 2001).
Geology N15

1,000

METERS

AMSL

–1,000

–2,000

0 2 4 KILOMETERS

0 1 2 MILES

Merensky Reef

UG2 Chromitite

Lower zone

Main zone

Critical zone

Rustenburg Layered Suite Orebody

10

20

0

KILOMETERS

VERTICAL EXAGGERATION ×3

VERTICAL SCALE GREATLY EXAGGERATED

Marble Hall dome

Fold and line of intrusives

Land surface

Land surface

Crocodile River dome

0 30 60 90 KILOMETERS

0 15 30 MILES

Ventersdorp Supergroup

Transvaal Sequence

Lebowa Granite Suite and 

 Rooiberg Group 

Rustenburg Layered Suite

Alkaline intrusions

Younger sedimentary and 

 intrusive rocks

EXPLANATION FOR SECTION A–A'

EXPLANATION FOR SECTION B–B'

Bushveld Complex

!

!

Pretoria

Pietersburg

SOUTH AFRICA

BOTSWANA

A

A

B

A'

A'

B'

B B'

Marble Hall dome

24° S

25° S

28° E 30° E

0 30 60 90 KILOMETERS

0 15 30 MILES

PGE fig N9 Bv_reefs_leases_geology_v5.ai

Merensky Reef and 

 UG2 Chromitite

Merensky Reef

UG2 Chromitite

EXPLANATION FOR MAP

Project lease

Orebody outcrop

Platreef

Rustenburg Layered Suite

of the Bushveld Complex

A A' Line of section

20°E 30°E

20°S

30°S SOUTH

AFRICA

BOTSWANA

ZIMBABWE

Indian Ocean

Map area

Figure N9. Map showing the Rustenburg Layered Suite of the Bushveld Complex, South Africa, the surface trace of significant 

orebodies, and cross sections through the central area and northeastern limb. Cross section A–A’ is based on geophysical 

modeling and suggests that the mafic and ultramafic rocks that make up the eastern and western portion of the complex are 

connected at depth. Modified from Webb and others (2004). Cross section B–B’ shows the remarkable continuity of the layers that 

host the Merensky Reef and UG2 Chromitite platinum-group-element deposits. Modified from Bauba Platinum (2012). Rustenburg 

Layered Suite is derived from Martini and others (2001); project leases and orebodies are from Zientek and others (2014).
N16 Critical Mineral Resources of the United States—Platinum-Group Elements

Figure N10. Photograph of the UG2 Chromitite 

at the Karee Mine in the western part of the 

Bushveld Complex, South Africa. The black layer 

at the base of the exposure is the main layer of 

the UG2 Chromitite. It has an undulating contact 

with the underlying, light-colored anorthosite 

layer. Several thin chromitite seams overlie 

the main chromitite layer. Most of the exposed 

rock is brown pyroxenite associated with the 

UG2 cyclic unit; a sharp contact separates the 

pyroxenite layer from the overlying light-colored 

anorthosite layer. Photograph by Wolfgang 

Maier, University of Cardiff, Wales, United 

Kingdom.

Merensky cyclic unit

Pyroxenite Pegmatoidal pyroxenite Layered norites

PGE Figure N11

Change in direction of rock face

Iron oxide

weathering

Iron oxide

weathering

A B

Figure N11. The base of the Merensky cyclic unit, a pegmatoidal pyroxenite, which contains the platinum-group-element-rich 

Merensky Reef. A, Photograph of the base of the Merensky cyclic unit on the Hackney farm property in the eastern part of the 

Bushveld Complex, South Africa. The basal contact of the cyclic unit is annotated by the yellow line. B, Same photograph with 

an overlay showing major features, including the dark, pegmatoidal pyroxenite cutting layering in the underlying lighter-colored 

norites. Photograph by Wolfgang Maier, University of Cardiff, Wales, United Kingdom. 
30° E 31° E

17° S

19° S

21° S

North chamber South chamber

subchamber

Darwendale

subchamber

Sebakwe

subchamber

Wedza

subchamber

Selukwe

subchamber

Musengezi

Hartley Complex

Selukwe

Complex

Wedza Complex

Musengezi Complex

C

B

ZAMBIA

DRC

SOUTH

AFRICA

Map area

24° E 36° E

10° S

20° S

0 10 20 KILOMETERS

0 10 MILES

30 40

20

PGE fig N12 GreatDyke with cross sections.ai

C C'

B B'

A A'

C'

B'

A A'

0 1 2 KILOMETERS

0 1 MILES

3 4

2

0

VERTICAL EXAGGERATION ×2

1 2 KILOMETERS

0 1 MILES

3 4

2

0.5

0

1.0

1.5

KILOMETERS KILOMETERS KILOMETERS

+10

–10

100

200

300

Gravity unit

Gravity unit

1

0

2

3

4

5

6

7

8

1

0

2

3

4

5

6

7

8

+10

–10

100

200

300

400

9

10

0 1 2 KILOMETERS

0 1 MILES

3 4

2

EXPLANATION FOR CROSS SECTIONS

Gabbronorite

Bronzitite

Serpentinite

Dunite/harzburgite

Granite

Quartz gabbro/gabbronorite

Websterite and Main Sulphide Zone

Fault

Bouguer gravity anomaly

Gravity residual from model

ZIMBABWE

Land surface

Land surface

Land surface

EXPLANATION FOR MAP

Mafic sequence

Ultramafic sequence,

 P1 layer, and Main

 Sulphide Zone

Ultramafic sequence, 

 with cyclic units

A A' Line of section

Geology N17

Figure N12. Geologic map and cross sections of the Great Dyke, Zimbabwe. Geologic mapping and geophysical studies show 

that the intrusion has a dike-like form; however, igneous layering dips gently into the center of the intrusion. The Main Sulphide 

Zone, which is a reef-type platinum-group-element deposit, is confined to the P1 layer of the ultramafic sequence (as shown on 

the map) and to the corresponding websterite layer shown on cross section C–C’. Geologic map generalized from Worst (1957); 

geologic sections generalized from Podmore and Wilson (1987).
N18 Critical Mineral Resources of the United States—Platinum-Group Elements

(c) J–M Reef, Stillwater Complex, Montana. The 

Neoarchean Stillwater Complex (2,704±5 Ma, using the 

uranium-lead zircon technique for determining age; Premo and 

others, 1990) is a mafic to ultramafic layered intrusion exposed 

in south-central Montana (fig. N13; Zientek and others, 2002). 

This intrusion is the only feature associated with the Stillwater 

LIP (Ernst and Buchan, 2001). More than 5,500 m of layered 

rocks are exposed, which can be traced for 48 km along strike. 

The J–M Reef, which is a reef-type PGE deposit, consists of 

0.5 to 3 volume percent magmatic sulfide minerals that are 

associated with a distinctive olivine-bearing cyclic unit that 

can be mapped both on the surface and underground. The 

actual distribution of sulfide minerals within the cyclic unit 

is determined by detailed mapping, drilling, and sampling, 

however. The J–M Reef, which is open at depth, has been 

traced for 42 km of the 48-km strike length of the complex 

and at least 2 km down the dip of layering. The J–M Reef 

is currently the sole source of primary PGE production and 

reserves in the United States. Since 1986, the Stillwater 

and East Boulder Mines (fig. N14) have produced roughly 

305 metric tons of PGEs from the J–M Reef deposit. The 

overall deposit grade is about 15 g/t palladium and platinum 

(U.S. Bureau of Mines, 1933–96; U.S. Geological Survey, 

1997–2016; Abbott and others, 2011).

(d) Contact-Type Deposits, Duluth Complex, Minnesota.

The Keweenawan LIP resulted from the eruption of lavas and 

emplacement of igneous intrusions between 1,109 and 1,087 Ma 

along a linear belt where Earth’s crust subsided (the Midcontinent Rift). The rift extends for more than 2,500 km from 

Kansas northward beneath Lake Superior and then southeast 

through Michigan (fig. N15); however, exposures of rocks 

related to the rift are found only in the Lake Superior region. 

Igneous rocks associated with this LIP cover more than 

160,000 km2 (Ernst and Buchan, 2001). LIP and rift-related 

igneous rocks exposed in the Lake Superior region include flood 

basalts, mafic intrusions, and minor rhyolite lava flows. The 

Mesoproterozoic Duluth Complex in Minnesota is composed 

of several discrete intrusions formed from mafic to felsic 

magmas that were emplaced between 1,108 and 1,098 Ma 

into older rocks, which include Paleoproterozoic sedimentary 

rocks and Archean granite-greenstone terranes (Peterson and 

Severson, 2002), and coeval LIP and rift-related flood basalts 

and hypabyssal (subvolcanic) intrusions.

Contact-type mineralization that includes a few percent 

sulfide minerals is laterally extensive along the base of 

some of the intrusions (mainly the South Kawishiwi and the 

Partridge River intrusions) along the western margin of the 

Duluth Complex. The mineralized interval of igneous rocks 

along the contact varies from tens to hundreds of meters in 

thickness. Mineral exploration studies since the early 1950s 

indicate that about 4 billion metric tons of mineralized rock 

containing 0.6 weight percent copper, 0.2 weight percent 

nickel, and 0.655 g/t PGEs may be present (Listerud and 

Meineke, 1977; Naldrett, 2010b). Because the proportion of 

sulfide minerals varies along strike, economic cutoff grades 

are used to define deposits along the contact zone.

Hydrothermal and Sedimentary Deposits

Anomalous concentrations of PGEs or the presence 

of platinum-group minerals have been reported in a variety 

of hydrothermal and sedimentary deposit types (table N2 

at back of chapter; Wilde and others, 2003; Wilde, 2005). 

Hydrothermal mineral deposits form by the interaction of 

hot water and rock. Water is a good solvent and hot water, 

called hydrothermal fluid, can efficiently move and deposit 

materials that occur in some mineral deposits. Sedimentary 

mineral deposits are economic concentrations of minerals 

in sedimentary rocks that formed directly from water in 

oceans, lakes, or in underground reservoirs. Although their 

occurrence is tantalizing, “a world-class deposit in which 

the PGE[s] are primary commodities and hydrothermal 

[processes] were indisputably the primary mechanism of 

concentration is yet to be discovered” (Wood, 2002, p. 233). 

Nevertheless, economic deposits of hydrothermal platinum 

and palladium may exist (Wilde and others, 2003).

Experimental and theoretical investigations by many 

authors have helped to identify the constraints under 

which the PGEs can be mobilized by fluids at relatively 

low temperatures (that is, at approximately 300 °C or 

less) (Wood and others, 1992; Gammons and Bloom, 

1993; Wood, 2002; Hanley, 2005; Colombo and others 

2008; Barnes and Liu, 2012; and references cited therein). 

Significant concentrations of platinum and palladium can 

dissolve into solution as chloride complexes only under 

highly oxidizing or highly acidic conditions. Palladium and 

platinum can be transported as bisulfide complexes in acidic 

to neutral solutions under reduced and moderate oxidation 

conditions. Hydroxide, thiosulfate, and organic complexes 

may be the dominant forms of dissolved platinum and 

palladium in low-temperature, near-surface environments.

Chloride complexes play a significant role in the 

mass transfer of PGEs in (a) porphyry copper deposits, 

especially the porphyry copper-gold subtype with island 

arc affinities and those associated with alkaline igneous 

rocks; (b) unconformity-related uranium-gold-platinumpalladium deposits; and (c) sediment-hosted strata-bound 

copper deposits (Wood, 2002). Bisulfide complexes may 

be responsible for (a) transporting platinum, palladium, and 

gold in the footwall copper-nickel-PGE ores at Sudbury, 

Ontario, and in nickel-molybdenum-PGE-gold enriched 

black shale deposits; (b) the remobilization of PGEs or 

recrystallization of platinum-group minerals in ophiolites 

and other ultramafic rocks during serpentinization and 

metamorphism; and (c) the remobilization of PGEs from 

mineralization of primary magmatic origin by deuteric or 

hydrothermal processes (Wood, 2002).

Manganese crusts that precipitate on the sides of 

seamounts are the definitive example of a sedimentary 

deposit enriched in PGEs (table N2). These crusts take 

millions of years to accumulate; the metals are extracted 

from seawater by adsorption processes (Halbach and 

others, 1989).
Geology N19

!

!

0 3 6 MILES

0 3 6 9 12 KILOMETERS

KILOMETERS

VERTICAL SCALE GREATLY EXAGGERATED

DATUM IS SEA LEVEL

0 MILE0 2 KILOMETERS 1

0.5 1

1.5 0.5

110°15' W 110°00' W 109°45' W 45°30' N

45°20' N

44°10' N

B

B

Limestone

Stillwater

Mine

Stillwater River West Fishtail Creek West Fork Stillwater River Boulder River Iron Creek East Boulder Plateau Stillwater Plateau

East Boulder Mine portal

East Boulder River

PGE Fig N13 Stillwater area geology

Dry Fork

thrust

Iron Creek–

Brownlee Creek

fault

A

A

A'

A'

Nye

Unconsolidated depositsPhanerozoic sedimentary rocksArchean quartz monzonite

Upper Banded seriesMiddle Banded seriesLower Banded seriesUltramafic and Basal series

Archean metasedimentary rocks

EXPLANATION

Lower Tertiary intrusive rocksStillwater Complex

Bronzitite zonePeridotite zone and Basal series

Fault (on section only)J–M ReefChromite seam Line of section

 (on section only)

A A'

N

49°30' N

45°0' N

105° W 110° W 115° W

MONTANA

IDAHO WYOMING

ALBERTA

CANADA

UNITED STATES

SASKATCHEWAN

Map area

3.02.52.01.51.0

ted to the north, exposing Geologic map and cross section of the Stillwater Complex, Montana. The layered cumulates that make up the complex have been til Figure N13.

a section through the entire intrusion, from the ultramafic rocks at its base (Ultramafic and Basal series) to mafic rocks (Lower, Middle, and Upper Banded series) exposed 

the Lower Banded series and has been beneath the younger Phanerozoic sedimentary rocks. The platinum-group-element mineralized interval, the J–M Reef, crops out in 

traced for 42 kilometers on strike. Modified from Jones and others (1960) and Zientek and others (2005).
N20 Critical Mineral Resources of the United States—Platinum-Group Elements

Figure N14. Photograph of the Stillwater Mine in south-central Montana, looking southeast. The pond is the 

tailings impoundment. The buildings in the lower left include the shaft, the mill complex, and offices. Photograph 

by Michael L. Zientek, 2006.

Residual Deposits Formed By Weathering 

and Laterites

In areas with tropical or forested warm to temperate 

climates, weathered red subsoil rich in secondary oxides forms 

on ultramafic rocks. These subsoils, or laterites, are a significant source of nickel (the other significant source of nickel is 

magmatic ore deposits). Laterites contain about 70 percent of 

the world’s nickel resources, have been mined for more than 

100 years, and account for about 40 percent of the world’s 

nickel production (Gleeson and others, 2003; Dalvi and others, 

2004; Mudd, 2010; Berger and others, 2011).

PGE enrichments are found in some laterites developed 

on ultramafic rocks, and an ongoing debate among geologists 

concerns whether this enrichment is related to the mobilization 

of the PGEs during the laterization process or to the residual 

accumulation of preexisting platinum-group minerals. Samples 

of lateritized ophiolitic mantle harzburgite from the Pirogues 

River area of New Caledonia contain up to 2 ppm (but average 

500 ppb) platinum. The continuity of the PGE-enriched 

laterite layers relative to the discontinuous distribution of 

PGE-enriched chromitites in the ophiolites initially suggested 

that the PGEs could be concentrated through the lateritization 

process (Augé and Legendre, 1994); however, recent work 

indicates that the PGE-enriched zones are the result of the 

concentration of residual platinum-group mineral particles 

that were present in the unweathered ultramafic rock (Traoré 

and others, 2008). The PGE-enriched saprolitic horizon of the 

Falcondo nickel-laterite deposit in the Dominican Republic 

also appears to be a residue of primary platinum-group 

minerals that were dispersed in ophiolite-related ultramafic 

rocks (Proenza and others, 2010).

Another PGE-enriched laterite occurrence, the Syerston 

nickel-cobalt-platinum deposit, overlies the dunitic core of 

the easternmost of three plugs that make up the Tout Intrusive 

Complex, which is a Uralian-type intrusion in the Fifield area 

in New South Wales, Australia (Teluk, 2001). Preferential 

weathering of the ovoid dunite body formed a slab-like laterite 

body. The laterite extends to a depth of 40 m and has zones 

of nickel-cobalt enrichment that average 10 m in thickness in 

the middle of the weathering profile. An interval of residually 

concentrated platinum largely coincides with the nickel-cobalt 

enriched layers. Primary platinum enrichment in the host 

dunite may explain the distribution of platinum in the Syerston 

laterite profile, the exceptional grades (commonly in excess of 

2.0 g/t platinum), and the coarse grains of the platinum-group 

minerals. Published resources are 137 million metric tons 

containing 0.24 g/t platinum; the planned nickel-cobalt open 

pit area (high-grade goethite zone) accounts for 91.6 million 

metric tons of the resource and contains 0.70 weight percent 

nickel and 0.12 weight percent cobalt (Teluk, 2001).
Geology N21

100° W

50° N

40° N

90° W

MN

WI

IL

IN

KY

MI

OH

SD

ND

NE

KS MO

IA

Animikie Group

Archean rocks

Diabase to gabbro 

 intrusion Other layered series intrusions

Bald Eagle intrusion

South Kawishiwi intrusion

Partridge River intrusion

Anorthositic rocks

Felsic rocks, undivided

Duluth Complex 

Thomson Formation

Biwabik Iron Formation

Ni-Cu-PGE deposit

Ni-Cu-PGE exploration target

North Shore Volcanic Group

Beaver Bay Complex intrusion

EXPLANATION

Mesaba

Northmet

Wetlegs

Serpentine

Wyman

Creek

Nickel

Lake

Dunka Pit 2

Dunka Pit 1

Dunka Pit 4

South Filson

Creek 1 and 2

Dunka Pit 3

Maturi West

Birch Lake

Maturi

Spruce Road

92°00' W 91°30' W

47°30' N

47°45' N

UNITED STATES

CANADA

Map

area

0 4 8 12 16 KILOMETERS

0 4 8 MILES

PGE fig N15 (Duluth_figure V4)

BEi

BEi

oi

BBCi

NSVG

NSVG

NSVG

NSVG

NSVG

BBCi

oi

oi

oi

oi

oi

oi

oi

SKi

SKi

PRi

PRi

Ar

Ar

Ar Ar

Ar

Ar

Ar

Ar

Ar

Fr

Tf

dgi

A

Tf

BIf

BIf

Fr

A

dgi

dgi

Lake

Superior

Lake Michigan

Duluth Complex

Midcontinent Rift 

 mafic igneous rocks

EXPLANATION

Figure N15. Map showing the geology along the western margin of the Duluth Complex, Minnesota, with the 

surface projection of nickel–copper–platinum-group-element (Ni-Cu-PGE) deposits and exploration targets. 

Exploration targets occur along the outer margins of known deposits (Birch Lake and Maturi) or as areas 

separated from known deposits (Maturi West). The location map shows the exposed and covered areas of the 

Midcontinent Rift along with the Duluth Complex. Map based on information in Miller and others (2002) and 

Parker and Eggleston (2012). Geology in the index map is derived from Muehlberger (1997).
N22 Critical Mineral Resources of the United States—Platinum-Group Elements

Platinum Placer Deposits

Erosion and weathering of rocks can form surficial 

mineral deposits. Placer deposits form when heavy particles 

of rocks or minerals are physically concentrated by the 

action of moving water. Placer deposits can form in active 

streams or rivers or along the shoreline of oceans or lakes. 

In 1824, placer platinum deposits were discovered on the 

Orulikha River, north of the Nizhny Tagil Massif in the central 

Ural Mountains (Urals) of Russia (table N3; fig. N16; Duparc 

and Tikonowitch, 1920; Pushkarev, 2001). From then until 

1922, approximately 330 metric tons of metal was recovered 

from placer deposits in the Urals. At least one company 

is still exploring for paleoplacers that are buried beneath 

younger deposits of sediment and for placers that may be 

on river terraces along the Tylay and Kos’va Rivers in the 

Urals (Eurasia Mining plc, 2012). So far, the company has 

delineated resources in placers that contain 0.6 metric tons 

of platinum; average grades range from 0.303 to 0.363 gram 

per cubic meter (Eurasia Mining plc, 2011).

In the second half of the 20th century, platinum placer 

deposits were discovered at Kondyor in eastern Siberia 

(fig. N17; Shcheka and others, 2004) and in the Koryak area at 

the northern end of the Kamchatka Peninsula in eastern Russia 

(Tolstykh and others, 2004). The Kondyor placer deposits 

produced about 85 metric tons of platinum from 1984 to 2011. 

After a decade of mining, approximately 45 metric tons of 

platinum was recovered in the Koryak area (Hurst, 2005).

The detrital platinum minerals in the placer deposits 

in Russia are clearly derived from small mafic to ultramafic 

plutons. The drainages surrounding the Nizhny Tagil Massif 

contained some of the most productive placer deposits in the 

Urals (fig. N16B). Placer mining near Nizhny Tagil began 

around 1839. Platinum in bedrock was first discovered in 1892 

at a deposit called Krutoy Log in the Nizhny Tagil Massif 

(fig. N16B; Anikina and others, 2014). These plutons, which 

are the sources of platinum placer deposits, are circular or 

elliptical in plan view, pipe-like when viewed in profile (that 

is, in a cross-sectional view), and generally range in areal 

extent from about 10 to 40 km2

 (Johan, 2002). The rocks that 

make up the plutons are a core of dunite that is concentrically 

surrounded by borders of clinopyroxenite, hornblendite, 

and gabbroic rocks. The zoned ultramafic complexes have 

rock textures that are consistent with (a) crystallization from 

magma at depth; (b) deformation, recrystallization, and 

annealing at high temperature; and (c) metasomatic alteration 

(Burg and others, 2009; Shmelev and Filippova, 2010). 

Rocks in these small mafic to ultramafic plutons make up a 

distinct rock association called Uralian-type complexes. These 

complexes are enigmatic because lode occurrences of platinum 

mineralization found in the plutons are inconsequential 

compared with the amount of placer platinum minerals found 

in the streams draining from the plutons.

Platinum mineralization in Uralian-type complexes is 

closely associated with chromitite pods, vein-like bodies, or 

schlieren within the dunite core of the plutons. In the Nizhny 

Tagil Massif, more than 1,600 chromitite occurrences have 

been mapped; most do not exceed 50 centimeters (cm) in 

length and a few centimeters in thickness. Two types of 

chromitite are distinguished: (a) massive chromitite that grades 

into disseminated chromite in the dunite host, and (b) chromitite 

marked by serpentine rims that has a sharp boundary with the 

dunite host rock. The second type of chromitite is exceptionally enriched in platinum (Pushkarev and Anikina, 2002). 

Although the chromitites are small, the abundance of platinum 

can be quite high. For example, at the Krutoy Log deposit, 

chromite-platinum ore at the top of the orebody extended for 

a distance of 2 m in one adit. From this small volume of rock, 

about 30 kilograms (reported as 965 troy ounces) of native 

platinum metals was recovered (Mertie, 1969).

The platinum-group minerals in the placer deposits 

provide clues about the rocks from which they are derived. 

The platinum-group minerals are dominated by platinum-iron 

alloys but also contain some iridium and osmium alloys; few 

minerals with palladium or rhodium are found. The platinum 

nuggets in the placer deposits can be quite large (more than 

10 cm in diameter and weighing up to several kilograms); by 

comparison, in most contact- and reef-type PGE ore deposits, 

the platinum-group minerals rarely exceed a few hundred 

microns in diameter (1 micron equals 0.0001 cm) and weigh 

less than 0.0000001 kilograms. Some of the detrital platinumgroup minerals have crystal faces that reflect their internal 

crystalline structure (figs. N18A and N18B). Minerals that 

display crystal faces usually grow from a liquid or a vapor. 

Many nuggets contain inclusions of chromite and olivine 

(fig. N18C), or in some cases, form a matrix interstitial to 

these minerals. These textures suggest that platinum-iron 

alloys formed from magmas that were crystallizing chromite 

and olivine. Collectively, these observations indicate that 

the platinum nuggets are not derived from concentrations 

of magmatic sulfide minerals; rather, it appears that the 

nuggets were crystallized directly from silicate melt or a 

magmatic vapor.

Platinum placers are derived from similar ultramafic 

plutons in Alaska; British Columbia, Canada; Colombia; 

Ethiopia; and New South Wales, Australia (table N3). The 

largest alluvial PGE resource in the United States occurs south 

of Goodnews Bay in Alaska (fig. N19). The Goodnews Bay 

placer deposits were discovered in 1926 and mined between 

1927 and 1975. Approximately 18 metric tons of platinum was 

recovered from the Salmon River and its tributaries (Johnston, 

1962; Mertie, 1976; Barker, 1986; Tolstykh and others, 2002). 

Significant placer resources may remain in the unmined, 

deeply buried ground of the lower Salmon River Valley as 

well as in waste material (tailings) from previous dredging 

operations (Southworth and Foley, 1986).
Geology N23

Table N3. Areas with significant placer platinum production, and estimates of cumulative production and grades.

[g/m3, gram per cubic meter; km, kilometer; n.d., no data; PGE, platinum-group element]

Location

Cumulative 

production 

(metric tons)

Dates Grade(s) Reference(s)

Ural Mountains, 

Russia

330 to 435 1824 to 

present

Initial tenors of placers in the Martian, Syssim, 

Tschauch, and Wyssim Rivers, which drain 

the Nizhniy Tagil area, were about 400 g/m3

; 

by World War I, the grades of placers in these 

rivers were 0.41 to 35 g/m3

.

The average grade of the placers in the Nizhniy Tagil 

district early in its history is estimated to be about 

20 g/m3

.

Grades currently estimated for the West Kitlim placer 

deposit are 0.303 to 0.363 g/m3

.

Lumb (1920); Eurasia 

Mining plc (2011); 

Malitch and Thalhammer (2002); Pushkarev 

and Anikina (2002); 

Pushkarev (2001); 

Teluk (2001)

San Juan and Atrato 

Rivers districts, 

Colombia

107 1778 to 1965 n.d. Mertie (1969)

Kondyor, Russia 85 1984 to 2011 Average grade of the placer deposit is 1.6 g/m3

 PGEs. 

Drainages within the intrusive body have grades of 

4 g/m3

. The grades 50 km downstream are about 

0.5 g/m3

.

Russian Platinum 

(2011a); Shcheka 

and others (2004)

Kamchatka, Russia 45 Approximately 

10 years of 

mining beginning 

after 1991

The average mined grade is 0.28 g/m3 platinum. Hurst (2005)

Goodnews Bay, 

Alaska

19 to 22 1927 to 1975 At Squirrel Creek, high-grade placers contained 

4.1 g/m3, with an average tenor of 1.2 g/m3.

At Clara Creek, early small-scale mining of placers 

with a grade of 3.25 g/m3

, with an average tenor

 of 0.81 g/m3

.

A cleanup a short distance from the mouth of 

Fox Gulch indicated grades of 0.81 to 1.2 g/m3

.

Average grade of 470 samples is 0.192 g/m3

, with 

a maximum value of 1.6 g/m3

.

Johnston (1962); Mertie 

(1976); Barker (1986); 

Southworth and Foley 

(1986); Pacific North 

West Capital Corp. 

(2006); Chatterjee and 

Bandopadhyay (2011)

Yubdo, Ethiopia 2.0 1926 to 1956 Eluvium—Grades were initially 0.3 to 0.5 g/m3

 but 

decreased to 0.11 g/m3

 as mining moved away 

from the core of the intrusion.

Alluvium—Grades were 0.25 to 0.83 g/m3

 in test pits.

Molly (1959)

Tulameen district, 

British Columbia, 

Canada

0.6 1885 to 1932 n.d. Mertie (1969)

Fifield district, New 

South Wales, 

Australia

0.6 1893 to mid1960s; most 

production from 

first few years 

of operation

n.d. Mertie (1969); Platina 

Resources Ltd. (2006)
N24 Critical Mineral Resources of the United States—Platinum-Group Elements

59°36' E 59°40' E

57°44' N

57°40' N

58° E 60° E 62° E

60° N

58° N

Prostokishenka River

Is River

Isovsky

District

Zholtaya Sopka

Massif

Denezhkin

Kamen Complex

Kumbinsky

Complex

Kytlimsky

Complex

Kosvinsky

Massif

Konzhakovsky

Massif

Veresovoborsky

Massif

Svetloborsky Massif

Kachkanarsky

Complex

Pavdinsky

Complex

A

B

Krutoy Log

Tura Riv e r

Area of map B

RUSSIA

KAZAKHSTAN

Area of map A

30° E 60° E 90° E

60° N

50° N

EXPLANATION

Undeformed cover of sedimentary rocks

Gabbro

Pyroxenite and dunite

Orogenic belts and passive continental margin

Orogenic belts

Archean and Proterozoic basement rocks

Platinum-producing area

Sedimentary rocks

Basalt, andesite, 

 and green shale

Fine grained amphibole-

 plagioclase rocks

Amphibolite

Clinopyroxenites 

 and wehrlites

Dunite

Gabbro-pyroxenite

Nizhny Tagil Complex Meta-basalts

Map A Map B

Platinum deposit

PGE fig N16 Urals_NizhnyTigil_v4

Rivers from Petrov and others, 2007

Asia North Albers Equal Area Conic Projection

Central meridian, 60° E. 

Latitude of origin, 30° N.

Nizhny Tagil

Massif

Kytlim

Vissim

area

Is area

Denezhkin

Kamen area

area

0 15 30 45 60 KILOMETERS

0 15 30 MILES

1 2 3 KILOMETERS

0 0.5 1 1.5 MILES

0

Tylay R Kos’va R

Figure N16. Maps illustrating the distribution of platinum deposits in the Ural Mountains, Russia. A, Generalized geology,

location of platinum placer districts, and platinum-producing areas in the Ural Mountains, Russia. Placer platinum 

deposits are derived from pyroxenitic and dunitic intrusions; the platinum-iron alloy derived from the Veresovoborsky 

Massif and the Svetloborsky Massif occurs in placer deposits that extend about 100 kilometers downstream along the 

Is and Tura Rivers. Modified from Zonenshain and others (1988), Anderson and Martineau (2002), and Petrov and others 

(2007). B, Geology of the Uralian-type Nizhny Tagil Massif and the location of the Krutoy Log, where the first bedrock 

source of the platinum deposit was discovered. Platinum minerals were concentrated in chromitites that occur in the 

dunitic core of the lithologically zoned pluton. Modified from Levin and others (2010). 
Geology N25

N

Metasedimentary

hornfels

Placer

Dunite deposits

Clinopyroxenite

N

134°30' E 134°50' E

57°

40'

N

57°

30'

N

A B

MONGOLIA

RUSSIA

CHINA

JAPAN

PACIFIC

OCEAN

120° E 170° E

50° N

30° N

Diffuse koswite and 

 phlogopite-rich stockwork

Felsic intrusions

Gabbro

Clinopyroxenite

Dunite

Metasedimentary hornfels

Basement gneiss and marbles

EXPLANATION

Imagery source: Esri, DigitalGlobe, 

GeoEye, Earthstar Geographics, 

CNES/Airbus DS, USDA, USGS, AEX, 

Getmapping, Aerogrid, IGN, IGP, 

swisstopo, and the GIS User Community

Image from Esri, DigitalGlobe, GeoEye, 

Earthstar Geographics, CNES/Airbus DS, 

USDA, USGS, AEX, Getmapping, Aerogrid, 

IGN, IGP, swisstopo, and the GIS User 

Community

Image courtesy of NASA/GSFC/

METI/Japan Space Systems, and 

U.S./Japan ASTER Science Team

0 2 4 6 KILOMETERS

0 21 3 MILES

PGE Figure N17

Map

area

Figure N17. Geology and imagery of the Uralian-type Kondyor Massif, which is located in eastern Siberia, Russia, north of the city of 

Khabarovsk. A, Geologic map of the massif, showing that the circular pluton consists of a central core of dunite, which is surrounded 

by clinopyroxenite and gabbro. Emplacement of the pluton metamorphosed the surrounding sedimentary rocks, forming hornfels. 

Geology simplified from Burg and others (2009). B, Oblique view of the Kondyor Massif showing the core of dunite surrounded by 

clinopyroxenites. This circular intrusion is about 10 kilometers in diameter; the prominent topographic ridge, which is up to 600 meters 

high, is underlain by contact metamorphosed rocks that surround the intrusion. The intrusion is the source of coarse crystals of 

platinum-iron alloy coated with gold that are recovered in streams that drain from the complex. An example of a coarse platinum 

nugget recovered from the Kondyor Massif is shown in the black-and-white picture. This view was created by draping a simulated 

natural color Advanced Spaceborne Thermal Emission and Reflection Radiometer ASTER composite image over an ASTER-derived 

digital elevation model; the image was acquired on June 10, 2006. Photograph courtesy of NASA/GSFC/METI/Japan Space Systems, 

and U.S./Japan ASTER Science Team. The image of the platinum nugget is from Malitch (1999).
N26 Critical Mineral Resources of the United States—Platinum-Group Elements

PGE Fig N18 (plate 2 from USGS bulletin 193 cropped.ai)

Chromite

Olivine Platinum-iron

alloy

A B

C

Figure N18. Photograph and lithograph showing the morphology of platinum-iron-alloy nuggets derived from 

Uralian-type intrusions. A, Photograph of silvery-gray nugget of isoferroplatinum from the Kondyor Massif, Russia. 

The nugget, which is about 4 millimeters (mm) long, consists of two intergrown crystals of isoferroplatinum; 

each crystal forms a cube with smooth crystal faces and sharp edges and corners (Photograph courtesy of 

TreasureMountainMining.com). B, Sketch of the isoferroplatinum nugget illustrating the cube-shaped crystals. 

C, Lithograph of a platinum nugget with olivine and octahedral chromite, from the Tulameen district, British Columbia, 

Canada. The longest dimension on the nugget is about 10 mm. In this nugget, platinum-iron alloy occurs interstitial to 

(in between) magmatic crystals of olivine and chromite. The texture implies that the platinum-iron alloy grew slightly 

later than the earlier formed olivine and chromite. Image is from Kemp (1902). 
Geology N27

Red Mountain

Kuskokwim Bay

Susie

Mountain

Smalls River

Salmon River

Clara Creek

Platinum Creek

Squirrel

Creek

161°45' W 161°35' W

59°00' N

58°55' N

Bering

Sea

Beaufort Sea

Gulf of

Alaska

0.5 1.5 1.0

KILOMETERS

MILES

0 1 2 3

0

EXPLANATION

PGE fig N19 (Goodnews Bay V6)

Platinum placer tailings

Quaternary deposits, primarily alluvium

Hornblende-dominant mafic and ultramafic rocks

Clinopyroxene-dominant ultramafic rocks

Peridotite

Dunite

Metasedimentary and metavolcanic rocks

Fault 

N

ALASKA

B

162° W 160° W

60° N

59° N

Goodnews Bay ALASKA

0 10 20 30 40 KILOMETERS

0 10 20 MILES

Permissive tract

Overlap assemblage, 

 Kuskokwim Group

Togiak

Nyack

Goodnews Bay

Kilbuck

Terranes

EXPLANATION

Area of

map below

Bering

Sea

A

Figure N19. Maps illustrating platinum-group-element (PGE) resources in southeastern Alaska. A, Location of the 

permissive tract assessed for undiscovered placer platinum resources and the major geologic provinces (terranes) it 

encompasses. The permissive tract is the area where undiscovered PGE resources could be present (Schmidt and others, 

2007). B, Geologic map of the Goodnews Bay area showing where platinum placer deposits have been mined (platinum 

placer tailings). The placer platinum deposits can be followed upstream to the Red Mountain pluton, a Uralian-type intrusion 

that could be the bedrock source of the platinum minerals found in the placers. Modified from Tolstykh and others (2002). 
N28 Critical Mineral Resources of the United States—Platinum-Group Elements

Resources and Production

Discovered mineral resources include the total amount 

of identified mineral resources that are in the ground as 

well as the cumulative past production (U.S. Geological 

Survey National Mineral Resource Assessment Team, 

2000). This section begins with an overview of historic and 

current primary and secondary production of PGEs. Primary 

production is when metal that occurs in rocks is mined 

and transferred to above-ground material stock; secondary 

production is from recycled material or scrap. The section 

also includes a discussion of the identified mineral resources 

of PGEs. Identified mineral resources refer to the mineralized 

material in the ground whose location, grade, quality, and 

quantity are known or can be estimated from specific geologic 

evidence (U.S. Bureau of Mines and U.S. Geological Survey, 

1980). Identified mineral resources represent the current stock 

of material in the ground (mineral inventory) that might be 

mined. The amounts of identified resources discussed this 

chapter were current as of 2012.

Production

Up to 1920, almost all of the world’s PGE production 

came from placer deposits in Russia and Colombia. The 

discovery of nickel deposits in solid rock in the Sudbury 

area of Ontario, Canada, in the 1800s (Giblin, 1984) roughly 

coincided with technological innovations that used nickel to 

strengthen traditional steels. This technological innovation 

led to extensive development of the Sudbury deposits; in the 

first half of the 20th century, Sudbury produced approximately 

80 percent of the world’s nickel (Prevec, 1997). PGEs were 

recovered as a byproduct of nickel mining in Sudbury; 

by 1934, Canada had become became the world’s leading 

producer of PGEs.

Most of the PGEs produced today are from mineral 

deposits that were discovered in the Noril’sk-Talnakh area 

of Russia in 1919 (Kunilov, 1994; Likhachev, 1994) and 

in South Africa in the 1920s (Wagner, 1929). Significant 

development of these deposits did not begin until the 1960s 

when industrial demands for PGEs increased. About the same 

time that new and extraordinarily PGE-enriched deposits were 

discovered in the Noril’sk-Talnakh area in Russia (Kunilov, 

1994), Anglo American Platinum Ltd. gained a controlling 

stake in the PGE industry in South Africa (Mattera, 2008)

and South Africa had an economic boom (Rouillard, 1997). 

As a result, beginning in the 1960s, PGE production increased 

in South Africa and the Soviet Union (Russia), such that 

South Africa became, and still is, the leading producer of 

platinum and rhodium, and Russia was and is the world’s 

leading producer of palladium (fig. N20).

Approximately 14,200 metric tons of PGEs was 

produced from 1900 to 2011, and roughly 95 percent of that 

amount (about 13,500 metric tons) was produced from 1960 

onward. The breakdown of production by country shows that 

about 90 percent of production since 1900 has come from 

South Africa and Russia; Canada, the United States, and 

Zimbabwe accounted for 5 percent, 2 percent, and 1 percent 

of production, respectively (fig. N21).

A secondary supply of platinum, palladium, and rhodium 

is obtained through the recycling of catalytic converters 

from end-of-life vehicles, jewelry, and electronic equipment. 

Recycling volumes are sensitive to PGE prices; higher prices 

typically lead to higher recycling volumes owing to the greater 

incentive to recycle. Recycled platinum, palladium, and 

rhodium provide a significant proportion of the total supply 

and are sufficient to close the gap between mine production 

and consumption. For example, in 2011, about 24 percent 

Others

Zimbabwe

United States

Canada

Russia/U.S.S.R.

South Africa

EXPLANATION

Year

1960 1970 1980 1990 2000 2010

0

100

200

300

400

500

600

Platinum-group-element production, in metric tons

Figure N20. Graph showing world platinum-groupelement production, by country and year, from 1960 to 2011. 

The layers of the graph are placed one above the other, 

forming a cumulative total. Data are from U.S. Bureau 

of Mines (1933–34, 1933–96) and U.S. Geological Survey 

(1997 –2016).
Resources and Production N29

South Africa

7,049 t Russia

5,046 t

Canada

727 t

United States

321 t

Zimbabwe

125 t

Others

254 t

Figure N21. PGE fig N21Pie chart showing world platinumgroup-element (PGE) production from 1960 to 2011, 

by country and amount (in metric tons [t]) . Data are 

from U.S. Bureau of Mines (1933–34,1933–96) and 

U.S. Geological Survey (1997–2016). 

of the total supply of platinum and palladium and about 

27 percent of the total supply of rhodium were obtained 

through recycling (Butler, 2012).

Identified Resources

Exploration and mining companies have delineated 

approximately 100,000 metric tons of PGE (and minor 

amounts of gold) resources in mineral deposits around the 

world that remain to be developed (table N4). This estimate 

was derived by compiling mineral inventory information about 

in situ estimates of the tonnage and grade of mineralized rock 

with “realistic prospects of eventual economic extraction” 

(Weatherstone, 2008) published by exploration and mining 

companies and geological surveys. To be considered an 

identified resource for this compilation, the evaluation of 

tonnage and grade of mineralized rock must be based on 

(a) direct sampling of the ore, and (b) industry standard 

resource estimation practices. The estimate of 100,000 metric 

tons is comparable to the 91,000 metric tons recently 

reported by Mudd (2012) (table N4). Most of the difference 

is accounted for by the larger estimate for the Platreef deposit 

of the Bushveld Complex in this chapter. For context, the 

total net consumption of PGEs in 2012 was approximately 

460 metric tons (Platinum Today, 2013c).

Table N4. Identified platinum-group-element and gold resources, summarized by deposit type and location.

[Data are from table N10 at the end of this chapter. Numbers may not add to totals because of rounding]

Deposit 

type Deposit

PGE and gold resources 

compiled for this chapter

(metric tons)

PGE and gold resources 

compiled by Mudd, 2012 

(metric tons)

Reef Merensky Reef and UG2 Chromitite, Bushveld 

Complex, South Africa 

58,000 56,000

Main Sulphide Zone, Great Dyke, Zimbabwe 8,200 8,700

J–M Reef, Stillwater Complex, Montana 2,200 620

Other areas 2,700 1,700

Total, reef-type deposits 71,000 67,000

Contact Platreef, Bushveld Complex, South Africa 17,000 7,700

Other areas 3,100 1,600

Total, contact-type deposits 20,000 9,300

Conduit Noril’sk-Talnakh area, Russia 10,000 11,000

Other areas 1,200 2,100

Total, conduit-type deposits 11,000 13,000

Other All other areas 990 1,200

Grand total, all three types of deposits 100,000 91,000
N30 Critical Mineral Resources of the United States—Platinum-Group Elements

Tonnage and Grade Relations for Magmatic 

Ore Deposits

Tonnage and total PGE and gold grade relations for identified resources in magmatic deposits are shown in figure N22. 

Estimates of total PGE and gold grade can vary over about 

three orders of magnitude, whereas tonnage can vary over 

five orders of magnitude. Tonnage and grade are not strongly 

correlated, which is typical for mineral deposits of a given 

type. Some conduit-type deposits have lower PGE and gold 

grades than reef-type deposits because conduit-type deposits 

are polymetallic deposits in which copper, nickel, PGEs, and 

gold all contribute to the value of the ore. In reef-type deposits, 

the value of the ore is primarily in the PGE and gold concentrations. In terms of contained metal, most deposits contain 

more than 1 metric ton of PGEs and gold. For any values less 

than 1 metric ton, it simply is not economic to determine the 

tonnage and grade. In 2012, world mine production was about 

450 metric tons of PGEs and world usage was about 615 metric 

tons (Loferski, 2013b). The shaded area in figure N22 

highlights those deposits that have more than 1 year’s worth 

of contained PGEs, given current rates of production and 

usage. Only a fraction of the deposits in the world have enough 

contained metal to satisfy multiple years of world demand.

Mineral deposits occur rarely in Earth’s crust, and 

large ones are especially uncommon (Singer, 1995). 

Most of the known metal for many mineral commodities 

is contained in a few, very large deposits (Singer and 

DeYoung, 1980). For most types of mineral deposits, the 

grade of the ore varies by about one order of magnitude and 

the tonnage varies by several orders of magnitude among 

deposits. As a result, contained metal correlates strongly 

with tonnage. For porphyry copper deposits (which are the 

world’s leading source of copper), for example, 10 percent 

of the deposits accounts for approximately 60 percent of the 

identified copper resource whereas the bottom 50 percent 

of the deposits accounts for less than 5 percent of the 

identified copper resources (fig. N23A). For the PGEs, 

this correlation is more pronounced. Ten percent of PGE 

deposits account for more than 80 percent of identified 

PGE resources whereas the bottom 50 percent of the PGE 

deposits accounts for less than 1 percent of identified PGE 

resources. Thirty percent of PGE deposits contain more 

than 97 percent of the world’s identified PGE resources, and 

almost all these deposits are associated with three areas—

the Bushveld Complex in South Africa, the Great Dyke 

in Zimbabwe, and the Noril’sk-Talnakh area in Russia 

(fig. N23B).

0.01

0.1

1

10

Concentration of platinum-group elements and gold, in grams per metric ton

0.1 1 10 100 1,000 10,000

Deposit tonnage, in million metric tons

100 t

10,000 t

1,000 t

0.1 t

1 t

10 t

20

Conduit

Other

Reef

Deposit type

Contact

EXPLANATION

Bushveld Complex

Great Dyke

Noril'sk-Talnakh

Stillwater Complex

Wellgreen

Syerston

Intrusion (designated 

 by fill color)

Eagle Nickel

PGE fig N22 (Bivariate tonnage vs grade_v2)

Figure N22. Plot showing the relation between tonnage and grade of remaining resources for conduit-type, 

reef-type, and other types of deposits enriched in platinum-group elements (PGEs). Diagonal lines are isolines 

that show amounts of contained PGEs and gold in the deposits, in metric tons (t). Annual world production and 

usage of PGEs, as of 2012, was about 450 metric tons. The shaded area on the figure highlights those deposits 

that are large enough to contain more than 1 year’s worth of the world’s PGE needs.
Resources and Production N31

0

20

40

60

80

100

Percent of contained metal

0 10 20 30 40 50 60 70 80 90 100

Percent of deposits

EXPLANATION

Porphyry copper deposits

PGE deposits

10% of deposits

contain 

approximately 

60% of resource

10% of deposits contain 

approximately 85% of resource 

50% of deposits contain 

less than 1% of resource

50% of deposits contain

less than 5% of resource

0

20

40

60

80

100

Percent of contained metal (PGE and gold)

0 20

Percent of deposits

EXPLANATION

Noril'sk-Talnakh

Great Dyke

Duluth Complex

Stillwater Complex

Other deposits

Bushveld Complex

10 30

A

B

Area of figure B

Figure N23. Graphs showing the percent of contained metal against percent of deposits for A, the 

world’s platinum-group-element (PGE) and porphyry copper deposits, and B, the top 30 percent of the 

world’s PGE deposits. In A, the curves are derived from compilations of known deposits and show that 

most of the contained metal in a given deposit type is associated with a small percentage of very large 

deposits. Both deposit types show the effect, but it is more pronounced for PGE deposits. In B, the 

curve focuses on the largest deposits, showing the data points and the name of the igneous intrusion 

hosting the deposit. Most of the large deposits are associated with the Bushveld Complex in South 

Africa, the Great Dyke in Zimbabwe, and sills in the Noril’sk-Talnakh area of Russia. Data for porphyry 

copper are from Singer and others (2008). PGE data are from table N10 at the end of this chapter.
N32 Critical Mineral Resources of the United States—Platinum-Group Elements

Seventy-two percent of the world’s PGE resources are 

found in reef-type and contact-type deposits in the Bushveld 

Complex, South Africa (table N4). The conduit-type magmatic 

deposits in the Noril’sk-Talnakh area, Russia, account 

for 10 percent of the world’s identified PGE resources. 

Eight percent of the world’s identified PGE resources are in 

the Great Dyke, Zimbabwe. In fact, more than 97 percent 

of all the world’s identified PGE resources are contained in 

only 14 intrusions or intrusive complexes (table N5 at back 

of chapter).

When considered by deposit type alone, 69 percent 

of identified PGE resources are associated with reef-type 

deposits, 20 percent are associated with contact-type deposits, 

and 11 percent are associated with conduit-type deposits 

(table N4). For reef-type deposits, 81 percent of identified 

resources have been delineated in the Bushveld Complex 

(South Africa); 12 percent in the Great Dyke (Zimbabwe); 

and 3 percent in the Stillwater Complex (Alaska). For 

contact-type deposits, 85 percent of the world’s identified 

resources are found in the Bushveld Complex. For conduittype deposits, about 90 percent of the identified resources 

are found in the Noril’sk-Talnakh area of Russia.

For simplicity, the previous discussion has focused on 

total PGE production and resources; however, individual 

PGEs do have different values and uses. For example, 

platinum and rhodium are more valuable economically than 

palladium, and the iridium market is much smaller than the 

markets for platinum and palladium. The proportions of the 

PGEs differ among deposits, however (fig. N24). The deposits 

in South Africa and Zimbabwe have a higher proportion 

of platinum than any other major deposit in the world. The 

proportion of palladium is much higher in the deposits in the 

Noril’sk-Talnakh area in Russia.

Undiscovered (Hypothetical and Speculative) 

Resources (United States and Global)

In addition to the resources of PGEs that have been 

discovered (identified), additional PGEs may be present as 

undiscovered resources in areas adjacent to identified magmatic 

deposits and in deposits that remain to be discovered.

Extensions to Identified Deposits

Additional PGE resources may be present in areas 

adjacent to or near deposits that now dominate world supply. 

Although resources for reef-type and contact-type PGE 

deposits are formally delineated in advance of mining, 

the extent of the deposits is never completely delineated 

because the upfront costs would be too great. Instead, mining 

companies conduct enough research and work to ensure 

that the deposit contains adequate resources to justify mine 

development. It is possible, therefore, that some deposits are 

larger than previously thought.

In the Bushveld Complex, South Africa, relatively 

unexplored extensions of the Merensky Reef and the 

UG2 Chromitite deposits could contain an additional 

33,000 metric tons of platinum and 32,000 metric tons of 

palladium to a depth of 3 km (Zientek and others, 2014). 

These reef-type PGE deposits are in the form of vast sheetlike orebodies. Mining has started at the surface and has 

progressed to depths of 2 km in some regions (fig. N25). 

Likewise, the Platreef in the Bushveld Complex could 

contain an estimated 1,100 metric tons of platinum and 

nearly 1,400 metric tons of palladium (based on a mean 

estimate to a depth of 1 km). The Great Dyke of Zimbabwe 

Figure N24. Graph showing contained 

platinum-group element (PGE) and gold metal 

against the ratio of palladium to platinum 

for the major PGE deposits of the world. The 

only identified deposits in which platinum is 

the dominant metal are associated with the 

Bushveld Complex in South Africa and the Great 

Dyke in Zimbabwe. PGE data are from table N10 

at the end of this chapter. MR, Merensky Reef; 

Pd, palladium; Pt, platinum, UG2, UG2 Chromitite

0

5,000

10,000

15,000

20,000

Total contained PGE and gold, in metric tons

0 1 2 3 4 5

Ratio of palladium to platinum

Bushveld Complex

Duluth Complex

Great Dyke

Noril'sk Talnakh

Stillwater Complex

Intrusion or area

EXPLANATION

Pt-dominant Pd-dominant

Talnakh

Sheba’s

Ridge

War

Springs

MR-East

MR-West

UG2-East

UG2-West

Platreef deposits

Pseudo

Reef

Noril’sk

PGE fig N24 (total contained PGM by Pd-Pt)
Resources and Production N33

Approximate

northeast

Merensky Reef surface trace

11

2.0

1.0

0.5

0

1.5

0 1 2 3 4 5 6 7 8 9 10

KILOMETERS

PGE fig N25 (BV_3D)

 KILOMETERS

EXPLANATION

0

2

Merensky Reef mined-out area

Interpolated depth of Merensky Reef,

 in kilometers

20°E 30°E

20°S

30°S SOUTH

AFRICA

BOTSWANA

ZIMBABWE

Indian Ocean

Merensky

Reef

Figure N25. Three-dimensional block diagram showing the Merensky Reef interpolated down to 2 kilometers in the southern 

area of the western limb of the Bushveld Complex, South Africa. Mined out areas are shown in gray, and the Merensky Reef 

trace is shown as a black line. The Merensky Reef continues at depth beyond what is shown in this illustration.

could contain up to 6,900 metric tons of undiscovered 

platinum, palladium, and rhodium (Zientek and others, 2014).

Additional PGE resources in the Noril’sk-Talnakh area in 

Russia are difficult to assess. The extent of the orebodies for 

which resources are reported is not revealed in publicly accessible technical information released by the mining company. 

The northernmost mine in the Talnakh area, the Skalisty Mine, 

appears to be midway along the axis of the intrusion that hosts 

the massive sulfide orebodies (fig. N7A). If existing maps, 

similar to the one shown in figure N7A, are approximately 

correct, additional resources associated with massive sulfide 

orebodies in the Talnakh area could be present.

Roughly 305 metric tons of platinum and palladium 

have been produced from the J–M Reef in the Stillwater 

Complex in Montana, and drill-based estimates indicate that 

another 2,200 metric tons are present (U.S. Bureau of Mines, 

1933–96; U.S. Geological Survey, 1997–2016; Abbott and 

others, 2011). Mining has progressed to depths of 1,800 m 

below the surface, but the bottom of the ore deposit has not 

been reached. Rough geologic estimates suggest that another 

1,000 to 6,200 metric tons of platinum and palladium could 

be present at depth (Zientek, 1993; Zientek and others, 

2002). As with the Bushveld Complex, the J–M Reef has 

been traced for longer distances along strike than down dip. 

The downdip projection of the J–M Reef is complicated 

by structural deformation associated with the uplift of the 

Beartooth Mountains. Gravity anomalies and inclusion 

suites in younger, cross-cutting intrusions, however, demonstrate that a portion of the complex is concealed at depth 

(Brozdowski, 1985; Kleinkopf, 1985). As a result, structural 

and geophysical models would be needed to constrain what 

rocks may be present at depth.
N34 Critical Mineral Resources of the United States—Platinum-Group Elements

Undiscovered Reef-Type and Contact-Type Deposits

Based on literature research, there are at least 200 layered 

intrusions in the world, although the actual number could 

be two to three times higher. An intrusion does not have to 

be large to have a reef-type deposit; intrusions as small as 

a few tens of kilometers in length with as little as 500 m 

of layered cumulates are known to host reef-type mineralization (table N6 at back of chapter; fig. N26). Well-explored 

intrusions have at the most 1- or 2-m-thick reef-type deposits 

(and several subeconomic occurrences) that occur in thousands of meters of layered rock. Many small intrusions could 

be evaluated for their reef-type potential to assess whether 

an economic deposit could be present.

For example, the identified PGE resources in Finland 

are 91 metric tons of platinum and 237 metric tons of 

palladium in seven reef-type and contact-type deposits. An 

assessment by the Geological Survey of Finland estimates 

that a mean number of 29 and 23 contact-type and reef-type 

deposits, respectively, occur in Finland. They estimate that 

the median contained metal contents would be 5,500 metric 

tons of platinum and 12,000 metric tons of palladium in the 

undiscovered contact-type and reef-type deposits (Rasilainen 

and others, 2010).

Undiscovered magmatic mineral deposits could be 

present in many areas of the United States and the world. The 

intrusions of the Duluth Complex, United States, and other 

smaller intrusive complexes of the Midcontinent Rift in the 

Lake Superior region have high potential to contain undiscovered copper-nickel-PGE sulfide deposits (fig. N15; Miller 

and others, 2002). In fact, recent exploration has dramatically 

expanded the mineral resource inventory of contact-type 

deposits along the western margin of the Duluth Complex. 

Exploration results have led to new exploration models, 

which indicate additional areas that may have the potential for 

undiscovered deposits. Furthermore, the Duluth Complex has 

many large layered intrusions that may have the potential to 

host reef-type PGE deposits. New models, new data, and tools 

for integration and visualization could result in discoveries of 

mineralized rocks.

Ernst and Buchan (2001) cataloged more than 300 LIP 

events worldwide; they determined that more than 100 LIP 

events have associated sill complexes or layered intrusions. 

All are prospective for the occurrence of magmatic mineral 

deposits. In the United States, Ernst and Buchan identified 

about 25 LIP events, of which several have sill complexes and 

layered intrusions. The obvious exploration targets for reeftype deposits in the United States are the Stillwater Complex 

in Montana and the dozens of intrusions that make up the 

Duluth Complex in Minnesota and associated rocks of the 

Keweenawan LIP. Other layered mafic intrusions with reeftype potential occur in the Wichita Mountains of Oklahoma, 

100

1,000

10,000

30,000

Stratigrahic thickness, in meters

0.1 1 100 1,000 10,000

Areal extent, in square kilometers

Intrusions with platinum-group-element (PGE) reef-type 

 deposits, chromitite layers, or magnetite layers

Intrusions with no known mineralization

Approximate extent of

 intrusions with PGE

reef-type deposits, 

 chromitite layers, 

or magnetite layers

EXPLANATION

50

10

PGE fig N26 (Intrusion area vs thickness) Figure N26. Graph illustrating the exposed area and 

stratigraphic thickness of cumulates in more than 

200 intrusions from around the world. 

which are part of a Cambrian LIP. The Lake Owens Complex 

in Wyoming, which is a small layered intrusion, is also 

reported to have reef-type mineralization (Loucks, 1991). 

Other large layered intrusions are known from subsurface 

oil and gas exploration in west Texas and New Mexico; at 

present, they are too deep to be targets for mineral exploration 

(Keller and others, 1989; Kargi and Barnes, 1995; Adams and 

others, 1997).

To have a significant effect on global supply, a 

large-tonnage deposit would first have to be discovered. 

Large-tonnage reef-type deposits can occur only in large 

layered intrusions. A survey of the literature shows there are 

several large intrusions that do not have identified reef-type 

mineralization (such as the Tete Complex, the Dufek intrusion, 

the Molopo Farms Complex, the Windimurra Complex, and 

so on) (table N6). Some of the intrusions are poorly exposed 

and others have been intensely deformed and metamorphosed, 

which makes them difficult to explore for deposits. The 

economic risk associated with exploring these intrusions 

would be relatively high compared with identifying additional 

resources in extensions of known reef-type deposits.
Resources and Production N35

Undiscovered Conduit-Type Deposits in the 

United States—Nikolai Greenstone and 

Eastern Wrangellia

Several LIPs with sill complexes occur in the 

United States (Ernst and Buchan, 2001); however, one 

in particular stands out for its potential for undiscovered 

conduit-type nickel-copper-PGE mineralization (Hulbert, 

1997; Hulbert and Stone, 2006; Schmidt and Rogers, 

2007). The Wrangellia LIP is known from extensive flood 

basalts and related rocks of the Middle to Late Triassic 

Nikolai Greenstone that is part of the Wrangellia terrane 

accreted to the western margin of North America (Berg 

and others, 1972; Jones and others, 1972; Berg and others, 

1978; Nokleberg and others, 1994; Plafker and Berg, 1994). 

The Wrangellia terrane can be traced discontinuously for 

more than 2,000 km from the Wrangell and Amphitheater 

Mountains in southern Alaska through Canada’s southwestern 

Yukon Territory, to the Queen Charlotte and Vancouver Islands 

in British Columbia, Canada.

Sill-like mafic to ultramafic intrusive complexes in 

Wrangellia are the remnants of subvolcanic magma chambers 

that fed the thick, overlying flood basalts and picritic 

pyroclastics of the Nikolai Greenstone. A conduit-type 

deposit in the Wellgreen intrusion, Yukon Territory, contains 

a resource of 461 million metric tons grading 0.32 weight 

percent nickel, 0.26 weight percent copper, 0.4 ppm platinum, 

0.34 ppm palladium, and 0.17 ppm gold (McCracken, 2011; 

Carter and others, 2012).

Alaska’s Amphitheater Mountains and the area of southcentral Alaska are also underlain by several mafic to ultramafic 

complexes that are part of the Nikolai LIP (fig. N27; Schmidt 

and Rogers, 2007; Glen and others, 2011). The largest and 

most exposed mafic to ultramafic sills include the Fish Lake 

and Tangle (FL–T) Complexes, which occur along the flanks 

of the Amphitheater Mountains synform. Smaller mafic to 

ultramafic complexes include Canwell, Eureka, and Rainy. 

The occurrence of picritic volcanic rocks (high-magnesium 

olivine basalt) near some of the larger mineralized ultramafic 

bodies is consistent with mineralized occurrences that have 

high nickel-to-copper ratios and elevated PGE concentrations. 

Grab samples from three different mafic to ultramafic 

complexes have extremely high metal concentrations, as 

follows: 2.0 weight percent nickel, 1.3 weight percent copper, 

and 7.9 g/t PGE+gold+silver from Rainy; 13.6 weight percent 

nickel, 2.9 weight percent copper, and 26.0 g/t PGEs from 

Canwell; and 6.7 weight percent nickel, 1.0 weight percent 

copper, and 1.5 g/t PGEs from Alpha (same as Fish Lake) 

(Pure Nickel, Inc., 2013).

Undiscovered Platinum Placer Deposits in Alaska

Undiscovered marine placer deposits could be associated 

with paleochannels or beach deposits in the Goodnews Bay 

region in Alaska. A geophysical survey indicates the presence 

of ultramafic rocks offshore, which could be a source of placer 

platinum (Barker and others, 1988; Barker and Lamal, 1989). 

Seismic analysis indicates that the bottoms of the paleochannels that drained these now-offshore ultramafic rocks are 

about 20 to 40 m deep and covered with modern sediments 

(Oommen, 2006; Oommen and others, 2008). Platinum may 

have accumulated in marine sand bars that are aligned parallel 

or subparallel to the current Alaskan coastline (Oommen, 

2006; Oommen and others, 2008). Assuming that the sand bars 

represent submerged paleobeaches exposed during periods of 

glaciation in the Pleistocene Epoch, intensified wave sorting 

and concentration of heavy minerals (including entrained 

platinum grains) may have taken place. Spatial distribution 

patterns of platinum and sediments offshore of the Goodnews 

Bay Complex suggest relatively high platinum concentrations 

near Goodnews Bay and Carter Bay (Oommen, 2006; 

Oommen and others, 2008).

Schmidt and others (2007) delineated a permissive tract 

(fig. N19A) for placer PGE-gold deposits hosted in Quaternary 

unconsolidated alluvial and marine sediments that includes 

onshore regions and areas up to 10 km offshore that are 

downstream of known mafic to ultramafic complexes within 

the Goodnews Bay and Togiak terranes (Hoare and Coonrad, 

1978; Decker and others, 1994). They estimate that the tract 

has a 90 percent probability of containing at least 1 metric 

ton of platinum and a 10 percent probability of containing 

at least 17 metric tons; the mean and median estimates are 

5 and 7 metric tons, respectively.

Metals in Stored Waste Products

PGEs may also be present in material discarded during 

the mining and beneficiation of magmatic ore deposits. When 

ore materials are processed to recover the PGEs, some losses 

always occur; in other words, it is impossible to recover 

100 percent of the metal in the mined material. For example, 

tailings derived from processing ores from the Noril’skTalnakh area in Russia may contain 800 to 1,000 metric 

tons of PGEs; an additional 100 metric tons of PGEs may 

be present in stored pyrrhotite concentrates and iron cakes 

(Petrov and others, 2013). Some of the metals in stored waste 

products could be recovered if appropriate technologies are 

developed. For example, ultrafine grinding techniques are 

currently being used to reprocess tailings from the Merensky 

Reef mines in South Africa (Buys and others, 2013).
N36 Critical Mineral Resources of the United States—Platinum-Group Elements

146°20' W 146°00' W 145°40' W

63°20' N

63°10' N

Fish Lake

Complex

 Rainy Complex

Canwell

Complex

Tangle

Complex

Eureka Complex 

Tangle

Lakes

Fielding

Lake

Landmark

Gap Lake Glacier

Lake

Delta River

Fish

Lake

Summit Lake

Creek

Phelan

Sevenmile Lake

Trng

Trnb

Trnb

Trng

Trng

Trnb

Trng

Trnu

Trnu

Trnu

Trnu

Trnb

Trnb

Trnu

PTar

Q

Q

Q

Q

Q

Q

Trng

Trnb

Trnb

Q

Q

Trnu

Trnb

HIGHWAY

HIGHWAY

DENALI

Eureka Glacier

Canwell Glacier

PGE fig N27 (Nikolai Belt from Geological Society of America Special Papers-2007-Schmidt-2007.2431(24)_v2)

0 2 4 6

0 2 4 6 8 10 12 KILOMETERS

MILES

60° N

70° N

RICHARDSON

Nickel–copper–platinum-group-

 element prospect

Undifferentiated Quaternary deposits and

 Tertiary to Pennsylvanian bedrock

Trnb Triassic Nikolai basalt

Trng Triassic mafic (gabbroic) rocks

Trnu Triassic ultramafic rocks

EXPLANATION

PTar Permian-Triassic argillaceous rocks 

Q

CANADA

UNITED

STATES

ALASKA

Map

area

180° 150° W 120° W

Figure N27. Geologic map of the Amphitheater Mountains and south-central Alaska showing the location and names of 

mafic-ultramafic complexes that are part of the Nikolai large igneous province (Schmidt and Rogers, 2007). 
Environmental Considerations N37

Exploration for New Deposits

The methods of locating PGE deposits can vary. In very 

rare cases, PGE-enriched sulfide mineralization has been 

found by panning. For example, the Merensky Reef of the 

Bushveld Complex, South Africa (fig. N9), was discovered 

in 1924 as the result of panning in a dry river bed and finding 

platinum minerals in the concentrate (Cawthorn, 1999). 

In most cases, chemical analysis of samples is required to 

identify PGE-enriched rocks. Because of the refractory nature 

of PGEs and their low concentration in most materials, an 

analytical technique that preconcentrates PGEs before analysis 

and ensures complete digestion of the sample is required. The 

preferred method to analyze for PGEs is fire assay.

A practical and effective exploration approach is to look 

for anomalous concentrations of copper, platinum, and palladium in residual or transported material derived from a larger 

volume of rock, such as soil, stream sediments, and talus fines 

(Cameron and Hattori, 2005). Soil chemistry has been used 

to delineate mineralized rocks in the Platreef of the Bushveld 

Complex (Frick, 1985). The J–M Reef of the Stillwater 

Complex (fig. N13) was discovered in 1974 by analyzing 

soil and talus fines for platinum and palladium (Conn, 1979; 

Zientek and others, 2005). Contact-type mineralization in the 

Duluth Complex (fig. N15) is currently being explored using 

regional geochemical surveys of glacial till (Duluth Metals 

Ltd., 2011).

Surface-water geochemistry led to the discovery of 

the ore deposits in the Noril’sk-Talnakh area in Russia 

(figs. N3 and N6; Kunilov, 1994). A surface water 

geochemical survey showed elevated concentrations of SO4

2–

in samples from the Talnakh River. This led to the discovery 

of mineralized boulders in the area, which were traced to the 

base of the Kharayelakh Mountains. A study of talus along the 

slope of Otdel’naya Mountain led to the discovery of mineralized outcrops.

Geophysical methods used in PGE exploration map 

physical property contrasts, which primarily indicate the 

presence of magnetite, silicate minerals, and sulfide minerals 

that may be associated with PGE mineralization (Balch, 

2005). If the PGEs are associated with base-metal sulfide 

minerals that have some degree of interconnection in the 

rock (and are electrically conductive), airborne and ground 

electromagnetics and induced polarization surveys can be used 

to detect mineralized rocks. High-resolution aeromagnetic 

surveys can be used to map igneous layering and tectonic 

structures (Campbell, 2006) but generally do not give direct 

indication of mineralized rock. Gravity studies are used to 

determine the subsurface extent of rocks with variable density 

and are particularly well suited to map and model the extent 

and volume of mafic and ultramafic igneous rocks (Webb 

and others, 2004). Once a rock layer that contains reef-type 

mineralization has been identified, seismic studies can be 

used to map the subsurface extent of the rocks (Davison and 

Chunnett, 1999; Chunnett and Rompel, 2004).

Lithogeochemical studies can be used to focus exploration activities (Maier and Barnes, 2005). For example, changes 

in the ratio of PGEs to sulfur or copper in rocks can be used 

to target specific intervals in cumulate stratigraphy where 

reef-type PGE mineralization could be present. In layered 

intrusions, nickel depletion in olivine is thought to indicate 

indirectly the presence of magmatic sulfide mineralization. 

Metal depletion in lavas flows is thought to indicate the 

presence of sulfide mineralization in the sills and dikes that 

served as feeders to the flows.

High-density drilling and sampling are required to 

delineate mineral resources and reserves in PGE deposits. 

For the J–M Reef, drilling on 15-m spacing is used to define 

proven mineral reserves; probable mineral reserves are 

delineated by projecting data 300 m from drill holes (Abbott 

and others, 2011). In the reef-type deposits in the Bushveld 

Complex, South Africa, measured mineral resources are 

defined with holes spaced 250 to 300 m apart, indicated 

resources are defined with holes spaced 500 to 600 m apart, 

and inferred mineral resources are defined with holes spaced 

800 to 2,000 m apart. By comparison, massive sulfide ores 

in the Noril’sk-Talnakh area, Russia, are explored on a 

32- by 60-m rectangular grid (Kunilov, 1994).

Environmental Considerations

The mineralogy of the ore and waste material associated 

with reef-type, contact-type, and conduit-type deposits are 

similar. Common sulfide minerals include chalcopyrite, 

pentlandite, and pyrrhotite. Secondary phases that typically form from primary sulfide minerals include covellite, 

goethite, gypsum, jarosite, marcasite, native sulfur, and 

pyrite (McGregor and others, 1998; Johnson and others, 

2000). Common primary gangue minerals include chromite, 

olivine, plagioclase, and pyroxene. Alteration products include 

chlorite, magnetite, serpentine, talc, and minor amounts of 

carbonate and clay minerals.

The abundance of sulfide minerals defines the environmental and geologic characteristics of PGE-enriched 

magmatic sulfide deposits. The high proportion of sulfide 

minerals distinguishes some conduit-type deposits from reefand contact-type deposits. The overall pyrrhotite content of 

ore, waste rock, and tailings determines their acid-generating 

potential. Acid generated by the oxidative weathering of 

pyrrhotite can attack associated ore sulfide minerals and 

gangue silicate minerals, releasing a variety of potentially 

toxic elements—particularly cobalt, copper, iron, manganese, 

nickel, and, to a lesser extent, aluminum, cadmium, chromium, 

lead, and zinc to solution or to secondary solid phases 

(McGregor and others, 1998; Johnson and others, 2000). 

Massive ores characteristic of some conduit-type deposits 

have greater than 50 volume percent sulfide minerals and 

pose a greater challenge for mining and processing the ores 

without causing environmental problems. Both the low-sulfide 
N38 Critical Mineral Resources of the United States—Platinum-Group Elements

reef-type and contact-type deposits have significantly lower 

sulfide contents, usually less than a few volume percent. 

Therefore, acid generation in mine wastes is less of a challenge to manage.

Sources and Fate in the Environment

The abundance of PGEs in the upper continental crust 

ranges from tens to hundreds of parts per trillion (ppt); 

platinum is the most abundant, followed by palladium, 

osmium, ruthenium, iridium, and rhodium (table N7; Taylor 

and McLennan, 1995; Park and others, 2012). Background 

PGE concentrations are also low in other environmental 

media, such as water, sediment, soil, and plants.

Experimental studies suggest that the maximum solubility of platinum in water at room temperature (25 °C) under 

oxygenated conditions ranges between 0.02 and 195 ppb, 

depending upon the pH of the water (Wood, 1991; Azaroual 

and others, 2001). The solubility of palladium in water at 

room temperature (25 °C) under oxygenated conditions ranges 

between 1 and 3,400 ppb, depending upon the pH of the water 

(Wood, 1991). Because of its greater solubility, palladium is 

more mobile in the environment than platinum.

PGE concentrations in water are generally in the 

10 –2 ppt range. Rhodium (0.082 ppt) has the highest mean 

concentration in seawater, followed by palladium (0.062 ppt), 

platinum (0.026 ppt), and osmium (0.005 ppt); ruthenium 

and iridium are present well below the ppt range (Bruland 

and Lohan, 2003). Platinum concentrations range from 

0.22 to 78,000 ppt in surface water and from 3 to 38 ppt 

in groundwater (Ravindra and others, 2004).

Environmental baseline studies generally show low 

concentrations of PGEs and trace elements in water near 

undeveloped deposits. In the vicinity of the contact-type 

NorthMet deposit in northern Minnesota, surface water 

downstream of the deposit in the Partridge River watershed 

has neutral pH (7.0), a low dissolved sulfate concentration 

(7.2 ppm), and moderate alkalinity (31 ppm calcium carbonate 

[CaCO3]) and hardness (30.3 ppm CaCO3). Dissolved platinum 

and palladium concentration are below the detection limit 

(<25 ppb), and the concentrations of other trace elements are 

low (table N8 at back of chapter; Minnesota Department of 

Natural Resources and the U.S. Army Corps of Engineers, 

2009). Near the buried Eagle deposit, which is a conduit-type 

deposit in northern Michigan, maximum concentrations of iron 

(190 ppb), nickel (59 ppb), and zinc (88 ppb) in groundwater 

near the deposit reflect the presence of the orebody, whereas 

maximum concentrations of copper (<5 ppb) and cobalt 

(<10 ppb) are below their detection limits. In contrast, the 

concentrations of copper, nickel, and other trace elements in 

surface waters over the buried deposit are indistinguishable 

from those from regional water samples (Kennecott Eagle 

Minerals Co., 2006).

In fluvial and pelagic-marine sediment away from any 

known deposits, total PGE concentrations are less than a few 

tens of ppb (table N7). Platinum and palladium are found in 

higher concentrations than are (in decreasing order of abundance) ruthenium, iridium, osmium, and rhodium.

Fuchs and Rose (1974) investigated the geochemical 

behavior of platinum and palladium in the weathering cycle in 

soils around the Stillwater Complex in Montana. They found 

that palladium was depleted in the surface soil horizons but 

was concentrated at depth, reaching a maximum concentration 

of approximately 40 ppb. Platinum did not show systematic 

trends with depth; a maximum concentration of approximately 

180 ppb was recorded. The differences in the behavior of 

palladium and platinum in soils reflects the higher solubility 

of palladium in water compared with that of platinum in water 

(Fuchs and Rose, 1974).

Table N7. Platinum-group-element concentrations in samples of upper crust, loess, river sediment, and marine pelagic sediment.

[n.d., no data]

Material

Platinum Palladium Ruthenium Iridium Osmium Rhodium

Reference(s) (part per billion)

Upper continental 

crust

0.599 0.526 0.03 0.022 0.05 0.018 Taylor and McLennan (1995); 

Park and others (2012)

Loess (China) 0.751 0.546 0.03 0.023 n.d. 0.018 Park and others (2012)

River sediment 

(Mölndal River, 

Sweden, and 

Stour River, 

England)

<0.29 to 

53.9

0.08 to

38.7

<0.15 to 

3.73

<0.03 to 

2.69

n.d. <0.11 to

9.4

Ravindra and others (2004);

Moldovan and others (2001); 

de Vos and others (2002)

Marine pelagic 

sediment

0.4 to

21.9

1.2 to

9.4

< 0.2 to

2.2

<0.2 to

1.2

<0.05 to 

0.81

n.d. Goldberg and Koide (1990)
Environmental Considerations N39

The results of an exploration program illustrate the 

natural background variation of platinum and palladium 

in a layered intrusion that hosts a major reef-type deposit. 

Following the studies by Fuchs and Rose, more than 

10,900 samples of soil and talus fines were collected and 

analyzed for platinum and palladium as part of the exploration 

program that led to the discovery of the J–M Reef at the 

Stillwater Complex (Conn, 1979; Zientek and others, 2005). 

Palladium was below the detection limit for 14 percent of 

the locations; almost one-half of the samples had platinum 

concentrations that were below the detection limit. For most of 

these samples, the detection limit for palladium and platinum 

was 2 ppb and 20 ppb, respectively. The median values for 

samples in the study above the detection limit were 8 ppb and 

20 ppb for palladium and platinum, respectively. Maximum 

values measured for palladium and platinum were 6.4 ppm and 

5.76 ppm, respectively.

Platinum and palladium data are characterized by 

positive-skewed, multimodal distributions. More than 

95 percent of the samples belong to a population consisting of 

unmineralized material with less than 30 to 40 ppb palladium 

and less than 40 to 50 ppb platinum. The remaining approximately 460 samples appear to be made up of two populations. 

The range of values for the largest population in the remaining 

5 percent varies from about 40 ppb to 200 ppb palladium and 

about 50 to about 200 ppb platinum. About 50 samples make 

up the third population and consist of material with more than 

200 ppb platinum or palladium.

Biogeochemical surveys show that platinum and 

palladium can concentrate in plants near mineralized outcrops. 

Riese and Arp (1983) conducted a geochemical orientation 

survey of the J–M Reef in the Stillwater Complex and 

found accumulations of platinum in stems of Douglas-fir 

(Pseudotsuga menziesii) up to thousands of parts per billion 

in ashed samples. In the northern forests of Saskatchewan, 

Canada, platinum and palladium tend to be concentrated in 

the twigs and trunks of black spruce (Picea mariana) and 

jack pine (Pinus banksiana) and in the stems of Labrador tea 

(Rhododendron groenlandicum). Spruce was sampled close to 

a worked-out nickel-copper deposit in northern Saskatchewan 

that contained 3,000 ppb platinum and 6,000 ppb palladium. 

The ashed twigs yielded up to 880 ppb platinum and 1,350 ppb 

palladium compared with background levels of below 10 ppb 

platinum and 2 ppb palladium (Dunn, 1986). Coker and others 

(1991) found greater enrichment of palladium than of platinum 

in vegetation around several PGE deposits in Canada—some 

that had been mined and some that had not been mined.

Anthropogenic sources of PGEs in the environment 

include catalytic converters used in modern automobiles (Ek 

and others, 2004; Ravindra and others, 2004; Wiseman and 

Zereini, 2009), platinum-based chemotherapy drugs (Ravindra 

and others, 2004), and smelter emissions (Chen and others, 

2009). During the release of exhaust gases from automobile 

engines, the catalytic converters are chemically and physically 

stressed and release PGE-containing particulate matter. Most 

of the PGEs in the particles are in the metallic state, but about 

1 percent is in an oxidized form that could be bioavailable 

(Ravindra and others, 2004). Concentrations of PGE particles 

are found in roadside dust, soils, and water. In platinumbearing drugs that are used to treat cancer, the platinum is 

excreted by the patients following treatment and ends up in the 

hospital sewage. The contribution of PGEs to the environment 

through this method is minor, however, compared with that 

from catalytic converters (Ravindra and others, 2004). Tiny 

amounts of osmium (0.25 to 23 femtograms per gram of water; 

1 femtogram is 10 –15 grams) are found in rainwater and snow 

from around the world (Chen and others, 2009). The most 

likely sources of this metal are smelters that process PGE ores. 

The osmium concentrations are so low that they are difficult to 

analyze; however, it is possible that the global appearance of 

osmium could someday be used as an environmental tracer.

Platinum-Group-Element Mining and MineralProcessing Methods

Selective underground mining techniques are used for 

PGE deposits that occur in narrow reefs, such as those in the 

Bushveld Complex (South Africa) and the Stillwater Complex 

(Montana). The underground mining is performed by using 

either handheld pneumatic drills or mechanized drills to bore 

holes in the ore horizon, which are then filled with explosives. 

After blasting, the ore is removed from the mining stope and 

transported to the surface for further processing. Open pit 

mining is used for deposits in which the PGEs occur in greater 

thicknesses and at shallow depth, such as in the Lac des Iles 

Mine in Canada and the Platreef of the Bushveld Complex.

After ore is removed from the mines, it is crushed and 

ground to reduce the particle size and free the PGE-containing 

minerals from the rock matrix. The ore is then concentrated 

by using froth flotation circuits. For flotation, the ground rock 

particles are mixed with water and various reagents and air 

is pumped through the liquid, creating bubbles to which the 

PGE-containing minerals adhere. These bubbles float to the 

surface of the flotation vats and are removed as froth. The 

material is reground and refloated to produce a concentrate 

of nickel-iron-copper-PGE sulfide minerals for further 

processing. The greatest losses of PGEs occur in the early 

stages of processing, including crushing, milling, and froth 

flotation, owing to the diversity of PGE mineralogy. Much 

research is devoted to increasing PGE recovery rates in these 

early stages.

The concentrate is dried and then smelted in an electric 

arc furnace at temperatures that can exceed 1,500 oC. During 

smelting, the concentrate forms a liquid, which separates into 

two layers. A matte containing the valuable metals is separated 

from a silicate- and oxide-rich slag, which is discarded. At 

the Stillwater Mining Co. smelter, the gases released from 
N40 Critical Mineral Resources of the United States—Platinum-Group Elements

smelting operations undergo a gas-to-liquid scrubbing process, 

which removes approximately 99.8 percent of the sulfur 

dioxide. The sulfur dioxide is converted to gypsum, which is 

sold to local farmers for use as a soil amendment.

After smelting, the PGE-rich matte is transferred to 

converters, through which air is blown to remove iron and 

sulfur. The converter matte is processed at the base-metals 

refinery to separate the base metals from the PGEs by selective 

leach processing. The final stage is PGE refining by various 

hydrometallurgical techniques, including solvent extraction, 

precipitation, and dissolution using chloride solutions, in 

which the six PGEs are separated from each other to a purity 

of more than 99.99 percent. The refined PGEs can be produced 

in various forms, including as ingots, grains, or a fine powder 

known as “sponge.”

The chromium-rich ores of the Bushveld Complex’s 

UG2 Chromitite present special difficulties for beneficiation. 

The high chromium content of the ores requires especially 

high temperatures for smelting, and the low sulfide content 

both increases the matte viscosity and lowers the amount of 

matte formed. A processing technique, known as ConRoast, 

has been employed by some Bushveld processors specifically for use with high-chromium PGE ores. The ConRoast 

technique removes sulfur from the concentrate by roasting the 

ore at high temperatures in oxidizing conditions followed by 

smelting in a direct-current arc furnace in reducing conditions, 

which uses an iron-based alloy as a collector of cobalt, copper, 

nickel, and PGEs. The ConRoast technique can be used for 

concentrates containing any amount of chromite (Jones, 2002).

Ores from the Duluth Complex (Minnesota) present 

another challenging processing problem. Their low-grade 

nature makes smelting inefficient and uneconomic. A hightemperature chloride-assisted pressure-leaching process, 

known as PLATSOLTM, has been developed to process the 

Duluth ores. The PLATSOLTM process is an alternative to 

smelting and is capable of processing low-grade PGE ores. 

The nickel-copper-PGE ore concentrates are processed in 

an autoclave in an oxidizing environment in which PGEs, as 

well as gold, are solubilized and can be recovered directly 

after leaching by adsorption or precipitation with sulfide ions. 

Finally, base-metal sulfides are recovered either by precipitation, ion exchange, or solvent extraction-electrowinning 

(Baxter and others, 2005).

Mine Waste Characteristics

The majority of solid mine waste includes waste rock, 

mill tailings, slag, and smelter dust. Waste rock is uneconomic 

rock that must be removed to access the ore and is disposed 

of on site. It can also be used for construction on site if tests 

determine that it will not generate acid-rock drainage when it 

is exposed to the atmosphere and water. Tailings, which is the 

waste material from processing ore, can be pumped as slurry 

or trucked dry to a tailings storage facility on site. Tailings 

storage facilities are typically impoundments surrounded by a 

retaining dam. Tailings can also be disposed of underground 

in mined-out portions of the mine. The crushing and milling 

process increases the porosity of the solid waste and thus its 

volume by a factor of approximately two, which means that 

only about one-half of the waste from ore processing can be 

returned underground. Smelter slag, which is glassy waste 

matter separated from metals during the refining of ore, also 

has both physical and chemical characteristics of environmental concern and is typically held in storage facilities. Slag 

is also used in some mining and construction applications. 

Smelters produce gaseous emissions that may contain particulate material (smelter dust). If the emissions are not treated 

before the gases are vented to the atmosphere, the particulate 

material will settle near the smelter. For most modern smelting 

facilities, the gaseous emissions are treated to remove particulates, which are reprocessed to recover metals or put into a 

storage system.

Waste rock and tailings are usually found at all mines, 

but their proportions vary based on the mining method. The 

amount of waste rock is significantly greater at open pit mines 

than at underground mines. Reef-type deposits are generally 

mined by underground methods because of the narrow width 

of the ore horizon. Contact-type and conduit-type ores are 

mined by either open pit or underground methods, depending 

upon the geometry and depth of the orebodies.

Acid-Base Accounts

Metals and many other trace elements tend to be more 

soluble at low pH than at neutral or high pH. Therefore, the 

acid-generating or acid-neutralizing potentials of waste rock, 

tailings, and other solid waste material help determine the 

possible environmental risk of mineral resource development. 

The balance between the acid-generating and the acidneutralizing potential of rocks and mine waste is expressed as 

an “acid-base account” (Price, 2009; International Network 

for Acid Prevention, 2011). The release of acid is primarily 

related to the presence of pyrrhotite in the rocks and waste. 

The presence of carbonate minerals, such as calcite, and some 

silicate minerals, such as feldspar, olivine, or pyroxene, in 

rocks and wastes can neutralize acidic solutions. Acid-base 

accounting values are commonly expressed in terms of 

kilograms of calcium carbonate per metric ton (kg CaCO3

/t) of 

waste material.

Acid-Base Accounts for Reef-Type and 

Contact-Type Deposits

Acid rock drainage from reef-type and contact-type 

deposits is unlikely because the ores and their host rocks 

contain low proportions of sulfide minerals. Jambor and others 

(2000) proposed a threshold of 0.3 weight percent sulfur 

present in sulfide minerals for acid generation. The sulfur 

content of ore in the J–M Reef at the Stillwater Complex 
Environmental Considerations N41

in Montana ranges from 0.13 to 0.49 weight percent, with 

concentrations in tailings ranging from 0.05 to 0.08 weight 

percent. The sulfur content of the waste rock ranges from 

0.01 to 0.04 weight percent (table N9 at back of chapter). 

Similarly, the ores and host rocks for the UG2 Chromitite, 

the Merensky Reef, and the Platreef deposits have low sulfur 

concentrations (Wilson and Chunnett, 2006; Naldrett and 

others, 2009). Acid-base accounting studies, leach tests, and 

site monitoring demonstrate that there is limited potential 

for the generation of hazardous metal leachates from ores, 

waste rocks, or tailings from most reef-type and contact-type 

deposits (table N9).

Acid-Base Accounts for Conduit-Type Deposits

For some conduit-type deposits with massive ores, the 

ore is processed to separate out and produce concentrates of 

iron-, copper-, and nickel-bearing sulfide minerals; the concentrated material is then further processed to extract the copper 

and nickel metals, and the iron-bearing sulfide minerals, 

mainly pyrrhotite, are discarded as waste. This process 

results in waste material with high acid-generating potential; 

the presence of minor amounts of carbonate minerals in the 

ore assemblage and the lower reactivity of the neutralizing 

minerals, such as olivine and pyroxene, offer minimal shortterm acid-neutralizing potential. The Eagle deposit in northern 

Michigan has the potential to generate acid leachate similar to 

that found in mineralogically similar tailings from Sudbury, 

Ontario, Canada (McGregor and others, 1998; Johnson and 

others, 2000). Johnson and others (2000) documented the 

generation of low pH (down to 3) waters with high dissolved 

concentrations of iron (up to 9,800 ppm), sulfate (up to 

2,400 ppm), aluminum (up to 1,130 ppm), and nickel (up to 

698 ppm) in groundwater in the tailings pile at the Nickel Rim 

Mine at Sudbury, Ontario, Canada. Copper (up to 3.5 ppm) 

and cobalt (up to 2.5 ppm) were also significant constituents.

Air Quality Assessments

An air quality assessment of the heavily industrialized 

western part of the Bushveld Complex, where the Merensky 

Reef and the UG2 Chromitite are mined, found that sulfur 

dioxide, nitrous oxide, and carbon monoxide concentrations 

are at acceptable levels using South African and European 

air quality standards (Venter and others, 2012). The major 

contributing sources were high-stack industry emissions for 

sulfur dioxide (metallurgical complexes and smelters), and 

household combustion for nitrous oxide and carbon monoxide. 

The levels of ozone and PM10 frequently exceeded standards 

(PM10 refers to the total mass concentration of particulate 

matter up to 10 micrometers [μm] in size). Ozone is related 

to regional sources (wildfires, and coal and wood combustion 

for cooking and heating). The source of PM10 was identified 

as local household combustion (Hirsikko and others, 2012; 

Venter and others, 2012).

Mining operations in the Noril’sk-Talnakh area of 

Russia emit large amounts of sulfur dioxide and nickel and 

other metals (Council on Ethics, 2009; Bellona Foundation, 

2010). According to the Norwegian Government’s Council on 

Ethics for the Government Pension Fund-Global, atmospheric 

emissions of sulfur dioxide, which lead to acid rain, are about 

2,000,000 metric tons per year, and atmospheric emissions 

of copper, nickel, and cobalt are conservatively estimated to 

be about 500, 450, and 50 metric tons per year, respectively. 

Smelting operations release metals into the atmosphere in 

the form of dust or particulates. Smelter emissions have been 

shown to cause death or significant damage to vegetation up 

to 200 km from the mining operations.

Human Health Concerns

The metallic forms of PGEs are generally considered 

to be inert. Health hazards specifically related to PGEs 

affect only individuals who are occupationally exposed to 

manmade PGE compounds, especially workers in preciousmetal refineries. Platinum-based drugs, such as cisplatin and 

carboplatin, are used in cancer treatment. Broader human 

health effects of PGEs appear to be limited because of the low 

concentrations of PGEs in the environment (Hoppstock and 

Sures, 2004). Ravindra and others (2004) and Hoppstock and 

Sures (2004) reviewed human health risks associated with 

PGEs, and Kielhorn and others (2002) reviewed the human 

health effects of palladium specifically. The more significant 

human health risks associated with PGE mining are related to 

the trace elements associated with PGE deposits.

As discussed in a previous section, the mining operations 

at the Noril’sk-Talnakh area in Russia affect air quality; 

surface and groundwater are also affected by the industrial 

activities. Respiratory diseases and various forms of cancer 

have been shown to be more prevalent in the local population 

of the Noril’sk-Talnakh area than in other regions of Russia 

(Council on Ethics, 2009). Industrial activity in this area has 

been shown to be related to increased incidences of acute 

respiratory illness and a higher prevalence of oncological 

disease, particularly lung cancer, and may be related to other 

possible negative health effects, such as weakening of the 

immune system, reproductive health issues, an increase in 

the rate of disease in children, and reduced life expectancy 

(Council on Ethics, 2009).

Ecological Health Concerns

The risks to the ecosystem from mining PGEs and 

associated minerals are mainly those related to acid mine 

drainage, which primarily affects aquatic environments. The 

surface water chemistry downstream from platinum mine 

waste dumps in Zimbabwe was investigated by Meck and 

others (2006). The surface water in the vicinity of platinum 

dumps was found to be near neutral or to have slightly alkaline 

pH and to contain concentrations of chromium, cobalt, copper, 
N42 Critical Mineral Resources of the United States—Platinum-Group Elements

lead, and nickel that exceeded chronic surface-water criteria 

established by the U.S. Environmental Protection Agency 

(Suter, 1996; U.S. Environmental Protection Agency, 2009). 

Differences between upstream and downstream sample sites 

were most noticeable for nickel and approached a difference 

of 500 ppb.

Ecological risks can also be associated with soils and mill 

tailings. The toxicity of rehabilitated and nonrehabilitated

mill tailings from reef-type deposits in South Africa to earthworms (Eisenia fetdia) was investigated by Jubileus (2008) 

and Maboeta and others (2008) using laboratory bioassay 

techniques. Sublethal impairment of earthworms (that is, 

effects on the growth and hatching success of cocoons) in 

soil samples decreased with distance from the tailing storage 

facilities. Anomalous concentrations of chromium, copper, 

nickel, and zinc were found in the tailings relative to the 

regional background soils. Wahl and others (2012) reached 

similar conclusions based on mesofauna surveys from the 

same tailing storage facilities.

Studies of soil and grass show anthropogenic platinum 

enrichment near mines in the Bushveld Complex. Thousands 

of soil and stream sediment samples collected at the Bushveld 

Complex show anomalous concentrations of precious metals 

related to bedrock sources and mining areas (Wilhelm and 

others, 1997). The results of the soil and stream sediment 

surveys mapped two bedrock sources of PGEs that are 

concordant to magmatic layering in the Bushveld Complex. 

The survey also mapped large anomalies near the locations 

of the large PGE mines in the western part of the Bushveld 

Complex; anomalies near the mines exceed 200 ppb PGEs. 

Rauch and Fatoki (2013) found about 700 ppb platinum in 

soils at a smelter. The maximum concentration in grass was 

256 ppb platinum, again found at the smelter.

Carbon Footprint

The PGEs and their mining have several links to carbon 

cycling in the environment. One of the most important uses 

of PGEs is in catalytic converters (Loferski, 2012a, b), which 

help to limit the amount of carbon and nitrogen gases and 

other air pollutants that enter the environment through vehicle 

exhaust. Fuel cells are an emerging energy technology. Proton 

exchange membrane (PEM) fuel cells use hydrogen as fuel, 

combining it with oxygen to form water; these fuel cells rely 

on platinum catalysts (Mehta and Cooper, 2003).

The trend towards the mining of lower grade ore in 

existing mines is expected to have a negative effect on 

greenhouse gas emissions. Lower grade ores require more 

energy to produce a unit of PGEs than do higher grade ores. 

The trend towards increased greenhouse gas emissions per 

unit of PGEs produced (measured in tons of carbon dioxide 

[CO2 ] per kilogram of PGEs) over time is evident in data from 

South Africa that date back to 2002 (Glaister and Mudd, 2010). 

In 2002, unit greenhouse gas emissions varied from about 

20 to 40 metric tons for different mine properties. In 2008, the 

range of values increased to about 30 to 65 metric tons.

The calcium- and magnesium-rich silicate minerals 

common in mine wastes from PGE deposits represent an 

important sink for atmospheric carbon dioxide because these 

minerals react fairly readily with this greenhouse gas. The 

potential for mineral carbonation of PGE tailings was evaluated by Vogeli and others (2011). They ranked South African 

PGE tailings in terms of their carbonation potential on the 

basis of their mineralogy and the grain size and found that 

finer-grained, more olivine-rich, and plagioclase-poor tailings 

would be best for mineral carbonation. Although the mass 

of PGE tailings globally is inadequate to reduce global 

atmospheric carbon dioxide concentrations significantly, 

the carbonation of calcium- and magnesium-rich tailings 

can contribute to reducing the carbon footprint of individual 

mining operations.

Mine Closure

The procedures used to close PGE mines depends 

primarily on the method of mining and the character of the 

waste material. Reef-type deposits are typically developed 

using underground methods. Contact-type and conduit-type 

deposits can be developed by underground or open pit methods, 

depending upon the geometry of the orebody. For all these 

deposit types, open pit mining results in at least three different 

features after mining: the open pit, waste rock piles, and tailings 

storage facilities. Because of the scale of typical PGE mining 

activities, backfilling open pits is usually not practical. If the 

water table is above the bottom of the pit, the pit will fill with 

water and form a lake. The water quality of the lake will depend 

upon a number of factors, including the nature of the wall rock, 

the extent of interconnected underground mine workings that 

open to the pit, the water level in the pit, the local hydrology, 

and the climate, among others (Castendyk and Eary, 2009).

The long-term fate of tailing storage facilities depends 

upon the nature of the tailings and the method of construction 

of the facility. Some facilities can be regraded, capped, and 

revegetated. Others are designed to have a water cover in 

perpetuity to limit sulfide oxidation. Either type may have 

seepage that may require some form of water treatment.

The long-term fate of waste rock piles may include 

regrading, capping, and revegetation. Depending upon the 

acid-generating potential of the material, some piles may 

require some form of water treatment. An emerging issue for 

historical PGE tailings and piles is that older ore-processing 

techniques were not as efficient as those used today. As a 

consequence, historical tailings and waste piles can carry 

significant PGE grades.
Problems and Future Research N43

Problems and Future Research

The supply of PGE from primary and secondary sources 

currently meets society’s demand for these metals. Even if 

demand increases (as it assuredly will), the in-ground identified mineral resources appear to be sufficient to supply the 

world’s needs for several decades. The location of most of 

the in-ground resources in a few big deposits in a few places 

makes the primary production of PGE susceptible to supply 

disruption, however. The following section discusses this 

problem in more detail.

Future Demand and Possible Risks to Supply

The demand for PGEs continues to increase. PGE use 

in automobile catalytic converters is expected to increase as 

more-stringent automobile emission standards are adopted 

worldwide. An emerging industrial use for platinum is in fuel 

cells that provide electricity for automobiles, homes, and businesses. Fuel cells produce electricity through electrochemical 

reactions by combining oxygen with a hydrogen-bearing fuel 

over a catalyst, such as platinum, and produce only water and 

carbon dioxide as emissions.

Various studies have compared anticipated demand for 

the PGEs with the amount of PGE ore that has been positively 

identified by mineral exploration. Along with the anticipated 

supply of PGEs by recycling, the studies suggest that there 

are sufficient PGE resources in the ground to meet projected 

platinum demand well into the middle of the 21st century 

(Tiax LLC, 2003; Wilburn and Bleiwas, 2004; Mudd, 2012; 

Wilburn, 2012). Mineral reserves in the Bushveld Complex 

would be sufficient to meet global platinum demand until 

2040 (assuming an annual increase in platinum consumption 

of 2 percent). The much larger volume of mineralized rock 

that has been classified as mineral resources, coupled with the 

potential for additional undiscovered resources to be found, 

indicates that the potential amount of PGEs that could be mined 

will not be a constraint to PGE supply for many decades. 

Disruptions to the PGE supply are more likely to be affected 

by social, environmental, political, and economic factors rather 

than geologic issues or resource depletion (Mudd, 2012).

Variations in the annual average prices of the PGEs 

during the past 40 years illustrate the types of events and 

policies that could affect global supply and demand for 

mineral commodities (fig. N28). Some events, such as the 

oil embargo in the mid-1970s and the global recession that 

began in 2008, affected all metal prices (not just the prices 

of PGEs). Other effects can be specifically related to legislation passed by one or more Governments; for example, in 

the mid-1970s, catalytic converters, which use PGEs as the 

catalyst, were installed in automobiles to meet air standards 

set in the Clean Air Act Amendments of 1970 (84 Stat. 1705, 

P.L. 91–604; Gerard and Lave, 2005; McCarthy and others, 

2011). Catalytic converters reduce harmful emissions from 

automobiles; the widespread adoption of catalytic converters, 

first in the United States in the 1970s, then in the European 

Union and Japan, increased demand for PGEs. Palladium 

supply was disrupted in 1999 and 2000 because Russian 

Government legislation temporarily blocked export of this 

metal (United Nations Conference on Trade and Development, 

undated). Other effects are related to problems with mineral 

production; for example, problems with a refinery in 

Rustenburg, South Africa, in 1989 caused a decrease in the 

world supply of rhodium. The PGE supply was also affected 

by work stoppages and miners’ strikes in South Africa in 1986, 

2011, and 2012 (Yager and others, 2012).

Production of PGEs requires power and water, both 

of which are in short supply in southern Africa. Africa 

depends on imports of oil and on production of synthetic 

fuels from coal to meet its fuel requirements (United Nations 

Conference on Trade and Development Secretariat, 1995). In 

January 2008, the South African mining industry briefly shut 

down almost all its operations because of the unpredictability 

of the power supply. Roughly one-quarter of the installed 

generating capacity was not available owing to system 

faults, planned maintenance, and a shortage of the coal used 

in power stations. The country was subject to short-notice 

blackouts, which made mining unsafe (Platinum Today, 2008). 

Expanding the mining capacity at the Bushveld Complex in 

South Africa is constrained by the power supply. Water is 

required to mine, process ore, and refine metals; if the water 

supply is restricted, then production is affected. In any country 

that is water-stressed, mining companies must strive to 

manage the supply and use of water to ensure continuation of 

operations (Anglo American Platinum Ltd., 2012).

The ultimate constraint on the development of PGE 

mineral resources may not be the presence of mineralized 

rock but rather the rock temperature. For example, the contact 

between the Main zone and the Lower Critical subzone in 

the Bushveld Complex can be traced to depths of 6 km in 

seismic surveys (Sargeant, 2001; Campbell, 2011), which 

indicates that mineralization associated with the Merensky 

Reef and the UG2 Chromitite may also be present at those 

depths (Cawthorn, 2010). These reefs are currently being 

mined at depths exceeding 2 km at the Northam Mine, where 

virgin rock temperatures of 70 °C are measured at a depth 

of 2,176 m (Northam Platinum Ltd., 2008). Anglo American 

Platinum Ltd., whose operations include several mines in 

the Bushveld Complex, considers a virgin rock temperature 

of 75 °C to be the limit of mining based on the anticipated 

technology, metal prices, and energy costs (Anglo American 

Platinum Ltd., 2011). By comparison, rock temperatures 

within the ore-bearing intrusions at the Talnakh area of Russia 

are up to 35 oC in the Taimyrsky Mine, 27 to 30 oC in the 

Skalisty Mine, and are anticipated to be 43 to 47 oC in the 

area that will be developed by the Gluboky Mine (Kunilov, 

1994). The massive sulfide ores in these Russian mines are 

highly reactive, however, and are prone to oxidation and 

spontaneous combustion.
N44 Critical Mineral Resources of the United States—Platinum-Group Elements

0

500

1,000

1,500

2,000

2,500

1870 1880 1890 1900 1910 1920 1930 1940 1950 1960 1970 1980 1990 2000 2010 2020

Year

0

500

1,000

1,500

2,000

Metal value, in 2010 dollars per troy ounce

0

500

1,000

1,500

0

2,000

4,000

6,000

1,000

3,000

5,000

0

1,000

2,000

3,000

4,000

0

200

400

600

800

100

300

500

700

Fig N28 (PGE yearly prices.ai)

World War I World War II Clean Air Act

of 1970

End of Cold War

a

c

d e

a

b

b a

c

f

Pt

Pd

Rh

Ir

Ru

Os

Figure N28. Graphs showing platinum-group-element prices (in constant U.S. dollars referenced to 2010) for platinum (Pt), 

palladium (Pd), rhodium (Rh), iridium (Ir), ruthenium (Ru), and osmium (Os) from 1880 to 2013. Major world events affecting pricing 

are shown as vertical dashed lines. Other major events shown are (a) investor speculation, (b) catalytic converter use begins 

and the Arab oil embargo, (c) shortfall of supplies of palladium from Russia, (d) disruption of supply because of problems at the 

Rustenburg refinery in South Africa, (e) increased demand for autocatalysts in Asian markets, and (f) increased ruthenium usage in 

electronics, especially in computer hard disk drives. National deflators are from the California Department of Finance (2012). Price 

data are from Plunkert and Jones (1999); Reese (1996); Hilliard (1998, 2000, 2004); Loferski (2012b, 2013c); and Platinum Today (2012). 
References Cited N45

Research Directions

Successful exploration strategies for PGEs have been 

based on the magmatic deposit paradigm. This approach has 

proven highly successful and will likely result in additional 

discoveries. Its utility has a down side, however, in that other 

geologic settings in which PGEs may be concentrated have 

not been studied. Models strongly influence the observations 

made in both research and mineral exploration. For example, 

most economic geologists would not look for world-class 

PGE deposits in ophiolites. The most common magmatic ore 

type in ophiolites are chromitite deposits; all contain PGEs 

but are enriched in iridium relative to platinum (Mungall, 

2002; Finnigan and others, 2008). There are a few examples 

of platinum-enriched chromitites in ophiolites, but these are 

too small to be of any economic consequence. Therefore, 

mineral exploration geologists have not searched for concentrations of magmatic sulfide minerals in ophiolites; however, 

a small magmatic sulfide deposit in the Acoje ophiolite, 

Philippines (Bacuta and others, 1990; Yumul, 2001), and 

occurrences in cumulates of the Shetland Ophiolite Complex, 

United Kingdom (Prichard and Lord, 1993), are examples that 

indicate that sulfide exsolution and PGE enrichment can take 

place in ophiolites. The examples of hydrothermal ore deposits 

with elevated PGEs are other reminders that an economic PGE 

deposit may be present in rocks that have not been explored.

Most researchers in the field of economic geology can 

supply a list of research topics that could be undertaken to 

alleviate concerns about the global availability of PGEs. One 

approach to addressing concerns is to find more deposits, but 

finding small deposits will not solve the problem. The deposits 

must be large (like one or more of the giants that currently 

dominate supply). Research to find new deposits can be basic, 

in which known deposits are mapped and characterized in 

order to understand how they form. The research can also be 

applied, in which the search for PGE deposits is extended 

into geologic settings where the deposits would be highly 

weathered, highly metamorphosed, or largely concealed. 

Additionally, research can focus on new ways to understand 

and integrate information so that new hypotheses can be 

proposed and tested. Research will be most effective if basic 

and applied research is integrated. It will also be more fruitful 

if expertise from many disciplines is involved in solving 

the problem.

Acknowledgments

Technical reviews by Art Bookstrom, Niki Wintzer, and 

Ennis Geraghty (Stillwater Mining Co.) greatly improved 

this chapter. Constructive comments were provided by 

Klaus Schulz and Dwight Bradley on an early version of the 

manuscript; John DeYoung and Suzanne Nicholson carefully 

looked through the document after the technical review was 

complete. Keith Labay helped create the world map, and 

Heather Parks constructed the first draft of the illustrations.

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Tables N2, N5, N6, and N8–N10
N72 Critical Mineral Resources of the United States—Platinum-Group Elements Table N2. Examples of rocks and ores with anomalous platinum-group-element concentrations that are not associated with magmatic deposits, by deposit type.—Continued [Units of measure: g/t, gram per metric ton; Mt, million metric tons; ppb, part per billion; ppm, part per million. Elements: Au, gold; Co, cobalt; Cu, copper; Fe, iron; Mo, molybdenum; Ni, nickel; Pd, palladium; PGE, platinum-group element; Pt, platinum; U, uranium; V, vanadium; Zn, zinc. n.d., no data; %, percent] Deposit type Site name and (or) location Comments Number of samples Platinum (ppm) Palladium (ppm) Gold (ppm) Other metals Reference(s) Black shale Jiangxi, China None 1 0.01 0.02 0.16 0.055% Mo; 0.01% Ni; 1.12% V Wilde (2005) Woodford Shale, Okla- homa Maximum value n.d. 0.048 n.d. 0.018 n.d. Orth and others (1988) Upper Woodford Shale, Indiana Range of values 4 0.026 to 0.063 0.008 to 0.016 n.d. n.d. Lechler and Hsu (1989) Brecciated quartz veins in fault zone Waterberg deposit, South Africa Sample cuts across ore zone; range of values

4 184 to 

1,747

n.d. n.d. n.d. Wagner (1929); 

Armitage and others (2007)

Picked samples; range of values 2 5,565 to 

5,735

n.d. n.d. n.d. Wagner (1929); 

Armitage and others (2007)

Copper skarn Carr Fork, Utah Upper Parnell Cu-Au skarn 3 n.d. 0.068 to 

0.210

0.450 to 

0.960

n.d. Cameron and

Garmoe (1987)

Au-pyrite-quartz ore shoot 6 n.d. 0.069 to 

0.340

0.100 to 

0.270

n.d. Cameron and

Garmoe (1987)

Gold–PGE mineralization

associated with reducedfacies sediment-hosted strata-bound coppermineralization

Ruwe Hill Mine, Katangan copperbelt, 

(Mutoshi Mine), Democratic Republic of the Congo

Samples from 60-meter gallery 56 10.29 3.72 4.68 n.d. Jebwab (2003) citing Buttgenbach 

(1908); Ball and Shaler (1914)

None n.d. 12 n.d. 9 n.d. Stevens (1905)

High-grade hydrothermal

manganese ore

Gibellini Mine, Nevada Range of values 6 0.167 to 

0.626

0.018 to 

0.045

n.d. n.d. Lechler and others 

(1988)

Homestake-type gold 

deposits

Dome Mine, Ontario, 

Cananda

None 16 n.d. 0.020 to 

0.262

0.12 to 

9.5

n.d. Fryer and others 

(1979)
Table N2 N7

Table N2. Examples of rocks and ores with anomalous platinum-group-element concentrations that are not associated with magmatic deposits, by deposit type.—Continued[Units of measure: g/t, gram per metric ton; Mt, million metric tons; ppb, part per billion; ppm, part per million. Elements: Au, gold; Co, cobalt; Cu, copper; Fe, iron; Mo, molybdenum; Ni, nickel; Pd, palladium; PGE, platinum-group element; Pt, platinum; U, uranium; V, vanadium; Zn, zinc. n.d., no data; %, percent]

Site name and (or) Number of Platinum Palladium Gold Other Deposit type Comments Reference(s) location samples (ppm) (ppm) (ppm) metals

Hydrogeneous iron- Central Pacific Range of values; mean value 34 0.14 to n.d. n.d. n.d. Halbach and others 

manganese crusts in parentheses 1.02 (1989)

on seamounts (0.51)

Federated States of Mean value 32 0.239 0.002 n.d. n.d. Hein and others 

Micronesia-Palau (2000)

Marshall Islands Mean value 77 0.634 0.003 n.d. n.d. Hein and others 

(2000)

Northwest of Marshall Mean value 19 0.501 <0.0054 n.d. n.d. Hein and others 

Islands (2000)

Johnston Island Mean value 99 0.244 n.d. n.d. n.d. Hein and others 

(2000)

California borderlands Mean value 60 0.082 n.d. n.d. n.d. Hein and others 

(2000)

Northwest Pacific Mean value 113 0.777 n.d. n.d. n.d. Hein and others 

(2000)

Hawaii Mean value 25 0.174 0.003 n.d. n.d. Hein and others 

(2000)

West Pacific Range of values 20 0.04282 to 0.00051 to 0.0006 to n.d. Xue and others 

0.50062 0.00199 >1.5 (2005)

Afanasiy-Nikitin Range of values 9 0.101 to 0.004 to 0.02 0.0025 to n.d. Banakar and others 

seamount 0.78 0.19 (2007)

Hydrothermal copper New Rambler, Wyoming Quartz-pyrite-chalcopyrite veins Average of 4 75 n.d. n.d. McCallum and 

vein associated with shears in mafic 87 samples others (1976)

complex; Approximately 

19.3 kilograms of PGEs produced

Limestone Welden Limestone, Maximum value n.d. 0.05 n.d. 0.014 n.d. Orth and others 

Oklahoma (1988)

Near-surface bonanza Serra Pelada, Brazil 43-meter interval in drill core n.d. 204 1,174 4,709 n.d. Cabral and others 

ore from sulfur- (2002)

deficient hydro -thermal system

3
N74 Critical Mineral Resources of the United States—Platinum-Group Elements Table N2. Examples of rocks and ores with anomalous platinum-group-element concentrations that are not associated with magmatic deposits, by deposit type.—Continued [Units of measure: g/t, gram per metric ton; Mt, million metric tons; ppb, part per billion; ppm, part per million. Elements: Au, gold; Co, cobalt; Cu, copper; Fe, iron; Mo, molybdenum; Ni, nickel; Pd, palladium; PGE, platinum-group element; Pt, platinum; U, uranium; V, vanadium; Zn, zinc. n.d., no data; %, percent] Deposit type Site name and (or) location Comments Number of samples Platinum (ppm) Palladium (ppm) Gold (ppm) Other metals Reference(s) Porphyry copper, copper-gold type St. Tomas II, Philippines Mean value 5 0.038 0.16 1.8 n.d. Tarkian and Koopmann (1995) Afton, British Columbia, Canada Range of values 10 0.0026 to 0.143 0.0193 to 1.179 1.226 to 7.764 n.d. Nixon (2003) Skouries, Greece Deposit has 206 Mt at 0.5% Cu, 75 ppb Pd, and 17 ppb Pt n.d. 0.017 0.075 n.d. n.d. Kiousis and others (2005) Porphyry copper, copper-molybdenum type Aksug, Sora, and Zhireken, Russia; Erdenetuin- Obo, Mongolia Range of values 31 <0.010 to 0.028 0.013 to 0.018 0.017 to 0.061 n.d. Sotnikov and others (2001) Reduced-facies sediment-hosted strata-bound copper Kalahari copperbelt, Namibia Range of values 3 0.034 to 0.12 0.0 to 0.008 0.007 to 0.046 2.5 to 3.3% Borg and others (1987) Secondary oxidized system developed on reduced-facies 

sediment-hosted strata-bound copper mineralization

Lubin area, Poland Red Kupferschiefer deposit; 

range of values for palladium and platinum; average for gold

6 0.163 to 

0.75

0.250 to 

1.759

15.419 

(average)

n.d. Pieczonka and 

others (2008)

Sediment-hosted goldarsenic deposits Sukhoi Log, Siberia, Russia Altered carbonaceous and pyritic meta-siltstone of the Khomolkho Formation 22 0.01 0.002 18.056 n.d. Wilde (2005)
Table N2 N75

Table N2. Examples of rocks and ores with anomalous platinum-group-element concentrations that are not associated with magmatic deposits, by deposit type.—Continued[Units of measure: g/t, gram per metric ton; Mt, million metric tons; ppb, part per billion; ppm, part per million. Elements: Au, gold; Co, cobalt; Cu, copper; Fe, iron; Mo, molybdenum; Ni, nickel; Pd, palladium; PGE, platinum-group element; Pt, platinum; U, uranium; V, vanadium; Zn, zinc. n.d., no data; %, percent]

Deposit type Site name and (or) location Comments Number of samples Platinum (ppm) Palladium (ppm) Gold (ppm) Other metals Reference(s)

Sediment-hosted nickelmolybdenum Kuusilampi and Kol- misoppi orebodies, 

Talvivaara deposit, Finland

15 to 25% sulfide in intensively 

folded black schists; samples 

with Ni ≥ 0.1%; deposit has 2.052 billion metric tons containing 0.22% Ni, 0.02% Co, 0.13% Cu, 0.05% Zn, and 0.0017% U; range of values. Mean value in parentheses

27 0.002 to 

0.027 

(0.008)

0.006 to 

0.073 (0.12)

<0.005 

to 0.0159 

(0.0023)

0.13% Cu; 

0.22% Ni; 

0.05% Zn

Loukola-Ruskeeniemi (1991); 

Loukola-Ruskeeniemi and Heino (1996); Talvivaara Mining Company plc (2013)

Huangjiawan Mine; 

Lower Cambrian Niutitang Formation, southern China

Polymetallic sulfide layers; 

mean value

7 0.295 0.3 0.334 3.0% Mo; 

2.6% Ni

Mao and others 

(2002)

Lower Cambrian 

Niutitang Formation, southern China

Polymetallic sulfide layers n.d. 0.405 0.38 0.32 5.7% Mo; 

3.6% Ni

Mao and others 

(2002)

Guizhou, southern China Polymetallic sulfide layers 8 0.19 0.24 0.17 3.23% Mo; 

3.55% Ni

Wilde (2005)

Nick, Yukon Territory, 

Canada

Representative mineralized 

samples

9 0.41 0.214 0.086 0.24% Mo; 

5.24% Ni

Hulbert and others 

(1992)

Western Hunan, China Polymetallic sulfide layers 1 0.3 0.33 0.249 18.28% Mo; 

5.15% Ni

Wilde (2005)

Unconformity uranium Jabiluka uranium–gold–

palladium ore, Northern Territory, Australia

Uranium ore, graphitic schist 27 n.d. 0.29 2.63 n.d. Wilde (2005)

Coronation Hill, Northern 

Territory, Australia

Deposit has 4.85 Mt containg 

4.85 g/t Au, 0.19 g/t Pt, and 0.65 g/t Pd

n.d. 0.19 0.65 4.85 n.d. Carville and others 

(1990)

Uranium–gold–PGE 

mineralization in carbonate-hematitepitchblende shear and fracture zone

Nicholson No. 2 zone, 

Saskatchewan, Canada

None n.d. 13.8 17.3 136 n.d. Hulbert and others 

(1988)
N76 Critical Mineral Resources of the United States—Platinum-Group Elements Table N5. Igneous intrusions and intrusive complexes that contain more than 97 percent of the world’s identified platinum-group-element (PGE) and gold resources, in order of total contained PGEs. [Data are from table N10 at the end of this chapter. Ma, mega-annum (106 years ago); PGE, platinum-group element; Ti-V, titaniferous-vanadiferous; n.d., no data] Igneous intrusion and country Age (Ma) Age reference Deposit types Names of deposits included in tally Resources (million metric tons) PGE and gold grade (grams per metric ton) Total contained PGEs (kilograms) Proportion of total resources, (percent) Active PGE production Bushveld Complex, South Africa 2,054.4

±1.3 Scoates and Friedman (2008) Contact and reef Merensky Reef, UG2 Chromitite, Platreef, and 

Sheba’s Ridge deposits

21,700 3.46 75,200,000 73 Yes, for PGE

Noril’sk-Talnakh 

area, Russia

290

±2.8 to 

226.7

±0.9

Malitch and others 

(2012)

Conduit Talnakh, Noril’sk 1, Maslovskoe,

and Chernogorskoye deposits

2,080 4.84 10,100,000 10 Yes, as a 

byproduct

Great Dyke, 

Zimbabwe

2,575.4

±0.7 Oberthür and 

others (2002)

Reef (sulfide) Main Sulphide Zone, 

all properties

2,260 3.63 8,190,000 7.9 Yes, for PGE

Stillwater Complex, 

Montana

2,704

±4 Premo and others 

(1990)

Reef (sulfide) J–M Reef, Stillwater, and 

East Boulder Mines

149 15.0 2,240,000 2.2 Yes, for PGE

Duluth Complex, 

Minnesota

1,095.94

±0.18 Hoaglund (2010) Contact Northmet, Birch Lake, and Maturi

(deposits with PGEs reported)

2,490 0.560 1,390,000 1.3 No

Onega intrusion, 

Russia

2,050±70 Geological Survey 

of Finland (2010)

Reef (Ti-V 

layers)

Pudozhgorskoe deposit 517 1.62 837,000 0.81 No

Fedorov-Pana intrusion, Russia 2,5326

±6; 

2,496±7

Nitkina (2006) Contact and 

reef

All deposits 458 1.70 780,000 0.75 No

Burakovskaya intrusion, Russia 2,433

±4;

2,430

±5

Bailly and others 

(2009)

Reef All deposits 587 1.10 645,000 0.62 No

Portimo Igneous 

Complex, Finland

n.d. n.d. Contact and 

reef

All deposits 382 1.37 524,000 0.51 No

Wellgreen intrusion, Canada 232.3±1 Mortensen and Hulbert (1992) Conduit n.d. 461 0.91 421,000 0.41 No

Sudbury Complex, 

Canada

1,850±1 Corfu and Lightfoot (1996) Sudbury Vale properties (PGEs reported) 105 2.36 249,000 0.24 Yes, as a byproduct

Skaergaard intrusion, Greenland 55.59

±0.13 Hamilton and 

Brooks (2004)

Reef (sulfide) Pd5 layer 104 2.18 227,000 0.22 No

Stella intrusion, 

South Africa

3,033.5

±0.3 Maier and others 

(2003)

Reef (Ti-V 

layers)

Kalplats project 137 1.53 210,000 0.20 No

Lac des Iles, 

Canada

2,689 ±1 Baker and others 

(2003)

Atypical; pipelike orebodies? n.d. 56 3.68 207,000 0.20 Yes, for PGE

Proportion 

of total

(percent)

Total contained

PGEs

(kilograms)

  Subtotal for first four intrusions in this list 92.3 95,700,000

Subtotal for all intrusions in this list 97.6 101,000,000

  Total inventory identified in table N10 of this chapter 100 104,000,000
Table N6 N77

Table N6. Areal extent and stratigraphic thickness of layered intrusions with reef-type platinum-group-element deposits and some examples of large intrusions with no known deposits, in order of areal extent.—Continued[Areal extent in parentheses is inferred from geophysical interpretations; —, no identified deposits or occurrences]

Intrusion Country or U.S. State Areal extent (square kilometers) Stratigraphic thickness of cumulates (meters) Name of reef deposits and occurrences Comments Reference(s)

Bushveld Complex 

(Rustenburg Layered Suite)

South Africa 12,000 

(69,000)

8,000 Merensky Reef, 

UG2 Chromitite, Pseudo Reef

None Vermaak and von Gruenewaldt (1986); Walraven 

(1986); Martini and others (2001); Barnes and Maier (2002); Cawthorn and others (2002)

Tete Complex Mozambique/

Malawi

5,500 Unknown — None Instituto Nacional de Geologica (1987); Maier 

and others (2001)

Dufek intrusion Antarctica 4,300

(7,700)

3,500 exposed, 

up to 9,000 

inferred

— Largely covered by ice; 

only 145 square kilometers

of rock is exposed

Ford (1976); Ford, Schmidt, and Boyd (1978); 

Ford, Schmidt, Boyd, and Nelson (1978); Ferris and others (1998; 2003)

Molopo Farms 

Complex

South Africa 3,900

(14,000)

3,200 — None von Gruenewaldt and others (1989); Walker and 

others (2010); Prendergast (2012)

Great Dyke Zimbabwe 3,100 3,350 Main Sulphide Zone, 

Boehmke Reef

None Worst (1957); Prendergast and Wilson (1989); 

Wilson and Tredoux (1990); Evans and others (1996); Wilson and Prendergast (1989, 2001); Armstrong and Wilson (2000); Oberthuer (2002)

Windimurra igneous 

complex

Australia 2,500 ~10,000 — None Parks and Hill (1986); Mathison and Ahmat 

(1996); Ivanic and others (2013)

Barro Alto Complex Brazil 2,000 10,100 — Granulite facies metamorphism Schobbenhaus and others (2004); Pimentel and others (2006); Ferreira-Filho and others (2010)

Fraser Range 

Metamorphics (Fraser Complex)

Australia 1,000 >16,000 — Metamorphosed and 

deformed

Meyers (1985); Spaggiari and others (2009)

Niquelândia layered 

complex

Brazil 770 17,750 — Intense deformation 

and amphibolite to granulate facies metamorphism

Schobbenhaus and others (2004); Pimentel 

and others (2006); Ferreira-Filho and others (2010)

Fiskenaesset 

intrusion

Greenland 730 550 Ghisler Reef Deformed with amphibolite 

to granulate facies metamorphism

Myers (1976, 1985); Allaart (1998); Appel and 

others (2011)

Rincon del Tigre 

Complex

Bolivia 640 4,600 Precious Metals Zone None Mitchell (1983); Montemurro (1984); 

Prendergast (2000)

Aganozero-Burakovsky intrusion Russia 640 6,200 Aganozero, Platinometal horizon “A,” 

Shalozerskoe (Kuk Ruchi Zone)

None Grokhovskaya and others (2005); Bailly and 

others (2009)
N78 Critical Mineral Resources of the United States—Platinum-Group Elements Table N6. Areal extent and stratigraphic thickness of layered intrusions with reef-type platinum-group-element deposits and some examples of large intrusions with no known deposits, in order of areal extent.—Continued [Areal extent in parentheses is inferred from geophysical interpretations; —, no identified deposits or occurrences] Intrusion Country or U.S. State Areal extent (square kilometers) Stratigraphic thickness of cumulates (meters) Name of reef deposits and occurrences Comments Reference(s) Pedra Branca Unit– Brazil 530 unknown Pedra Branca Intense deformation Angeli (2005); Barrueto and Hunt (2010) Ceará and Tróia and amphibolite to Complexes granulate facies metamorphism Coldwell Complex Canada 410 <2,500 Skipper Lake None Barrie and others (2002) Fedorov Pana Fennoscandia 340 3,750 Vostochno (eastern) Pans- None Alapieti and Lahtinen (2002); Mitrofanov intrusion koe; Western Panskoe and others (2005) Stillwater Complex Montana 190 6,500 J–M Reef, A–B None McCallum and others (1980); Segerstrom and (2,300) chromitite Carlson (1982); Raedeke and McCallum (1984); Kleinkopf (1985); Zientek and others (2002) Weld Range Com- Australia 77 8,500 Parks Reef None Parks (1998); Geological Survey of Western plex Australia (2004) Skaergaard intrusion Greenland 63 3,500 Platinova Reef None McBirney (1989); Nielsen and others (2005) Munni Munni Australia 63 5,450 Ferguson Reef None Barnes and others (1990); Geological Survey intrusion of Western Australia (2004) Penikat intrusion Finland 59 2,200 Sompujärvi, Ala-Pennika, None Alapieti and Lahtinen (2002); Rasilainen and and Paasivaara Reefs others (2010)

Sonju Lake intrusion Minnesota 44 1,200 Sonju Lake Reef None Miller (1999); Miller and others (2002)

Lake Owen intrusion Wisconsin 39 4,500 Unnamed occurrence None Houston and Orback (1976); Loucks and 

Glasscock (1989); Shive and others (1990); Loucks (1991); Green and Drouillard (1994)

Portimo Complex Finland 39 900 Siika-Kämä Reef, None Alapieti and Lahtinen (2002); Rasilainen and 

Rytikangas Reef others (2010)

Stella intrusion South Africa 16 1,000 Main Reef None Maier and others (2003); Carroll (2005); 

Lewins and others (2008); African Rainbow Minerals Ltd. (2010)

Panton intrusion Australia 13 1,600 Top and Middle Reefs None Perring and Vogt (1991); Hoatson and Blake (2000)
Table N8 N79

Table N8. Trace element geochemistry of waters from selected reef-type, contact-type, and conduit-type deposits.—Continued[Elements: Cd, cadmium; Co, cobalt; Cr, chromium; Cu, copper, Fe, iron; Mn, manganese; Ni, nickel; Pb, lead; Sb, antimony; Zn, zinc. n.d., no data]

Area or

subject

Media or

material pH

Cd Co Cr Cu Fe Mn Ni Pb Sb Zn

Reference(s) (part per billion)

Reef-type deposits

Pilanesberg Platinum 

Mines, Zimbabwe

Surface water 

downstream of waste dumps

Near 

neutral to 

slightly 

alkaline

60 260 2,370 60 n.d. n.d. 820 110 910 70 Meck and others 

(2006)

UG2 Chromitite, 

Bushveld Complex, South Africa

Leachate from 

leach test

5.1 <5 <25 370 <25 1.53 74 50 <20 n.d. 50 Bolton (2013)

Merensky Reef, 

Bushveld Complex, South Africa

Leachate from 

leach test

5.1 <5 <25 <25 <25 1.45 1,220 430 <20 n.d. 30 Bolton (2013)

Stillwater Mine, 

Montana, J–M Reef ore

Leachate from 

leach test

n.d. <0.0001 

to <0.1

n.d. <0.01

to <0.5

<0.01

to <0.5

n.d. n.d. <0.1

to <0.5

0.0004 

to <0.5

n.d. 0.01

to <0.5

Kirk and others 

(2006); Montana Department of 

Environmental Quality (2012b)

East Boulder Mine, 

Montana, J–M Reef ore

Leachate from 

leach test

n.d. <0.0001 n.d. <0.01 <0.01 n.d. n.d. <0.01 <0.0001 n.d. <0.01 Kirk and others 

(2006); Montana Department of 

Environmental Quality (2012b)

Stillwater Mine, 

Montana, waste rock

Leachate from 

leach test

n.d. <0.0001 n.d. <0.01 <0.01 n.d. n.d. <0.01 0.0007 n.d. 0.01 Kirk and others 

(2006); Montana Department of Environmental Quality (2012b)

East Boulder Mine, 

Montana, waste rock

Leachate from 

leach test

n.d. <0.0001 n.d. <0.01 <0.01 n.d. n.d. <0.01 0.0002 n.d. 0.02 Kirk and others 

(2006); Montana Department of Environmental Quality (2012b)

Stillwater Mine, 

Montana, tailings

Leachate from 

leach test

n.d. <0.0001 n.d. <0.01 <0.01 n.d. n.d. <0.01 0.0002 n.d. 0.01 Kirk and others 

(2006); Montana Department of Environmental Quality (2012b)
N80 Critical Mineral Resources of the United States—Platinum-Group Elements Table N8. Trace element geochemistry of waters from selected reef-type, contact-type, and conduit-type deposits.—Continued [Elements: Cd, cadmium; Co, cobalt; Cr, chromium; Cu, copper, Fe, iron; Mn, manganese; Ni, nickel; Pb, lead; Sb, antimony; Zn, zinc. n.d., no data] Area or subject Media or material pH Cd Co Cr Cu Fe Mn Ni Pb Sb Zn Reference(s) (part per billion) Reef-type deposits—Continued East Boulder Mine, Leachate from n.d. <0.0001 n.d. <0.01 <0.01 n.d. n.d. <0.01 <0.0001 n.d. 0.01 Kirk and others Montana, tailings leach test (2006); Montana Department of Environmental Quality (2012b) East Boulder Mine, Operational n.d. <0.0001 n.d. <0.001 0.002 0.025 0.005 <0.02 <0.002 n.d. <0.01 Kirk and others Montana, tailings wastewater (2006); Montana water Department of Environmental Quality (2012b) Stillwater Mine, Operational n.d. <0.0001 n.d. <0.01 0.003 0.025 <0.003 n.d. <0.003 <0.01 n.d. Kirk and others Montana, wastewater (2006); Montana tailings water Department of Environmental Quality (2012b) Contact-type deposits Northmet deposit, Surface water 7.0 <0.2 <1 2 n.d. 0.89 80 <5 <1 n.d. <10 Minnesota Department Minnesota downstream of Natural Resources, of the de- and the U.S. Army posit Corps of Engineers (2009)

Conduit-type deposits

Eagle Mine, Groundwater 8.7 <0.5 <10 <5 <5 0.19 220 59 <1 <6 88 Kennecott Eagle 

Michigan near the the Minerals Co. 

deposit (2006)

Water-quality criteria

Groundwater Standard n.d. 0.005 n.d. 0.100 1.3 0.300 0.050 0.1 0.015 n.d. 2.0 Montana Department 

criteria (State of Environmental of Montana) Quality (2012a)

Chronic surface- Standard n.d. n.d. n.d. n.d. 1.3 n.d. 0.50 0.610 n.d. n.d. 7.4 Suter (1996);

water criteria U.S. Environmental (U.S. Government) Protection Agency 

(2009)

Drinking water crite- Standard n.d. 0.003 n.d. 0.05 0.002 n.d. n.d. 0.07 0.01 0.02 n.d. World Health Organiria (World Health zation (2008)

Organization)
Table N9 N81

Table N9. Acid-base accounting for selected reef-type, contact-type, and conduit-type deposits.[Net neutralizing potential is the difference between the acid-generating and the acid-neutralizing potentials. A negative value indicates excess acid-generating potential and a positive value indicates excess acid-neutralizing potential. kg CaCO3/t, kilograms of calcium carbonate per metric ton; ?, identification uncertain; n.d., no data] 

Deposit 

and area Material

Sulfur content

(weight

percent) 

Acid-generating

potential 

Acid-neutralizing 

potential

Net neutralizing 

potential Reference(s)

(kg CaCO3 /t)

Reef-type deposits

J–M Reef, Stillwater

Complex, Montana

Waste rock <0.05 Low (<1) 20 to 70 n.d. Kirk and others (2006)

Merensky Reef, Pilanesberg 

Platinum Mine, Bushveld Complex, South Africa

Ore <0.07 2.07 15.12 7.29 Bolton (2013)

UG2 Chromitite, Pilanesberg 

Platinum Mine, Bushveld Complex, South Africa

Ore? <0.01 0.31 10.21 32.66 Bolton (2013)

Contact-type deposits

Babbit deposit, Duluth 

Complex, Minnesota

Waste rock n.d. 19.7 to 44.1 0.6 to 2.1 Slight excess of 

acid-generating 

potential

Lapakko (1994)

Duluth Complex, 

Minnesota

Tailings from metallurgical testing of Duluth 

Complex ores

n.d. 6.2 to 8.1 14 Slight excess of 

acid–neutralizing 

potential

Lapakko and Berndt (2009); Minnesota 

Department of Natural Resources, and the U.S. Army Corps of Engineers (2009)

Northmet deposit, Duluth

Complex, Minnesota 

Nonreactive waste rockReactive waste rock

n.d.n.d.

0.9

14.1

n.d.n.d.

n.d.n.d.

Minnesota Department of Natural Resources, 

and the U.S. Army Corps of Engineers (2009)

Turfspruit deposit, Platreef,

Bushveld Complex, South Africa

Ore 0.3 to 0.5 n.d. n.d. –5.56 to 2.8 Lishman (2009)

Waste rock 0.27 to 0.67 n.d. n.d. –11.21 to 0.78

Tailings 0.03 to 0.05 n.d. n.d. 12.6 to 13.44

Conduit-type deposits

Eagle deposit, Michigan Massive ore

Semimassive ore

Peridotite

Pyroxenite

Siltstone

Sandstone

Hornfels

35.6613.43

1.571.301.030.338.91

1,111

419

4940321049

0

414434193818

–1,111

–378

–5–6

–12

28

–31

Kennecott Eagle Minerals Co. (2006)
N82 Critical Mineral Resources of the United States—Platinum-Group Elements Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued [Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain] Deposit name Deposit type Intrusion Country Latitude Longitude Ore (Mt) Ni (percent) Cu (percent) Pt (g/t) Pd (g/t) Au (g/t) PGE and Au (g/t) Reference(s) Decimal degrees WGS 84 A–B chromitite, West Fork area Reef Stillwater Complex United States 45.38 –110.01 3.6 n.d. n.d. 0.685 1.71 n.d. 2.40 Zientek (1993) Aganozero Reef (chro- mitite) Burakovskaya intrusion Russia 62.20 36.52 200 n.d. n.d. n.d. n.d. 8.9 1.20 Geological Survey of Finland (2011) Aguablanca Synorogenic Aguablanca intrusion Spain 37.85 –6.18 17 0.51 0.43 0.241 0.217 0.121 0.579 Alvarez and others (2009) Ahmavaara Contact Portimo Finland 66.10 26.04 190 0.09 0.17 0.17 0.82 0.1 0.990 Geological Survey of Finland (2011) Akanani (P1 and P2) Contact Bushveld South Africa –23.97 28.85 290 0.22 0.12 1.57 1.91 0.219 3.70 Lonmin plc (2011) Akanvaara LC Reef (chro- mitite) Akanvaara Finland 67.20 28.20 27 n.d. n.d. 0.52 0.08 n.d. 0.600 Geological Survey of Finland (2011) Akanvaara UC Reef (chro- mitite) Akanvaara Finland 67.18 28.20 18 n.d. n.d. n.d. 0.92 n.d. 0.920 Geological Survey of Finland (2011) Ala-Penikkavaara (AP Reef) Reef Penikat intrusion Finland 65.82 25.01 3.5 n.d. 0.21 1.68 6.16 0.28 7.95 Geological Survey of Finland (2011) Benbow Contact Stillwater United States 45.36 –109.81 130 0.22 0.22 0.0068 0.0106 n.d. 0.0174 Zientek (1993) Birch Lake (Twin 

Metals)

Contact Duluth United States 47.73 –91.80 340 0.15 0.48 0.196 0.411 0.0949 0.702 Parker and Eggleston (2012)

Bird River (Page 

Zone and Ore Fault Zone)

Contact Bird River 

Sill

Canada 50.48 –95.51 5.2 0.34 0.18 0.088 0.352 0.0151 0.456 Ewert and others 

(2008)

Boikgantsho 

Project

Contact Bushveld South Africa –23.90 28.88 130 0.11 0.08 0.767 0.743 0.0724 1.58 Ferreira and Bisnath 

(2012)

Bokai and Unki 

East (Main Sulphide Zone)

Reef 

(sulfide)

Great Dyke Zimbabwe –19.74 30.04 210 0.21 0.17 1.88 1.48 0.332 3.85 Central African Mining

& Exploration Company Plc (2009);Anglo American Platinum Ltd. (2011)

Brattbakken Hydrothermal

or remobilized?

(stockwork)

n.d. Norway 63.94 12.34 1.5 1 0.8 0.7 n.d. n.d. 0.700 Geological Survey 

of Finland (2011)
Table N10 N83

Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued[Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain]

Deposit Latitude Longitude Ore Ni Cu Pt Pd Au PGE and Au Deposit type Intrusion Country Reference(s) name Decimal degrees WGS 84 (Mt) (percent) (percent) (g/t) (g/t) (g/t) (g/t)

Broken Hammer Offset type Sudbury Canada 46.77 –81.04 0.23 0.1 0.92 2.01 1.9 0.71 4.62 Doran and others 

(Sudbury area) deposit (2012)

Bystrinskoe Conduit Pechenga Russia 69.40 30.73 36 0.49 0.26 n.d. n.d. 0.008 0.0860 Geological Survey 

greenstone of Finland (2011)

belt 

Camp deposit Contact Stillwater United States 45.41 –110.08 5.9 0.42 0.23 0.018 0.139 n.d. 0.157 Zientek (1993)

Chernogorskoye Conduit Noril’sk Russia 69.15 88.27 140 0.22 0.29 n.d. n.d. n.d. 3.90 Russian Platinum 

deposit Talnakh (2011b)

Chinei intrusion. Contact Chinei Russia 56.48 118.58 23 0.05 0.65 0.23 1.04 0.17 1.44 Vladimir Chechetkin,

all deposits Research Geologist(reserves only) (retired), written 

comm., 2009

Dikolati Nickel Contact n.d. Botswana –22.03 27.74 4.1 0.7 0.5 n.d. n.d. n.d. 1.20 Discovery Metals 

project Ltd. (2013)

Eagle Nickel Conduit Eagle Intru- United States 46.75 –87.89 4.1 3.6 2.9 0.73 0.47 0.28 0.617 Rossell and 

sion Coombes (2005)

Eagle’s Nest Conduit Double Eagle Canada 52.74 –86.31 22 1.4 0.98 1 3.22 0.23 4.44 Burgess and others 

Complex (2012)

Expo Komatiitic n.d. Canada 61.56 –73.45 5.1 0.85 0.86 0.372 1.6 0.1 2.07 Puritch and others 

(2006)

Fedorovo Contact Imandra- Russia 67.51 35.00 410 0.09 0.15 0.311 0.933 0.084 1.33 Barrick Gold Corp. 

Varzuga (2008); Schissel 

greenstone and others (2002); belt Mitrofanov and 

others (2007); Rasilainen and others (2010)

Geordie Lake Contact Coldwell Canada 48.82 –86.49 40 n.d. 0.37 0.038 0.611 0.048 0.697 Drennan and Fell 

Complex (2010)

Gold-platinum Reef Burakovskaya Russia 62.07 36.45 23 n.d. n.d. n.d. n.d. 2.49 0.600 Geological Survey 

horizon “B” intrusion of Finland (2011)

Burakovka intrusion
N84 Critical Mineral Resources of the United States—Platinum-Group Elements Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued [Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain] Deposit name Deposit type Intrusion Country Latitude Longitude Ore (Mt) Ni (percent) Cu (percent) Pt (g/t) Pd (g/t) Au (g/t) PGE and Au (g/t) Reference(s) Decimal degrees WGS 84 Gold zone Reef (sul- Skaergaard Greenland 68.19 –31.67 110 n.d. n.d. 0.05 0.59 1.68 2.32 Secher and others fide) (2007) Great Lakes Contact Crystal Lake Canada 48.08 –89.63 46 0.18 0.33 0.206 0.686 0.0686 0.960 Postle and others Nickel gabbro (1986); Lightfoot and Lavigne (1995) Haukiaho Contact Koilismaa Finland 65.85 28.06 27 0.24 0.36 0.209 0.549 0.216 0.974 Iljina and others (2005) Hietaharju Komatiitic Kuhmo Finland 65.15 29.08 1.1 0.8 0.4 0.49 1.17 n.d. 1.66 Geological Survey of Finland (2011) Hitura Synorogenic Hitura Finland 63.84 25.05 4.4 0.52 0.18 0.03 0.04 n.d. 0.0700 Geological Survey of Finland (2011) Horizon of banded Reef Burakovskaya Russia 62.07 36.45 23 n.d. n.d. n.d. n.d. 1.3 1.10 Geological Survey subzone of intrusion of Finland (2011) Agano zero and Shalozero blocks Ikenskoe Conduit Kun-Manie Russia 55.33 132.53 36 0.45 0.13 0.2 0.2 n.d. 0.400 Armitage (2007) Massif Ivakkak Komatiitic n.d. Canada 61.44 –77.17 0.63 1.6 2 0.782 3.42 0.168 4.37 Puritch and others (2006) Ivanhoe (Turfspruit, Contact Bushveld South Africa –24.10 28.98 630 0.33 0.17 1.61 1.64 0.267 3.60 Ivanplats (Pty) Ltd. Rietfontein, (2013)

Macalacaskop)—selective underground

Jinbasoshan Conduit Jinbaoshan China 25.01 100.76 13 0.18 0.19 1.09 0.669 n.d. 1.76 Jinshan Gold Mines 

intrusion (2004)

Jinchuan, Conduit Jinchuan China 38.47 102.17 440 1 0.66 0.11 0.082 n.d. 0.197 Yonglin (2005)

Baijiazuizi intrusion

Kabanga Conduit Kabanga Tanzania –2.87 30.58 46 2.7 0.38 0.252 0.256 0.0873 0.595 Xstrata plc (2007)Kalplats Project Reef (Ti-V Amalia- South Africa –26.31 24.86 140 n.d. n.d. 0.696 0.788 0.0433 1.53 Lewins and others 

(Main Reef+LG+ layers) Kraaipan (2008); African 

UM+UUM+LM+ greenstone Rainbow MinerMR+MMW) belt als Ltd. (2011)

Karenhaugen Hydrothermal n.d. Norway 70.03 25.22 0.7 n.d. 0.57 0.31 0.87 n.d. 1.18 Geological Survey 

or remo- of Finland (2011)bilized?
Table N10 N85

Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued[Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain]

Deposit Latitude Longitude Ore Ni Cu Pt Pd Au PGE and Au Deposit type Intrusion Country Reference(s) name Decimal degrees WGS 84 (Mt) (percent) (percent) (g/t) (g/t) (g/t) (g/t)

Karhunjupukka Reef (Ti-V Pajala-Kolari Finland 67.29 24.48 5.2 n.d. 0.03 n.d. 0.1 n.d. 0.100 Geological Survey 

layers) of Finland (2011)

Karikjavr Conduit Pechenga Russia 69.30 31.43 22 0.63 0.5 n.d. n.d. n.d. 0.464 Geological Survey 

greenstone of Finland (2011)

belt

Kaukua Contact Koilismaa Finland 65.94 28.18 12 0.1 0.15 0.25 0.71 0.08 0.960 Geological Survey 

of Finland (2011)

Kevitsa Atypical; pipe- n.d. Finland 67.70 29.97 270 0.3 0.41 0.2 0.15 0.11 0.350 Geological Survey 

like ore- of Finland (2011)

bodies?

Kilvenjärvi Contact Portimo Finland 66.24 26.17 0.7 n.d. 2.7 1.12 7.27 0.8 8.39 Geological Survey 

of Finland (2011)

Kilvenjoki Offset type Portimo Finland 66.26 26.17 0.18 0.28 6.1 0.06 2.5 0.84 2.56 Geological Survey 

contact of Finland (2011)

deposit

Kilvenlatvalampi Contact Portimo Finland 66.25 26.27 3.2 n.d. 0.5 0.5 1.4 n.d. 1.90 Geological Survey 

of Finland (2011)

Kingashskoe and Komatiitic n.d. Russia 54.92 95.36 480 0.45 0.21 n.d. n.d. n.d. 0.630 OJSC MMC Norilsk 

Verkhnekin- Nickel (2009)gashskoye

Koitelainen LC Reef (chro- Koitelainen Finland 67.80 27.10 2 n.d. n.d. 0.48 0.9 n.d. 1.38 Geological Survey 

mitite) of Finland (2011)

Koitelainen UC Reef (chro- Koitelainen Finland 67.83 27.32 70 n.d. n.d. n.d. n.d. n.d. 1.10 Geological Survey 

mitite) of Finland (2011)

Koivukivalonaapa Reef? Portimo Finland 66.25 26.20 0.84 n.d. 0.1 2.1 6 0.17 8.10 Geological Survey 

of Finland (2011)

Konttijärvi Contact Portimo Finland 66.12 25.98 75 0.05 0.1 0.27 0.95 0.07 1.22 Geological Survey 

of Finland (2011)

Kootsel’vaara- Conduit Pechenga Russia 69.40 30.33 7.5 1.2 0.64 n.d. n.d. n.d. 0.185 Geological Survey 

Kammikivi greenstone of Finland (2011)

belt

Lac des Iles Atypical; pipe- Lac des Iles Canada 49.17 –89.62 56 0.09 0.06 0.258 3.2 0.222 3.68 North American 

like ore- Palladium Ltd. bodies? (2012)
N86 Critical Mineral Resources of the United States—Platinum-Group Elements Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued [Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain] Deposit name Deposit type Intrusion Country Latitude Longitude Ore (Mt) Ni (percent) Cu (percent) Pt (g/t) Pd (g/t) Au (g/t) PGE and Au (g/t) Reference(s) Decimal degrees WGS 84 Las Aguilas Conduit n.d. Argentina –33.11 –66.13 4.6 0.41 0.41 0.123 0.145 0.311 0.579 Carter and othrers (2012) Lavotta Contact Koilismaa Finland 65.80 28.14 3 0.21 0.26 0.18 0.26 0.2 0.640 Mattila and others (1976); Lahtinen (1983) Lomalampi Komatiitic n.d. Finland 67.91 26.92 3.1 0.17 0.06 0.269 0.122 0.074 0.391 Geological Survey of Finland (2011) Lower Pseudo Reef (sul- Bushveld South Africa –25.18 26.97 14 0.14 0.03 1.41 0.733 0.127 2.37 Harper and de Wit Reef fide) (2009); Platinum Australia Ltd. (2012) Lukkulaisvaara— Contacts of Olanga Russia 66.32 30.87 2.8 0.06 0.12 0.6 2.13 0.14 2.87 Zarubezhtsvetmet total resources macrodike Austrasia pty Ltd. and reef (2009) Main sulfide layer Reef (sul- Munni Munni Australia –21.14 116.84 24 0.09 0.15 1.1 1.5 0.2 2.90 Platina Resources (Ferguson Reef) fide) Ltd. (2010) Marathon Contact Coldwell Canada 48.79 –86.32 120 n.d. 0.24 0.222 0.751 0.072 1.05 Murahwi and others Complex (2011) Maslovskoe Conduit Noril’sk Russia 69.20 88.04 220 0.33 0.51 1.78 4.56 0.19 6.53 OJSC MMC Norilsk deposit Talnakh Nickel (2012)

Maturi (Nokomis) Contact Duluth United States 47.81 –91.73 1,600 0.18 0.56 0.151 0.344 0.0806 0.576 Parker and Egg-

(Twin Metals) leston (2012)

Mequillon Komatiitic n.d. Canada 61.51 –73.75 4.4 0.7 0.9 0.7 2.4 0.2 3.30 Puritch and others 

(2006)

Merensky Reef— Reef (sul- Bushveld South Africa –24.60 30.00 2,500 0.16 0.09 2.51 1.3 0.288 4.23 Zientek and others 

eastern limb fide) (2014)

Merensky Reef— Reef (sul- Bushveld South Africa –25.30 27.30 1,600 0.2 0.08 3.88 1.72 0.269 6.16 Zientek and others 

western limb fide) (2014)

Merensky Reef, Reef Bushveld South Africa –25.14 27.00 160 0.09 0.02 2.19 0.918 0.109 3.47 Platinum Australia 

Pseudo Reef, Ltd. (2011, 2012);

UG2 Platmin Ltd. (2011);

Realm Resources Ltd. (2011)

Mesamax Komatiitic n.d. Canada 61.58 –73.26 1.9 2.1 2.6 0.994 3.83 0.198 5.02 Puritch and others 

(2006)
Table N10 N87

Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued[Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain]

Deposit Latitude Longitude Ore Ni Cu Pt Pd Au PGE and Au Deposit type Intrusion Country Reference(s) name Decimal degrees WGS 84 (Mt) (percent) (percent) (g/t) (g/t) (g/t) (g/t)

Mimosa (Main Reef (sul- Great Dyke Zimbabwe –20.30 29.81 140 0.14 0.11 1.81 1.41 0.318 3.70 Aquarius Platinum 

Sulphide Zone) fide) Ltd. (2011); 

Impala Platinum Holdings Ltd. (2011)

Mogalakwena Contact Bushveld South Africa –24.00 28.92 3,500 n.d. n.d. 0.976 1.15 0.126 2.25 Anglo American 

Mine Platinum Ltd. 

(2011)

Mokopane Contact Bushveld South Africa –24.14 29.00 40 0.15 0.09 0.22 0.33 n.d. 0.550 Blackthorn Resources

Ltd. (2009)

Monchetun- Contact Imandra- Russia 67.88 32.75 48 n.d. n.d. n.d. n.d. n.d. 3.81 Geological Survey 

dorvskoe Varzuga of Finland (2011)

greenstone belt

Mouat Contact Stillwater United States 45.38 –109.90 130 0.31 0.29 0.0039 0.0105 n.d. 0.0144 Zientek (1993)

Mt. General’ Contact Mt. General’ Russia 69.42 31.08 53 0.27 0.46 0.2 2.05 n.d. 2.25 Geological Survey 

skaya skaya of Finland (2011)

Nadezhda Contacts of Olanga Russia 66.32 30.87 1.8 n.d. n.d. n.d. n.d. 3.57 Zarubezhtsvetmet 

macrodike Austrasia pty Ltd. 

(2009)

Nebo-Babel Conduit Giles Australia –26.10 127.71 390 0.3 0.3 n.d. n.d. n.d. 0.180 Western Mining 

Complex Corp. (2005)

Niittylampi Contact Portimo Finland 66.10 26.20 1 0.67 0.49 0.27 0.68 n.d. 0.950 Geological Survey 

of Finland (2011)

Nkomati Mine Conduit Uitkomst South Africa –25.73 30.60 340 0.34 0.13 0.23 0.597 0.0125 0.861 African Rainbow 

(MMZ and intrusion Minerals Ltd. PCMZ) (2006, 2011)

Noril’sk-1 deposit, Conduit Noril’sk Russia 69.23 88.18 81 0.32 0.44 1.68 4.17 0.172 6.19 OJSC MMC Norilsk 

all resource Talnakh Nickel (2011)

categories

Northern Limb Contact Bushveld South Africa –23.50 28.90 120 0.04 0.05 0.478 0.782 n.d. 1.26 McConnachie and 

projects— others (2011)Aurora and Hacra
N88 Critical Mineral Resources of the United States—Platinum-Group Elements Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued [Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain] Deposit name Deposit type Intrusion Country Latitude Longitude Ore (Mt) Ni (percent) Cu (percent) Pt (g/t) Pd (g/t) Au (g/t) PGE and Au (g/t) Reference(s) Decimal degrees WGS 84 Northmet Contact Duluth United States 47.62 –91.97 550 0.09 0.33 0.0823 0.298 0.0418 0.422 Desautels and (Polymet) Zurowski (2012) Nunatak-Brady Contact La Perouse United States 58.55 –136.93 90 0.53 0.33 0.116 0.113 n.d. 0.229 U.S. Bureau of Glacier Mines (1991); Czamanske and others (1981); Himmelberg and Loney (1981) Nutturalampi Contact Portimo Finland 66.26 26.31 0.4 n.d. 0.09 1.5 5.5 0.15 7.00 Geological Survey of Finland (2011) Nye Basin Contact Stillwater United States 45.37 –109.84 280 0.22 0.25 0.0032 0.0087 n.d. 0.0119 Zientek (1993) Owendale (includes Uralian Fifield Australia –32.69 147.46 13 n.d. n.d. 0.7 n.d. n.d. 0.700 Platina Resources Owendale North, Ltd. (2013) Cincinnati, and Milverton) Paasivaara Reef Reef (sul- Penikat Finland 65.81 24.98 5 n.d. 0.28 4.04 2.58 0.61 7.31 Geological Survey (PV Reef) fide) intrusion of Finland (2011) Parkin (Sudbury Offset type Sudbury Canada 46.83 –80.85 0.35 0.59 0.7 0.764 0.874 0.322 1.96 Bailey (2012) area) deposit

Parks Reef Reef (sul- Weld Range Australia –26.84 117.75 15 n.d. n.d. 0.6 0.5 n.d. 1.10 Parks (1998)

fide)

Pazhskoe Hydrothermal Onega Russia 62.71 35.76 2.8 n.d. n.d. 2.5 n.d. n.d. 2.50 Geological Survey 

(vein of Finland (2011)

swarm)

Pedra Branca Reef (chro- Troia Unit of Brazil –5.60 –39.96 6.6 0.23 0.03 n.d. n.d. n.d. 2.27 Anglo American 

mitite) the Cruzeta Platinum Ltd. 

Complex (2009)

Peura-aho Komatiitic Kuhmo Finland 65.20 29.12 0.49 0.6 0.27 0.27 0.58 n.d. 0.850 Geological Survey 

of Finland (2011)

Platinometal hori- Reef (chro- Burakovskaya Russia 62.07 36.45 210 n.d. n.d. n.d. n.d. n.d. 1.40 Geological Survey 

zon “A” mitite) intrusion of Finland (2011)

Pudozhgorskoe Reef (Ti-V Onega Russia 62.28 35.90 520 n.d. 0.13 n.d. n.d. 0.21 1.62 Geological Survey 

layers) of Finland (2011)

River Valley Contact River Valley Canada 46.69 –80.26 130 0.02 0.06 0.197 0.518 0.0372 0.772 McCracken (2012)

intrusion
Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued

Table N10 N89

[Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain]

Deposit Latitude Longitude Ore Ni Cu Pt Pd Au PGE and Au Deposit type Intrusion Country Reference(s) name Decimal degrees WGS 84 (Mt) (percent) (percent) (g/t) (g/t) (g/t) (g/t)

Rocky Claim Contact Stillwater United States 45.35 –109.80 49 0.28 0.26 0.0052 0.0107 n.d. 0.0159 Zientek (1993)

Group (Chrome Lake)

Rooiport Contact Bushveld South Africa –24.27 29.01 18 0.19 0.11 0.474 0.717 0.0896 1.28 Verbeek and Lomberg (2005)

Rusamo Contact Koilismaa Finland 65.78 27.95 1.5 0.24 0.39 0.266 0.384 0.15 0.800 Lahtinen (1983)

Selkirk Contact n.d. Botswana –21.21 27.77 140 0.23 0.27 n.d. n.d. n.d. 0.569 OJSC MMC Norilsk 

Nickel (2011)

Semiletka Conduit Pechenga Russia 69.38 30.33 17 0.65 0.35 n.d. n.d. 0.036 0.140 Geological Survey 

greenstone of Finland (2011)

belt

Shakespeare Conduit Nipissing Canada 46.35 –81.83 17 0.33 0.36 0.333 0.362 0.187 0.883 Carter and others 

Suite sill (2012)

Shalozerskoe (Kuk Reef (chro- Burakovskaya Russia 62.10 36.47 78 0.07 0.05 0.17 0.175 0.0245 0.369 Fletcher (2009)

Ruchi Zone) mitite) intrusion

Sheba’s Ridge Contact Bushveld South Africa –25.32 29.36 610 0.19 0.08 0.209 0.659 0.0637 0.932 Aquarius Platinum 

Ltd. (2011)

Showing No15 Reef? Olanga Russia 66.32 30.87 0.6 n.d. n.d. 2.5 4 n.d. 6.50 Geological Survey 

(massif Luk- of Finland (2011)kulaisvaara)

Showing No2 (mas- Reef? Olanga Russia 66.32 30.87 0.7 n.d. n.d. 3 6 n.d. 9.00 Geological Survey 

sif Lukkulai- of Finland (2011)svaara)

Siika-Kämä Reef (sul- Portimo Finland 66.22 26.48 49 0.08 0.1 0.67 2.45 0.07 3.12 Geological Survey 

(SK Reef) fide) of Finland (2011)

Sompujärvi Reef (sul- Penikat Finland 65.94 25.14 6.7 n.d. n.d. 3.08 5.36 0.1 8.92 Geological Survey 

(SJ Reef) fide) intrusion of Finland (2011)

Sputnik Conduit Pechenga Russia 69.40 30.67 17 1.4 0.77 n.d. n.d. 0.01 0.0350 Geological Survey 

greenstone of Finland (2011)

belt

Stillwater and Reef (sul- Stillwater United States 45.39 –109.88 150 n.d. n.d. 4.35 10.7 n.d. 15.0 Abbott and others 

East Boulder fide) (2011)

Stormyrplutten Synvolcanic n.d. Norway 64.78 13.47 0.2 0.08 0.5 0.331 0.313 n.d. 0.644 Geological Survey 

gabbroic of Finland (2011)

body

Sudbury-Vale Sudbury Sudbury Canada 46.54 –81.00 110 1.2 1.5 0.916 1.06 0.387 2.36 Vale S.A. (2012)
N90 Critical Mineral Resources of the United States—Platinum-Group Elements Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued [Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain] Deposit name Deposit type Intrusion Country Latitude Longitude Ore (Mt) Ni (percent) Cu (percent) Pt (g/t) Pd (g/t) Au (g/t) PGE and Au (g/t) Reference(s) Decimal degrees WGS 84 Suhanko Contact Portimo Finland 66.10 26.09 1 0.2 0.31 0.2 0.9 0.03 1.10 Geological Survey of Finland (2011) Syerston Laterite Fifield Australia n.d n.d 140 n.d. n.d. 0.24 n.d. n.d. 0.240 Teluk (2001) Talnakh ore field— Conduit Noril’sk Russia 69.53 88.45 85 0.89 3.8 2.11 8.49 0.614 10.8 OJSC MMC Norilsk cuprous Talnakh Nickel (2011) Talnakh ore field— Conduit Noril’sk Russia 69.53 88.45 1,400 0.51 1 0.842 2.9 0.19 3.91 OJSC MMC Norilsk disseminated Talnakh Nickel (2011) Talnakh ore field— Conduit Noril’sk Russia 69.53 88.45 150 2.8 3.3 1.26 6.28 0.195 8.04 OJSC MMC Norilsk rich Talnakh Nickel (2011) Thunder Bay North Contact n.d. Canada 48.75 –88.90 8.2 0.21 0.29 1.22 1.15 0.0756 2.45 Cole and El-Rassi project (2009) Top and Middle Reef (chro- Panton Sill Australia –17.77 127.83 14 0.27 0.08 2.19 2.39 0.31 5.20 Platinum Australia Reefs mitite) Ltd. (2003, 2011) Tundrovskoe Conduit Pechenga Russia 69.40 30.67 110 0.51 0.26 n.d. n.d. 0.008 0.0220 Geological Survey greenstone of Finland (2011) belt UG2—eastern Reef (chro- Bushveld South Africa –24.60 30.00 3,900 0.11 0.04 2.44 2.06 0.0779 5.03 Zientek and others limb mitite) (2014) UG2—western Reef (chro- Bushveld South Africa –25.30 27.30 3,400 0.12 n.d. 2.98 1.48 0.0333 5.05 Zientek and others 

limb mitite) (2014)

Upper Pseudo Reef (sul- Bushveld South Africa –25.18 26.97 7.4 0.17 0.04 2.02 1.07 0.128 3.40 Harper and de Wit 

Reef fide) (2009); Platinum 

Australia Ltd. (2012)

Vaara Komatiitic Kuhmo Finland 65.28 29.29 8.2 0.32 0.02 0.07 0.14 n.d. 0.210 Geological Survey 

of Finland (2011)

Vaaralampi Contact Portimo Finland 66.10 26.18 32 0.31 0.2 0.25 0.68 0.07 0.930 Geological Survey 

of Finland (2011)

Verkhnee Conduit Pechenga Russia 69.40 30.63 63 0.5 0.24 n.d. n.d. 0.02 0.099 Geological Survey 

greenstone of Finland (2011)

belt

Vodorazdelny Conduit Kun-Manie Russia 55.35 132.44 5.9 0.71 0.2 0.2 0.2 n.d. 0.300 Armitage (2007)

Massif

Volspruit project— Contact Bushveld South Africa –24.35 28.95 51 0.14 0.04 0.556 0.667 0.0222 1.24 Clay and others 

North Pit Mines (2010)
Table N10 N91

Table N10. Grade and tonnage of mineralized rock remaining in platinum-group-element-bearing mineral deposits.—Continued[Negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere. WGS 84, World Geodetic System of 1984. Elements: Au, gold; Cu, copper; Ni, nickel; Pd, palladium; PGE, platinum-group element(s); Pt, platinum. Abbreviations: Mt, million metric tons; g/t, gram per metric ton; LC, lower chromitite; LG, lower grade reef; LM, lower main reef; MMW, mid-main waste zone; MMZ, main mineralized zone; MR, mid reef; PCMZ, peridotite chromititic mineralized zone; Ti-V, titaniferous-vanadiferous; UC, upper chromitite; UM, upper main reef; UUM, post reef hanging wall; n.d., no data; ?, identification uncertain]

Deposit Latitude Longitude Ore Ni Cu Pt Pd Au PGE and Au Deposit type Intrusion Country Reference(s) name Decimal degrees WGS 84 (Mt) (percent) (percent) (g/t) (g/t) (g/t) (g/t)

Vostochno (east- Reef (sul- Imandra- Russia 67.40 35.95 7.6 0.23 0.17 n.d. n.d. n.d. 13.2 Geological Survey 

ern) Panskoe fide) Varzuga of Finland (2011)

greenstone belt

Vostochnoe—criti- Reef? Olanga Russia 66.32 30.90 0.52 n.d. n.d. 3.77 0.07 n.d. 3.84 Geological Survey 

cal zone of Finland (2011)

War Springs Contact Bushveld South Africa –24.22 29.05 47 n.d. n.d. 0.255 0.777 0.0777 1.11 Platmin Ltd. (2011)Wellgreen (indicat- Conduit Wellgreen Canada 61.47 –139.52 460 0.32 0.26 0.399 0.343 0.171 0.913 McCracken (2011); 

ed and inferred Carter and

pitshell) others (2012)

Western Panskoe Reef (sul- Imandra- Russia 67.53 35.27 35 0.1 0.25 n.d. n.d. n.d. 3.65 Geological Survey 

fide) Varzuga of Finland (2011)

greenstone belt

Yangliuping Conduit Yangliuping China 30.70 101.90 90 0.44 0.26 0.23 0.3 n.d. 0.563 Yao and others (2001); 

deposit Sill Song, Zhou, Cao, 

and others (2003); Song, Zhou, and Cao (2004)

Yli-Portimo Contact Portimo Finland 66.13 26.17 30 n.d. n.d. 0.18 1.12 0.11 1.30 Geological Survey 

of Finland (2011)

Zapolyarnoe Conduit Pechenga Russia 69.40 30.75 11 2.3 1.2 n.d. n.d. 0.008 0.430 Geological Survey 

greenstone of Finland (2011)

belt

Zhdanovskoe Conduit Pechenga Russia 69.44 30.67 1,000 0.57 0.25 n.d. n.d. 0.008 0.0580 Geological Survey 

greenstone of Finland (2011)

belt

Zimplats Hartley Reef (sul- Great Dyke Zimbabwe –17.99 30.45 220 n.d. n.d. 2 1.55 0.297 4.02 Impala Platinum 

Mine (Main fide) Holdings Ltd. 

Sulphide Zone) (2011)

Zimplats Ngezi Reef (sul- Great Dyke Zimbabwe –18.42 30.41 1,700 n.d. n.d. 1.77 1.37 0.262 3.55 Impala Platinum 

Mine (Main fide) Holdings Ltd. 

Sulphide Zone) (2011)

Rare-Earth Elements

By Bradley S. Van Gosen, Philip L. Verplanck, Robert R. Seal II, Keith R. Long, 

and Joseph Gambogi

Chapter O of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–O

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter O

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................O1

Introduction..................................................................................................................................................O2

Uses and Applications ......................................................................................................................O3

Availability of Supply .........................................................................................................................O4

Australia .....................................................................................................................................O4

China ...........................................................................................................................................O4

India.............................................................................................................................................O5

Malaysia.....................................................................................................................................O5

Russia..........................................................................................................................................O5

United States .............................................................................................................................O5

Conservation, Recycling, and Substitution ...........................................................................O5

Geology.........................................................................................................................................................O6

Geochemistry......................................................................................................................................O6

Rare-Earth-Element-Mineral Formation in Carbonatite Magmas ....................................O7

Rare-Earth-Element-Mineral Formation in Alkaline Magmas ...........................................O7

Rare-Earth-Element-Mineral Formation in Parental Magma Sources ............................O7

Rare-Earth-Element-Mineral Formation Through Hydrothermal Processes ..................O7

Aqueous Geochemistry of Rare-Earth Elements .................................................................O7

Mineralogy..........................................................................................................................................O7

Deposit Types ...................................................................................................................................O10

Resources and Production ......................................................................................................................O17

Identified Resources .......................................................................................................................O17

Undiscovered Resources ...............................................................................................................O18

Exploration for New Deposits .................................................................................................................O18

Environmental Considerations ................................................................................................................O20

Sources and Fate in the Environment ..........................................................................................O21

Mine Waste Characteristics ..........................................................................................................O21

Human Health Concerns .................................................................................................................O22

Ecological Health Concerns ...........................................................................................................O22

Carbon Footprint...............................................................................................................................O22

Mine Closure.....................................................................................................................................O23

Problems and Future Research ..............................................................................................................O23

References Cited.......................................................................................................................................O24

Suggested citation:

Van Gosen, B.S., Verplanck, P.L., Seal, R.R., II, Long, K.R., and Gambogi, Joseph, 2017, Rare-earth elements, chap. O 

of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—

Economic and environmental geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802,

p. O1– O31, http://dx.doi.org/10.3133/pp1802O.
Rare-Earth Elements

By Bradley S. Van Gosen, Philip L. Verplanck, Robert R. Seal II, Keith R. Long, and Joseph Gambogi

Abstract

The rare-earth elements (REEs) are 15 elements that 

range in atomic number from 57 (lanthanum) to 71 (lutetium); 

they are commonly referred to as the “lanthanides.” Yttrium 

(atomic number 39) is also commonly regarded as an REE 

because it shares chemical and physical similarities and has 

affinities with the lanthanides. Although REEs are not rare in 

terms of average crustal abundance, the concentrated deposits 

of REEs are limited in number.

Because of their unusual physical and chemical 

properties, the REEs have diverse defense, energy, industrial, 

and military technology applications. The glass industry is 

the leading consumer of REE raw materials, which are used 

for glass polishing and as additives that provide color and 

special optical properties to the glass. Lanthanum-based 

catalysts are used in petroleum refining, and cerium-based 

catalysts are used in automotive catalytic converters. The 

use of REEs in magnets is a rapidly increasing application. 

Neodymium-iron-boron magnets, which are the strongest 

known type of magnets, are used when space and weight are 

restrictions. Nickel-metal hydride batteries use anodes made 

of a lanthanum-based alloys.

China, which has led the world production of REEs 

for decades, accounted for more than 90 percent of global 

production and supply, on average, during the past decade. 

Citing a need to retain its limited REE resources to meet 

domestic requirements as well as concerns about the 

environmental effects of mining, China began placing 

restrictions on the supply of REEs in 2010 through the 

imposition of quotas, licenses, and taxes. As a result, the 

global rare-earth industry has increased its stockpiling of 

REEs; explored for deposits outside of China; and promoted 

new efforts to conserve, recycle, and substitute for REEs. 

New mine production began at Mount Weld in Western 

Australia, and numerous other exploration and development 

projects noted in this chapter are ongoing throughout 

the world.

The REE-bearing minerals are diverse and often complex 

in composition. At least 245 individual REE-bearing minerals 

are recognized; they are mainly carbonates, fluorocarbonates, 

and hydroxylcarbonates (n=42); oxides (n=59); silicates 

(n=85); and phosphates (n=26).

Many of the world’s significant REE deposits occur in 

carbonatites, which are carbonate igneous rocks. The REEs also 

have a strong genetic association with alkaline magmatism. 

The systematic geologic and chemical processes that explain 

these observations are not well understood. Economic or 

potentially economic REE deposits have been found in 

(a) carbonatites, (b) peralkaline igneous systems, (c) magmatic 

magnetite-hematite bodies, (d) iron oxide-copper-gold (IOCG) 

deposits, (e) xenotime-monazite accumulations in mafic gneiss, 

(f) ion-absorption clay deposits, and (g) monazite-xenotimebearing placer deposits. Carbonatites have been the world’s 

main source for the light REEs since the 1960s. Ion-adsorption 

clay deposits in southern China are the world’s primary source 

of the heavy REEs. Monazite-bearing placer deposits were 

important sources of REEs before the mid-1960s and may be 

again in the future. In recent years, REEs have been produced 

from large carbonatite bodies mined at the Mountain Pass 

deposit in California and, in China, at the Bayan Obo deposit 

in Nei Mongol Autonomous Region, the Maoniuping deposit 

in Sichuan Province, the Daluxiang deposit in Sichuan 

Province, and the Weishan deposit in Anhui Province. Alkaline 

igneous complexes have recently been targeted for exploration 

because of their enrichments in the heavy REEs.

Information relevant to the environmental aspects of REE 

mining is limited. Little is known about the aquatic toxicity of 

REEs. The United States lacks drinking water standards for 

REEs. The concentrations of REEs in environmental media are 

influenced by their low abundances in crustal rocks and their 

limited solubility in most groundwaters and surface waters. 

The scarcity of sulfide minerals, including pyrite, minimizes or 

eliminates concerns about acid-mine drainage for carbonatitehosted deposits and alkaline-intrusion-related REE deposits. 

For now, insights into environmental responses of REE mine 

wastes must rely on predictive models.
O2 Critical Mineral Resources of the United States—Rare-Earth Elements

Introduction

Until recently, the rare-earth elements (REEs) were

familiar to a relatively small number of people, such as chemists,

some geologists, and specialized materials and technology 

engineers. In the 21st century, the REEs have gained considerable visibility through news media and Internet coverage 

because of (a) recognition of the critical, specialized properties 

that the REEs contribute in components of our modern technology, combined with (b) the near monopoly on the production 

and supply of the REEs held by one country (China), and 

(c) the world’s heavy dependence on this controlled supply 

from China (Long, 2011; Tse, 2011; Hatch, 2012a).

Since the late 1990s, China has provided more than 

90 percent of the world’s supply of the REEs, on average 

(Long and others, 2010; Tse, 2011; Hatch, 2012a). In 2010, 

China announced that it intended to reduce its exports of REEs 

(Hatch, 2012a). At the same time, the use of REEs in clean 

energy and defense technologies was continuing to increase 

(Goonan, 2011; Long, 2011; Tse, 2011; Hatch, 2012a). China’s 

announcement regarding its REE export quotas raised concerns 

among nations that are highly dependent on specialized 

technologies, such as Japan, the United States, and members 

of the European Union. These countries, as well as the World 

Trade Organization, formally protested China’s reduced export 

quotas for REEs (Hatch, 2012a), and China agreed in 2014 to 

remove the restrictions. Nonetheless, the political and economic 

issues surrounding the global supply of REEs in the early 21st 

century has given these metals more visibility and a greater 

appreciation of their strategic importance than ever before. 

As a consequence, exploration activities to discover economic 

deposits of REEs and efforts to bring them into production 

increased dramatically. One industry analyst stated that, as of 

September 2012, during a high period of global exploration for 

REE deposits, his company was “tracking a total of 442 rareearth projects in progress associated with 270 different 

companies in 37 different countries” (Hatch, 2012b).

In recent years, expert panels convened by research 

institutes and Government agencies highlighted specific 

REEs as raw materials critical to evolving technologies, 

such as clean-energy applications, electronics, and high-tech 

military components. These reports also suggest that a high 

potential exists for disruptions in the supply of these REEs. 

As a result, several expert panel analyses ranked REEs high 

on the “criticality” factor of raw materials—that is, high 

importance to technologies and economies combined with 

high risk for supply disruptions. Recent reports by panels and 

agencies tasked to assess the criticality of REEs and other 

raw materials include those by the National Research Council 

(2008), U.S. Department of Energy (2010, 2011), European 

Commission (2014), American Physical Society (APS) Panel 

on Public Affairs and Materials Research Society (MRS) 

(2011), and the Resnick Institute (2011).

The REEs comprise 15 elements (table O1) that range in 

atomic number from 57 (lanthanum) to 71 (lutetium) on the 

periodic table of elements. They are commonly referred to as 

the “lanthanides.” The REE promethium (atomic number 61) 

is not included in discussions of REE deposits because the 

element is extremely rare and unstable in nature. Yttrium 

(atomic number 39) is commonly regarded as an REE because 

it shares chemical and physical similarities and has affinities 

with the lanthanides and because it typically occurs in the 

same deposits as other REEs.

Traditionally, the REEs are divided into the following 

two groups on the basis of atomic weight: (a) the light 

REEs (LREEs), which include lanthanum through gadolinium (atomic numbers 57 through 64), and (b) the heavy 

REEs (HREEs), which include terbium through lutetium 

(atomic numbers 65 through 71) (table O1). Yttrium, although 

light (atomic number 39), is included with the HREE group 

because of its common chemical and physical affiliations with 

the other HREEs.

Table O1. List of the rare-earth elements found in natural 

deposits—the “lanthanides” plus yttrium.

[Average crustal abundance values are from Lide (2004, p. 17); REE, rareearth element]

Element1 Symbol Atomic

number

Atomic

weight

Crustal 

abundance

(part per 

million)

Light REEs

Lanthanum La 57 138.91 39

Cerium Ce 58 140.12 66.5

Praseodymium Pr 59 140.91 9.2

Neodymium Nd 60 144.24 41.5

Samarium Sm 62 150.36 7.05

Europium Eu 63 151.96 2.0

Gadolinium Gd 64 157.25 6.2

Heavy REEs

Yttrium Y 39 88.91 33

Terbium Tb 65 158.92 1.2

Dysprosium Dy 66 162.50 5.2

Holmium Ho 67 164.93 1.3

Erbium Er 68 167.26 3.5

Thulium Tm 69 168.93 0.52

Ytterbium Yb 70 173.04 3.2

Lutetium Lu 71 174.97 0.8

1

Promethium (Pm, atomic number=61) is not included in this list because it 

is extremely rare in nature.
Introduction O3

Most of the REEs are not as rare as the group’s name 

suggests. They were named “rare-earth” elements because most 

were identified during the 18th and 19th centuries as “earths” 

(originally defined as materials that could not be changed 

further by the sources of heat), and, in comparison with other 

“earths,” such as lime or magnesia, they were relatively rare. 

Cerium is the most abundant REE (table O1), and it is actually 

more common in Earth’s crust than is copper or lead (Lide, 

2004, p. 17). All the REEs except promethium are more 

abundant than silver, gold, or platinum in Earth’s crust, on 

average (Lide, 2004, p. 17). Thus, REEs are not rare in terms 

of average crustal abundance, but concentrated and economic 

deposits of REEs are unusual (Adams and Staatz, 1973). 

Ore deposits of REEs are found in unusual rock types and 

uncommon minerals, as described in the sections that follow.

Uses and Applications

To take advantage of their unusual physical and chemical 

properties, the REEs are used in a variety of industrial and 

technology applications (Goonan, 2011; Long, 2011). Given 

their similar chemical nature, many different REEs have related 

or complementary uses; thus, it is more convenient to describe 

their uses by application rather than by individual element. In 

general, the lighter REEs and yttrium are cheaper, are produced 

in greater quantities, and are more extensively used than the 

heavier REEs. The least common and most expensive REEs, 

from holmium to lutetium, are limited to a very few, highly 

specialized, high-technology applications. Goonan (2011) 

provided a quantitative study of domestic REE consumption, 

including estimated consumption of individual REEs as of 

2008, addressed according to consumption categories.

The glass industry is the leading consumer of REE 

raw materials (Goonan, 2011), principally for purposes of 

glass polishing and as additives to provide color and special 

optical properties to glass. Cerium oxide is widely used in 

the production of glass types that require a precision polish, 

such as flat panel display screens. Cerium is also used to 

decolorize glass. Lanthanum and lutetium greatly increase the 

refractive index of optical glass. Lanthanum is widely used in 

camera lenses, whereas the more expensive lutetium is used in 

immersion lithography, which requires a high-refractive index. 

Erbium, holmium, neodymium, praseodymium, ytterbium, and 

yttrium are used as special colorants and to provide filtering 

and glare-reduction qualities for glass. Europium is a common 

dopant (doping agent) for optical fibers.

Catalysts are another major use for REEs. They include 

lanthanum-based catalysts used in petroleum refining and 

cerium-based catalysts used in automotive catalytic converters 

(Goonan, 2011). Catalysts enriched in REEs are essential 

to cracking (breaking down) heavy hydrocarbon molecules 

into smaller molecules, which enables petroleum refineries to 

obtain significantly more product per barrel of oil processed. 

Small amounts of neodymium, praseodymium, and yttrium are 

used as catalysts in catalytic converters to reduce automotive 

carbon monoxide emissions.

Permanent magnets made from alloys of REEs are a 

recent, rapidly growing application. Neodymium-iron-boron 

magnets are the strongest magnets known; they are used 

wherever space and weight are at a premium. Significant 

uses include hard disk drives, cell phones, electric motors 

for hybrid vehicles and windmills, and actuators in aircraft 

(Goonan, 2011; Long, 2011). Lesser amounts of dysprosium, 

gadolinium, and praseodymium are also used in these 

magnets. Dysprosium is of particular importance because 

substituting it for a small portion of neodymium improves 

high-temperature performance and resistance to demagnetization. Permanent magnets are used in such demanding 

applications as electric motors for hybrid cars and wind 

turbines. Samarium-cobalt magnets, though less powerful than 

neodymium magnets, have better heat tolerance and are used 

in lieu of neodymium magnets where heat stress is an issue.

Nickel-metal hydride batteries use a lanthanum-based 

alloy for anodes (Kopera, 2004). Found in many rechargeable 

consumer products and some hybrid vehicles, these battery 

alloys are also a significant application of cerium, neodymium, 

praseodymium, and samarium. Demand for REEs in battery 

applications is expected to decrease as lithium ion batteries 

displace nickel-metal hydride batteries (Anderson and PatiñoEcheverri, 2009). Nickel-metal hydride batteries in hybrid 

electric cars are significant users of lanthanum, however, 

employing as much as 10 to 15 kilograms per vehicle.

Cerium, lanthanum, neodymium, and praseodymium, 

commonly in the form of a mixed oxide known as mischmetal, 

are used in steelmaking to remove impurities, as well as in 

the production of special steel alloys (Goonan, 2011). These 

REEs, along with yttrium, individually or in combination, are 

also used in various special alloys of chromium, magnesium, 

molybdenum, tungsten, vanadium, and zirconium.

Many REEs, especially yttrium, cerium, lanthanum, europium, and terbium, are used individually or in combination to 

make phosphors for many types of cathode ray tube and flat 

panel display screens, and in some incandescent, fluorescent, 

and light-emitting diode lighting (Goonan, 2011; Long, 2011). 

Gadolinium phosphors are used in X-ray imaging and various 

medical applications, such as magnetic resonance imaging 

(MRI). Yttrium, lanthanum, cerium, neodymium, and praseodymium are used as pigments for ceramics.

Rare-earth elements are also used in synthetic gems, 

crystals for lasers, microwave equipment, superconductors, 

sensors, nuclear control rods, and cryo-coolers. Neodymium 

is the active constituent in a popular fertilizer in China. 

Significant potential new uses for REEs include their use as 

nanofilters and in memory devices, power converters, optical 

clocks, infrared decoy flares, and fusion energy. Future 

research is likely to find many new uses for the REEs.
O4 Critical Mineral Resources of the United States—Rare-Earth Elements

Availability of Supply

China has been the leading producer of REEs for 

decades (fig. O1), and since the late 1990s, it has accounted 

for more than 90 percent of global production, on average. 

When China began restricting the supply of REEs through 

the imposition of quotas, licenses, and taxes in 2010, 

it cited limited resources for domestic requirements and 

environmental concerns as the reasons (Tse, 2011). China’s 

restrictions on REEs changed the world rare-earth industry 

in several ways, including by fostering increased stockpiling 

of REEs; increased exploration and development of deposits 

outside of China; and new efforts to conserve, recycle, and 

find substitutes for and among the REEs. This is evidenced 

by new mine production that began in Australia in 2011 and 

in the United States in 2012, and by additional exploration 

and development projects in many other countries (Gambogi, 

2016). Global reserves of REEs were estimated on a rareearth-oxide (REO) basis to be 130 million metric tons and 

were led by, in decreasing order of reserves, China, Brazil, 

Australia, and India (Gambogi, 2015).

Australia

Heavy-mineral sands produced in Australia contain 

significant quantities of REEs; however, monazite has not 

been produced from heavy-mineral sands operations since 

the 1990s to avoid the concentration of naturally occurring 

radioactive minerals, primarily owing to the presence of 

thorium in monazite. Lynas Corporation Ltd. of Australia 

began production of rare-earth mineral concentrates at its 

Mount Weld carbonatite complex in Western Australia in 

2011. Shipments of mineral concentrates were delayed until 

2013 while processing operations were being commissioned in 

Malaysia. Lynas later increased its capacity to produce mineral 

concentrates to 22,000 metric tons per year of contained 

REO. Other REE projects underway in Australia include 

the Nolans Bore property in the Northern Territory and the 

Dubbo Zirconia deposit in New South Wales, both of which 

were at the advanced feasibility study stage (Gambogi, 2016). 

Australia’s reserves of REEs are estimated to be 3.2 million 

metric tons of REO (Gambogi, 2015).

China

China’s rare-earth mine production takes place in the 

Provinces of Fujian, Guangdong, Hunan, Jiangxi, Shandong, 

Sichuan, and Yunnan, and in the Autonomous Regions of 

Guangxi and Nei Mongol. The largest mining operation is 

Baotou Rare Earth’s Bayan Obo Mine, which produces iron 

ore as well as bastnaesite and monazite as the main rare-earth 

minerals. REE-bearing ion-adsorption clays are mined in various 

locations in the southeastern Provinces. In contrast to bastnaesite 

ores, ion-adsorption clays are valued for their medium REE and 

HREE content. Reserves of REEs in China are estimated to be 

55 million metric tons of contained REO (Gambogi, 2015).

REE figure 1. 

140,000

120,000

100,000

80,000

60,000

40,000

20,000

0

Rare-earth oxide mine production, in metric tons

1960 1965 1970 1975 1980 1985

Year

China

United States

Other countries

EXPLANATION

1990 1995 2000 2005 2010

Figure O1. Graph showing world mine production of rare-earth oxides, by country and year, from 

1960 to 2012. The layers of the graph are placed one above the other, forming a cumulative total. Data 

are from U.S. Bureau of Mines (1961–96) and U.S. Geological Survey (1997–2016).
Introduction O5

The leading export markets for China’s REEs are 

Japan, the United States, and France. The supply of REEs 

to consumers outside of China is determined not by mine 

capacity but by production and export quotas set by China’s 

Ministry of Commerce (MOFCOM) and the Ministry of 

Industry and Information Technology. Supply concerns were 

raised when the MOFCOM set China’s export quota for 

the second half of 2010 at 7,976 metric tons (a 64 percent 

decrease compared with the quota set in the first half of 2010), 

which reduced the total exports for 2010 to 20,258 metric 

tons (a 40 percent decrease compared with exports for the 

previous year). Following the sudden drop in exports, prices 

for rare-earth products in 2010 increased sharply (Ministry of 

Commerce 2010a, b). Export quotas for 2011, 2012, and 2013 

were set at 30,246 metric tons, 30,996 metric tons, and 30,999 

metric tons, respectively. Prices for most rare-earth products 

peaked in mid-2011. In 2014, China’s rare-earth export 

quotas were 30,611 metric tons, including 27,006 metric tons 

for LREEs and 3,605 metric tons for HREEs (Hatch, 2013, 

2014). Chinese-owned companies were allocated 74 percent 

of the rare-earth products export quota, and joint ventures 

with foreign partners were allocated 26 percent of the quota. 

Following a World Trade Organization ruling in 2014, China 

announced the end of export restrictions on rare earths (Yap, 

2015). According to the Ministry of Industry and Information 

Technology, the Government has ongoing efforts to reduce 

illegal mine production of rare earths, which was estimated to 

account for more than 40,000 metric tons of REE production 

in a single year (Global Times, 2013).

India

In India, REEs are produced from monazite contained in 

heavy-mineral sands. The two Government-owned producers 

are the Rare Earth Division of Indian Rare Earths Ltd. (IREL) 

and Kerala Minerals and Metals Ltd. (KMML). Heavy-mineral 

operations that produce monazite concentrate include IREL’s 

Manavalakurichi operation in the State of Tamil (which had a 

capacity of 6,000 metric tons per year) and KMML’s Chavara 

operation in Kerala State (240 metric tons per year) (Indian 

Bureau of Mines, 2015). IREL’s capacity to produce rare-earth 

compounds from monazite was limited, but was reported to 

be increasing. India’s reserves of REEs are estimated to be 

3.1 million metric tons of REO (Gambogi, 2015).

Malaysia

In the Malaysian States of Perak and Selangor, 

REE-bearing

monazite and xenotime are recovered from cassiterite mine 

tailings, which are referred to by the Malaysian mining 

industry as “amang.” Recent monazite production from amang 

was less than 500 metric tons per year. Malaysia’s reserves of 

REEs are estimated to be 30,000 metric tons of REO, which is 

less than 1 percent of global reserves (Gambogi, 2015, 2016).

Russia

In Russia, loparite mineral concentrates are produced at the 

Lovozero mining operation on the Kola Peninsula, Murmanskaya 

Oblast; the concentrates are shipped to the Solikamsk magnesium plant in Permskiy Kray. Rare-earth-bearing residues from 

the Solikamsk plant are exported for recovery of REEs. The 

Lovozero operation had the capacity to produce an estimated 

3,700 metric tons per year of REOs contained in mineral 

concentrates. Within a global total of 130 million metric tons, 

Russia’s reserves contributed to the 41 million metric tons of 

reserves from other counties (Gambogi, 2015).

United States

From 2012 to 2015, the United States was a leading 

producer of rare-earth minerals outside of China. All mine 

production of REEs during this period was from Molycorp, 

Inc.’s Mountain Pass Mine in San Bernardino County, 

California. Mine production at Mountain Pass began in 

1952 but was discontinued from 2002 to 2011 because of 

low rare-earth prices and environmental permitting issues. 

In 2012, Molycorp commissioned new mine and processing 

operations that had an initial planned (but never fully realized) 

production capacity of 19,000 metric tons per year of REOs. 

Because of substantial decreases in the market prices of the 

REEs, however, Molycorp idled operations at the Mountain 

Pass Mine in 2015. Exploration and development assessment 

projects in the United States include Bokan Mountain, Alaska; 

La Paz, Arizona; Diamond Creek, Idaho; Lemhi Pass in Idaho 

and Montana; Pea Ridge, Missouri; Elk Creek, Nebraska; 

Thor Mine, Nevada; Round Top, Texas; and Bear Lodge, 

Wyoming. Domestic reserves of REOs were estimated to be 

1.8 million metric tons (Gambogi, 2015).

Conservation, Recycling, and Substitution

Concerns about the availability of rare-earth supplies 

have stimulated efforts in recycling, substitution, and 

improving material efficiencies. These efforts are being led 

by the leading consumers of REEs.

In the United States, the Critical Materials Institute (a 

partnership of academia, Government, and industry) is working 

to diversify production, reduce waste, and develop substitutes 

for rare earths. The collaboration is expected to spend up to 

$120 million during a period of 5 years to support research 

focused on ensuring a reliable supply of rare earths and other 

critical materials (U.S. Department of Energy, 2013).

Belgian-based Umicore Group has developed a batteryrecycling program for nickel-metal-hydride (NiMH) batteries 

at its new recycling plant in Hoboken, Belgium. After the 

separation of nickel and iron, the company plans to process the 

REEs into a high-grade concentrate that would be separated 

into rare-earth materials at Rhodia Rare Earth Systems’ plant 

at La Rochelle, France (Umicore Group, 2011).
O6 Critical Mineral Resources of the United States—Rare-Earth Elements

In the European Union, a consortium is investigating the 

extraction of rare-earth metals from permanent magnets in electric vehicle motors. The project is funded by the German Federal 

Research Ministry, and the consortium is led by Siemens AG 

(Recycling International, 2011b). German research agency 

Fraunhofer Institute for Surface Engineering and Thin Films is 

examining the recycling of rare-earth metals from disused lightmetal alloys and from lanthanum from fluid catalytic cracking 

(FCC) catalysts (Recycling International, 2011a).

In France, Rhodia formed an alliance with Umicore to 

recycle rare-earth minerals from NiMH rechargeable batteries. 

Rhodia also planned to recycle rare-earth minerals from lowenergy light bulbs and magnets. By yearend 2012, recycling 

plants in La Rochelle and Saint-Fons, Rhône-Alpes Region, 

were operational. The company planned to conserve on its 

consumption of terbium by using reformulated phosphors 

(Recycling International, 2011b, c).

In Japan, Honda Motor Co., Ltd. has been working with 

Japan Metals & Chemicals Co. to recapture rare-earth metals 

from NiMH batteries. Honda expects to recover 80 percent of 

rare-earth metals contained in the used NiMH batteries and 

plans to collect the batteries globally (Honda Motor Co., Ltd., 

2013). Japan’s Hitachi, Ltd. is developing methods to recycle 

rare-earth magnets from hard disk drives, air conditioners, and 

other types of compressors. Hitachi Metals also has accelerated development of reduced-dysprosium sintered magnets 

using a dysprosium vapor-deposition and diffusion technology 

(Baba and others, 2013, p. 453).

In Vietnam, Japan’s Shin-Etsu Chemical Co., Ltd. 

constructed processing plants for the separation of rare earths 

and production of rare-earth magnets. The processing plants 

enable the recycling of rare-earth magnets used in appliances, 

such as air conditioner compressor motors (Showa Denko K.K., 

2010; Shin-Etsu Chemical Co., Ltd., 2014).

Geology

Geochemistry

The REEs are commonly found together in Earth’s crust 

because they share a trivalent charge (REEs3+) and similar 

ionic radii. Exceptions to the trivalent charge are cerium, 

which may occur in a Ce4+ valence state, and europium, which 

can exist as Eu2+. The lanthanides (lanthanum to lutetium) 

(table O1) have a unique property—with each increase in 

atomic number across the group, an electron is added to an 

inner incomplete subshell (4 f ) rather than to an outer level. 

This property of selectively filling inner levels results in 

a progressive decrease in the ionic radius of the trivalent 

lanthanide ions from La3+ to Lu3+ (fig. O2), referred to as 

“lanthanide contraction.” The systematic decrease in ionic 

radii across the lanthanide group (fig. O2) has two significant 

consequences: (a) it affects the fractionation of the REEs to 

a high degree, which allows for their individual separation in 

geologic environments and in ore processing, and (b) it lends 

12.5

12.0

11.5

11.0

10.5

10.0

9.5

9.0

13.0

Ionic radius, in nanometers

Y

La

Ce

Ce

Pr

Nd

Sm

Cerium ion (Ce4+)

Europium ion (Eu2+)

Rare-earth element 

 trivalent ion (3+)

Eu

Element by atomic number

Eu

Gd

Tb

Dy

Ho

Er

Tm

Yb Lu

REE figure 2

EXPLANATION

Figure O2. Graph showing radii (in nanometers) of the 

trivalent (3+) ions of the rare-earth elements (REEs) as well as 

of cerium in the +4 valence state (Ce4+) and europium in the 

+2 valence state (Eu2+). The graph illustrates the systematic 

decrease in the radius of the trivalent REE ions with 

increasing atomic number. Data plotted are for coordination 

number VIII. Data are from Shannon (1976). Elements, in order 

of increasing atomic number: Y, yttrium; La, lanthanum; Ce, cerium; 

Pr, praseodymium; Nd, neodymium; Sm, samarium; Eu, europium; 

Gd, gadolinium; Tb, terbium; Dy, dysprosium; Ho, holmium; Er, 

erbium; Tm, thulium; Yb, ytterbium; Lu, lutetium. Promethium (Pm) 

is not included in the list because it is extremely rare in nature.

the REEs unusual and useful chemical and physical properties, 

such as unique magnetic and (or) optical properties.

In magmatic systems, the high-charge of the REE ions 

impedes the ability of these elements to achieve charge balance 

and fit into the structure of the common rock-forming minerals, 

which have coordination sites best suited for valence states 

of 2 or 1. As a result, when common silicate minerals (such 

as amphiboles, feldspars, and olivine) crystallize, most of 

the REEs tend to remain in the coexisting melt. (In contrast, 

europium is often depleted in magmas because it is incorporated 

into feldspars owing to its Eu2+ valence state.) Successive 

generations of this process, referred to as crystal fractionation, 

increase the concentrations of REEs in the remaining melt 

until individual REE-mineral-rich phases crystallize. In addition to ionic charge (valence state), factors that can affect the 

partitioning of the REEs between a mineral and the coexisting 

magma can include temperature, pressure, fluid composition, 

magmatic convection, and variation in mineral settling velocities (Henderson, 1984; McKay, 1989; Kogarko and others, 

2006). Elements, such as the REEs, that do not tend to participate in the early mineral-formation processes are referred to as 

incompatible elements. In magmatic systems, other incompatible elements may also include hafnium, niobium, phosphorus, 

tantalum, thorium, titanium, scandium, uranium, and zirconium.
Geology O7

Rare-Earth-Element-Mineral Formation in 

Carbonatite Magmas

Many of the world’s REE deposits are associated with 

carbonatites, which are igneous rocks derived from carbonaterich magmas rather than silica-rich magmas. Carbonatites 

are defined as igneous rocks with greater than 50 modal 

percent carbonate minerals, usually calcite and dolomite, 

and, in general, they have the highest REE concentrations 

of any igneous rocks (Chakhmouradian and Zaitsev, 2012). 

Carbonatite magmas are derived from the mantle, but the 

processes responsible for the formation of carbonatite magmas 

are in debate. Possible processes responsible for carbonatite 

magma generation include (a) primary mantle melts or 

(b) evolution from a mantle-derived alkali melt by crystal 

fractionation. As the carbonatite magma crystalizes, REEs can 

be concentrated in the primary REE mineral phases or they 

can become enriched in late-stage magmas or fluids. Latestage REE-rich magmas can result in ore-grade REE mineralization that contains very large concentrations of LREEs.

Rare-Earth-Element-Mineral Formation in 

Alkaline Magmas

In addition to carbonatites, the REEs have a strong genetic 

association with alkaline igneous processes, particularly 

peralkaline magmatism. Alkaline igneous rocks crystallize 

from silicate magmas and fluids enriched in alkali elements 

and precipitate sodium- and potassium-bearing minerals, such 

as sodium- or potassium-rich amphiboles and pyroxenes. 

Alkaline igneous rocks are typically undersaturated in silica, 

and they therefore contain little or no quartz. Peralkaline 

igneous rocks, which are defined as those in which the 

molecular proportion of aluminum oxide (Al2O3) is less than 

that of sodium oxide (Na2O) and potassium oxide (K2O) 

combined (that is, Al2O3<Na2O+K2O), are associated with 

enrichments in REEs, typically more than most other igneous 

rocks.

Rare-Earth-Element-Mineral Formation in 

Parental Magma Sources

The REEs occur in modest concentrations in the upper 

mantle but by some processes can become highly enriched in 

carbonate and alkaline magmas that ascend from the mantle 

to the crust (Arzamastsev and others, 2001). A discussion of 

what geologic processes separate, enrich, and mobilize REEs 

in the upper mantle, ultimately forming ore-grade deposits of 

REE minerals in the crust, is beyond the scope of this chapter. 

Wyllie and others (1996) provide a comprehensive review 

and discussion of many studies that focus on the genesis of 

carbonatite magmas and their REE minerals. An overview of 

recent research toward REE sources and processes is discussed 

by Chakhmouradian and Zaitsev (2012).

Rare-Earth-Element-Mineral Formation Through 

Hydrothermal Processes

The formation of REE minerals in hydrothermal environments is a less investigated aspect of their formation, but is 

discussed by Gieré (1996). The mobilization of the REEs by 

hydrothermal processes is summarized by Williams-Jones and 

others (2012).

Aqueous Geochemistry of Rare-Earth Elements

With the development of inductively coupled plasmamass spectrometry (ICP–MS), analysis of REEs in aqueous 

solutions has become routine. As a result, the number of 

studies that analyze and interpret REEs has increased substantially in the past 15 years. Similar to many trace metals, the 

REEs occur as cations, which generally become less soluble 

with increasing pH (Verplanck and others, 2004). The REEs 

have a charge of +3, but under oxidizing conditions, cerium 

can have a charge of +4. Under reducing conditions, europium 

can have a +2 charge, but europium in the +2 oxidation state is 

generally limited to high-temperature environments.

The behavior of REEs in aqueous solutions can be described 

as a competition between the formation of stable aqueous 

complexes and the tendency of REEs to partition to solid phases, 

either by adsorption or (co)precipitation. The dominant inorganic 

ligands include carbonate, fluoride, hydroxide, phosphate, and 

sulfate. Other factors that potentially affect the partitioning 

between solid and aqueous phases include the composition 

of solid phases, pH, temperature, and redox conditions. The 

behavior of REEs in aqueous solutions is particularly relevant 

to REE ore deposit genesis in both magmatic and surficial 

environments because the stability of aqueous REE complexes 

plays an important role in their mobilization, transport, 

fractionation, and deposition (Williams-Jones and others, 

2012).

Mineralogy

The minerals that contain REEs are numerous, diverse, 

and often complex in composition. Jones and others (1996, 

appendix A) list 245 individual REE-bearing minerals, which 

include silicates (in which category the total number of 

REE-bearing minerals is 85); oxides (59); carbonates, fluorocarbonates, and hydroxylcarbonates (42); phosphates (26); 

arsenates, sulfates, and vanadates (19); halides (6); uranylcarbonates and uranyl-silicates (5); and borates (3). Table O2 

lists the REE- and yttrium-bearing minerals that historically 

have been important ore minerals for these elements.

Bastnaesite (also spelled bastnäsite) may be considered 

the most important REE mineral because it is the primary 

ore mineral of the world’s two largest REE deposits—the 

carbonatite deposit at Mountain Pass in California (fig. O3) 

and the iron-carbonatite deposits of Bayan Obo in China. 
O8 Critical Mineral Resources of the United States—Rare-Earth Elements

Figure O3. Photograph of the Mountain Pass Mine in San 

Bernardino County, California, which was the only active producer 

of rare-earth elements (REEs) in the United States in 2013. The 

orebody is a tabular carbonatite intrusion and is thought to be the 

largest REE resource in the United States. Molycorp, Inc. (2012), 

which owns and operates the mine, reported proven and probable 

reserves totaling 16.7 million metric tons of carbonatite ore at 

an average grade of 7.98 percent rare-earth oxide (REO) using a 

cutoff grade of 5 percent REO. The mine was placed on care-andmaintenance status in 2015. Photograph by Bradley S. Van Gosen.

The orebody of the Mountain Pass Mine—the Sulphide Queen 

carbonatite—contains about 10 to 15 percent bastnaesite and 

lesser amounts of parisite (Castor and Nason, 2004).

In carbonatites, REE mineral assemblages can be 

categorized by mineral-forming process into the following 

three groups (Mariano, 1989; Chakhmouradian and Zaitsev, 

2012): (a) primary magmatic mineralization in carbonatites 

(Mountain Pass, Calif.), (b) hydrothermal-metasomatic 

mineralization (Bayan Obo, China), and (c) supergene 

mineralization (Mount Weld in Western Australia, Australia, 

and Araxá in the State of Minas Gerais, Brazil). The principal 

REE minerals in carbonatites are fluorocarbonates (bastnaesite, parisite, synchysite), hydrated carbonates (ancylite), 

and phosphates (monazite). Less abundant REE minerals in 

carbonatites include REE-bearing britholite, burbankite, and 

fluorapatite (table O2). Most carbonatites display a multiphase 

petrogenetic evolution, including hydrothermal overprinting of 

primary phases; thus, mineralogical and textural characteristics can vary and be complex.

Bastnaesite is typically the most significant source 

of REEs in carbonatite-related deposits, such as at 

Mountain Pass (California) and Bear Lodge (Wyoming) in 

the United States, and Bayan Obo (Nei Mongol Autonomous 

Region), Maoniuping (Sichuan Province), Daluxiang (Sichuan 

Province), and Weishan (Anhui Province) in China (fig. O4). 

Mariano (1989) notes that strong enrichments in the LREEs 

occur in carbonatite deposits in which the dominant ore 

mineral is bastnaesite. Mariano (1989) states that bastnaesite 

of hydrothermal origin typically occurs as fine-grained fibrous 

or platy masses in vugs, microfractures, and veinlets, which 

are commonly associated with, in general order of abundance, 

quartz, fluorite, strontianite, barite, and hematite. Ancylite is 

an indicator of secondary carbonate hydrothermal processes; 

it is found intergrown with strontianite in altered carbonatites 

(Mariano, 1989). The REE ore minerals in the intensely 

altered, upper oxidized zones of the Bear Lodge deposit are 

minerals of the bastnaesite group, plus cerianite (table O2) 

and locally abundant monazite (Dahlberg, 2014).

The vast majority of peralkaline REE deposits display 

complex petrogenetic evolution, which can result in variable 

mineralogy and intricate textures. Important REE-yttriumbearing minerals in peralkaline deposits can include, in 

general order of abundance, fluorapatite (fluorine-rich apatite), 

eudialyte, loparite, xenotime, monazite, gadolinite, bastnaesite, 

parisite, synchysite, kainosite, mosandrite, britholite, allanite, 

brannerite, thalenite, iimoriite, and fergusonite (table O2). 

In general, REE mineralization in peralkaline intrusions is 

typically more enriched in the HREEs than is the case with 

carbonatite-hosted deposits (fig. O5). Complex ore mineralogy 

with hydrothermal overprinting of primary magmatic phases 

is typified by the peralkaline dikes at Bokan Mountain on 

Prince of Wales Island in southeastern Alaska, in which 

more than 20 REE-yttrium-bearing minerals representing 

multiple generations have been identified (Barker and Van 

Gosen, 2012; Dostal and others, 2014). As another example, 

45 REE-bearing minerals have been identified in the Khibiny 

alkaline complex in Murmanskaya Oblast’, northwestern 

Russia (Chakhmouradian and Zaitsev, 2012).

In metamorphic deposits, common REE-yttrium-bearing 

minerals are the phosphate minerals xenotime and monazite. 

Examples in gneiss occur in the Music Valley area of the 

northern part of Joshua Tree National Park in Riverside 

County, Calif. (Evans, 1964; Long and others, 2010).

The primary REE-rich minerals within magmatic 

magnetite-hematite deposits are REE-bearing fluorapatite 

along with xenotime and monazite. Examples of massive 

magnetite-hematite deposits with REE-bearing fluorapatite, 

xenotime, and monazite as accessory minerals are the Pea 

Ridge iron deposit in Washington County, Mo., and iron 

deposits of the Mineville iron district in Essex County, New 

York (both districts are summarized in Long and others, 2010).

Nepheline syenite complexes in the Kola Peninsula 

region of Murmanskaya Oblast’ in far northwestern Russia 

have been mined for their enrichments in niobium and REEs; 

the REEs occur mainly in loparite (table O2), which is a 

mineral of the perovskite group. This loparite reportedly 

contains more than 30 weight percent REO and is strongly 

enriched in the LREEs compared with the HREEs (Mariano, 

1989; Chakhmouradian and Mitchell, 1998).

Deep weathering of REE-bearing alkaline igneous rocks 

or carbonatite in humid tropical climates with moderate to 

high rainfall can form thick laterite zones; these laterites can 

contain economic accumulations of residual REE minerals, 

depending on the underlying lithology. Examples of REE 

minerals that occur in the intensely weathered zone (laterite) 
Geology O9

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

EXPLANATION

Active mines

CarbonatitePeralkaline igneousHeavy-mineral sandsClays

Advanced projects

CarbonatitePeralkaline igneousHeavy-mineral sandsOther

Steenkampskraal

Mountain Pass

Round Top

Bear Lodge

Bokan Mountain

Elk Creek

Montviel Foxtrot

Kvanefjeld,Sørensen,Zone 3 Two Tom

Strange

Lake Eldor

Nechalacho

Hoidas Lake

Clay-Howells

Kipawa and Zeus

Sarfartoq

Norra Kärr

Kutessay II Lavergne-Springer

Mrima Hill Ngualla

Songwe Hill

Lofdal

Araxá

Glenover

Zandkopsdrift

Wigu Hill

Bayan Obo

Weishan

South China clays

Maoniuping

Daluxiang

Dong Pao

Odisha

Kerala

Mount Weld

Hastings

Nolans Bore

Milo

Plant inStepnogorsk

Eco Ridge

Dubbo Zirconia

Cummins Range

TRE

Lovozero

Charley Creek

Canakli

Buena Norte

mining district

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection;

REE, Figure 4.

 World Geodetic System 1984 datum

Figure O4. World map showing locations of active or recently active rare-earth-element (REE) mines and ongoing advanced exploration projects. The exploration projects are at the assessment stage, meaning that they are being evaluated to determine if they are economic to develop. REE resource estimates, which are usually based on extensive drilling programs, have been conducted on all the projects. Tables O3 and O4 contain additional information about the deposits shown on this map. Mountain Pass Mine in California (United States) was placed on care-and-maintenace status in 2015.
O10 Critical Mineral Resources of the United States—Rare-Earth Elements

include, in order of abundance, monazite, rhapdophane, 

crandallite-group minerals (florencite, gorceixite, goyazite), 

and cerianite; these are secondary rather than primary phases 

(Mariano, 1989). Examples of thick residual REE deposits 

(that is, more than 100 meters [m] thick) occur at the Araxá 

deposit, which overlies the Barreiro carbonatite complex in 

the State of Minas Gerais, Brazil; the Catalão I carbonatite 

complex in the State of Goiás, Brazil; and the Mount Weld 

carbonatite in Western Australia, Australia.

Monazite is the dominant REE mineral in sediments 

derived from the erosion of igneous or high-grade metamorphic rocks. These sediments are deposited in fluvial (stream 

and river) environments and coastal areas. Monazite is dense 

and resistant to chemical and physical weathering. For these 

reasons, more than any other REE mineral, monazite survives 

the arduous trip from a bedrock source to its deposition in a 

stream, river, or shore area. Monazite is the principal REE 

mineral recovered by placer mining operations, particularly 

in coastal deposits.

Deposit Types

Deposits of REEs occur in diverse, generally uncommon 

geologic settings. The discussion that follows does not attempt 

to encompass all the many subeconomic geologic occurrences 

in which REEs have been found (such as those described 

by Mariano and Mariano, 2012). As noted earlier, REEs are 

quite common, but not often in economic concentrations. 

The economic or potentially economic REE deposits occur 

primarily in the following geologic settings:

• Carbonatites

• Peralkaline igneous systems

• Magmatic magnetite-hematite bodies

• Iron oxide-copper-gold (IOCG) deposits

• Xenotime-monazite accumulations in mafic gneiss

• Ion-absorption clay deposits

• Monazite-xenotime-bearing placer deposits

Carbonatites. Carbonatites have been the world’s main 

source for the LREEs since the 1960s. In 2016, REEs were 

being produced from large carbonatite bodies mined in 

China (Bayan Obo, Maoniuping, Daluxiang, and Weishan 

deposits) (table O3; fig. O4). Carbonatite intrusions occur 

in a variety of forms and carbonate compositions. Forms 

include stocks, tabular bodies, dikes, irregular-shaped masses, 

and veins. Carbonatites vary in composition from calcic 

(calcite-dominant; originally called soviet), to dolomitic 

(dolomite-dominant; called beforsite), to iron-carbonate (rich 

in ankerite [Ca(Mg,Fe,2+Mn)(CO3)2] or siderite [Fe2+ CO3]). 

Calcite and dolomite can also occur in roughly equal proportions. Carbonate composition can vary complexly across an 

individual carbonatite intrusion, such as is displayed in the 

Sulphide Queen carbonatite at Mountain Pass (Olson and 

others, 1954; Castor, 2008). Cross-cutting relations, mineralogy, and geochemistry have shown that carbonatite masses 

and dikes of different mineralogy and chemistry in a single 

district can represent multiple, discrete episodes of carbonate 

magma intrusion (Yang and others, 2010).

As noted earlier, in carbonatite deposits, the primary REE 

ore minerals are bastnaesite, parisite, and synchysite, ancylite, 

and monazite. Primary gangue (non-ore) minerals are carbonate 

minerals (calcite, dolomite, ankerite, siderite), which may be 

accompanied by aegerine-augite, barite, fluorite, fluorapatite, 

hematite, magnetite, phlogopite, quartz, and (or) strontianite. 

Accessory minerals—monazite, perovskite, pyrochlore, and 

thorite—are the most common uranium- and thorium-rich phases.

Table O2. List of selected rare-earth-element-bearing and 

yttrium-bearing ore minerals.

[Source: Jones and others (1996, Appendix A)]

Mineral name1 Chemical formula

Allanite (REE,Ca,Y)2

(Al,Fe3+)3(SiO4)3(OH)

Ancylite Sr(REE)(CO3 )

2

(OH)•H2O

Bastnaesite (REE)(CO3)F

Brannerite (U,Ca,Y,REE)(Ti,Fe)2

O6

Britholite (REE,Ca,Th)5(SiO4,PO4)3(OH,F)

Burbankite (Na,Ca)3

(Sr,Ba,Ce)3(CO3)5

Cerianite-(Ce) (Ce4+,Th)O2

Eudialyte Na4(Ca,REE)2(Fe2+,Mn,Y)ZrSi8O22(OH,Cl)2

Fergusonite-(Y) YNbO4

Florencite (REE)Al3(PO4 )2(OH)6

Fluorapatite (Ca,REE,Na)5(PO4 )

3

(F,OH)

Gadolinite (REE,Y)2

Fe2+Be2Si2O10

Gorceixite (Ba,REE)Al3(PO4)2(OH5•H2O)

Goyazite (Sr,REE)Al3

(PO4 )2(OH5•H2O)

Iimoriite-(Y) Y2

SiO4CO3

Kainosite Ca2(Y,REE)2Si4O12CO3•H2O

Loparite-(Ce) (Na,Ce,La,Ca,Sr)(Ti,Nb)O3

Monazite (REE,Th)PO4

Mosandrite (Ca,Na,REE)12(Ti,Zr)2Si7O31H6F4

Parisite Ca(REE)2(CO3 )3F2

Rhapdophane (REE)PO4•H2O

Synchysite Ca(REE)(CO3 )2

F

Thalenite-(Y) Y3

Si3O10OH

Xenotime YPO4

1

A more-extensive list that includes 245 individual rare-earth-elementbearing minerals is provided in Jones and others (1996, appendix A).
Geology O11

The geology and mineralogy of the world’s largest REE 

deposit and leading producer of LREEs—the Bayan Obo 

iron-carbonatite deposit in Nei Mongol Autonomous Region, 

China—are described by Yuan and others (1992), Xu and 

others (2010), Yang and others (2010, 2011), Kynicky and 

others (2012), and Verplanck and others (2014). The Bayan 

Obo deposit is a complex irregular mass of mixed iron 

deposits and carbonatite. Two other REE producers in China 

are hosted by carbonatites—the Maoniuping deposit and the 

Daluxiang deposit, both in Sichuan Province. Other carbonatites in China are described by Yang and Woolley (2006), 

Xu and others (2008, 2010), and Kynicky and others (2012).

The largest REE deposit in the United States is the 

Sulphide Queen carbonatite in Mountain Pass, Calif.; this is 

the only known carbonatite where the REE ore minerals—

bastnaesite and parasite—are interpreted to have crystallized 

directly from the magma (Mariano, 1989; Castor, 2008). In 

other carbonatites, the REE-bearing minerals are interpreted 

to have formed by late-phase hydrothermal processes. Another 

type of REE enrichment in carbonatite systems—supergene 

processes—involves deep weathering of a carbonatite in tropical 

environments, which forms an REE-enriched laterite deposit. 

Examples of thick laterites with economic accumulations of 

residual REE minerals are the Araxá deposit that overlies the 

Barreiro carbonatite complex in Brazil; the Catalão I carbonatite 

La Ce Pr Nd Sm Eu Gd Tb Dy Ho Er Tm Yb Lu Y

Ratio of ore concentration to chondrite concentration

EXPLANATION

REE fig 5

Music Valley deposit—Monazite-xenotime-biotite gneiss

Bokan Mountain deposit—Peralkaline igneous dikes and veins

Thor Lake/Nechalacho deposit—Peralkaline igneous intrusion

Foxtrot project deposit—Mineralized, metamorphosed felsic volcanics

Bear Lodge deposit—Hydrothermally altered carbonatite

Mountain Pass deposit—Magmatic carbonatite

1,000,000

100,000

10,000

1,000

100

10

1

Element

Figure O5. Chondrite-normalized plot showing the rare-earth-element (REE) distribution in six 

different types of North American REE deposits. The example deposits are located in Music 

Valley, California; Bokan Mountain, Alaska; Nechalacho (Thor Lake), Northwest Territories, 

Canada; Foxtrot project, Newfoundland and Labrador, Canada; Bear Lodge, Wyoming; and 

Mountain Pass, Calif. Notice the higher enrichments in the heavy REEs associated with 

peralkaline deposits, represented by the Bokan Mountain and Thor Lake deposits. Data for 

the Bear Lodge, Bokan Mountain, Mountain Pass, and Music Valley deposits are from author 

samples; Foxtrot project data are from Srivastava and others (2012); Thor Lake data are from 

Avalon Rare Metals Inc. (2015). Elements: La, lanthanum; Ce, cerium; Pr, praseodymium; 

Nd, neodymium; Sm, samarium; Eu, europium; Gd, gadolinium; Tb, terbium; Dy, dysprosium; 

Ho, holmium; Er, erbium; Tm, thulium; Yb, ytterbium; Lu, lutetium; Y, yttrium
O12 Critical Mineral Resources of the United States—Rare-Earth Elements

Table O3. Active rare-earth mines, by deposit type.

[Mt, million metric tons; REE, rare-earth element; Y, yttrium; REO, rare-earth oxide; NA, not available; —; none reported]

Deposit Location

Reported 

resource 

(Mt)

Reported grade

(total REE+

Y oxide, in 

weight percent)

Comment Reference(s)

Carbonatites

Bayan Obo Nei Mongol 

Autonomous 

Region, China

800 6 Estimated resource in the total 

deposit, not subdivided

Berger and others 

(2009)

Daluxiang 

(Dalucao)

Sichuan, China 15.2 5.0 About 0.76 Mt (estimated) 

of REOs

Hou and others (2009)

Maoniuping Sichuan, China 50.2 2.89 REO content of reserves is estimated 

to be more than 1.45 Mt

Xu and others (2008); 

Hou and others 

(2009); Xie and 

others (2009)

Weishan Shandong, China — — Tonnage and grade information 

are not available

NA

Mountain Pass California, 

United States

16.7 7.98 Resource represents proven and 

probable reserves using a cutoff 

grade of 5 percent REO. Placed 

on care-and-maintenance status 

in 2015.

Molycorp, Inc. (2012)

Mount Weld Western Australia, 

Australia

23.9 7.9 Tonnage represents the estimated 

combined total mineral resource 

as of January 2012 for two 

deposits at Mount Weld—the 

Central Lanthanide deposit and 

the Duncan deposit

Lynas Corporation Ltd. 

(2012)

Peralkaline igneous

Karnasurt Mountain, 

Lovozero deposit

Northern region, 

Russia

— — Loparite concentrate contains 

30 to 35 percent REO

Zaitsev and Kogarko 

(2012)

Heavy-mineral sand deposits

Buena Norte 

mining district

East coast of Brazil — — Historic and active producer 

of REEs from monazite in 

coastal sands

NA

Ion-adsorption clay deposits

Dong Pao Mine Vietnam — — Mine is reportedly in a late stage 

of development. Laterite clays 

overlie syenite intrusions

NA

South China clay 

deposits 

Jiangxi, Hunan, Fujian, 

Guangdong, and 

Guangxi Provinces, 

southern China

— About 0.05 

to 0.4

Numerous small mines. Little ore 

information is available. Best 

source of data may be Chi and 

Tian (2008)

Clark and Zheng 

(1991); Bao and 

Zhao (2008); 

Chi and Tian (2008)
Geology O13

complex in Brazil; and the Mount Weld carbonatite in Australia. 

Laterite-hosted REE deposits are unknown in the United States.

Some carbonatite intrusions are important sources of 

niobium, which is hosted primarily in the mineral pyrochlore 

[(Na,Ca)2Nb2O6(OH,F)]. More than 90 percent of the world’s 

supply of niobium comes from laterite deposits that are several 

hundred meters in thickness and that overlie the Barreiro 

carbonatite complex near Araxá (Papp, 2011). The niobiumrich laterites at Araxá are now being evaluated for the recovery 

of REEs as a byproduct of niobium production. The largest 

identified niobium resource in the United States is the buried 

Elk Creek carbonatite in southeastern Nebraska (table O4; 

Xu, 1996; Carlson and Treves, 2005).

Peralkaline igneous systems. Peralkaline igneous systems 

are inherently enriched in REEs, in some instances hosting 

high-grade deposits. Peralkaline igneous complexes have 

been targeted by recent exploration (fig. O4) because of their 

enrichments in the HREEs (fig. O5).

Alkaline igneous rocks make up a large grouping of rock 

types that are generally deficient in silicon dioxide (SiO2) 

relative to Na2O, K2O, and calcium oxide (CaO). As noted 

earlier, peralkaline igneous rocks are a subclass of alkaline 

rocks, defined as those in which the molecular proportion of 

Al2O3<Na2O+K2O.

Peralkaline igneous systems enriched in REEs come in a 

variety of forms, such as the following:

• Complexes exhibiting vertical and lateral zonation: 

(a) Kipawa complex, Quebec, Canada (Saucier and 

others, 2012); (b) Strange Lake complex, Quebec and 

Labrador, Canada (Quest Rare Minerals Ltd., 2012); 

(c) Nechalacho (Thor Lake), Northwest Territories, 

Canada (Avalon Rare Metals Inc., 2015); (d) Kutessay II 

project, Chuy Oblusu, Kyrgyzstan (Stans Energy Corp., 

2015); and (e) Tantalus Rare Earths (TRE) project, 

northern Madagascar (Tantalus Rare Earths AG, 2013).

• Complexes exhibiting layering: (a) Lovozero complex, 

Murmanskaya Oblast’, Russia (Féménias and others, 

2005; Kogarko and others, 2006; Chakhmouradian and 

Zaitsev, 2012), and (b) Ilimaussaq complex (includes 

the Kvenfjeld, Sørensen, and Zone 3 deposits), 

Kommune Kujalleq, Greenland (Greenland Minerals 

and Energy Ltd., 2012).

• Dikes and veins associated with peralkaline igneous 

complexes: (a) Bokan Mountain, Alaska (Robinson 

and others, 2011); (b) Hastings project, Western 

Australia, Australia (Hastings Rare Metals Ltd., 2015); 

and (c) Hoidas Lake deposit, Saskatchewan, Canada 

(Star Minerals Group Ltd., 2014).

• Plutons, stocks, plugs, and other peralkaline intrusions: 

(a) Norra Kärr deposit, Jönköpings län, Sweden (Tasman 

Metals Ltd., 2012); (b) Dubbo Zirconia project, New 

South Wales, Australia (Alkane Resources Ltd., 2015); 

(c) Two Tom deposit, Labrador, Canada (Daigle, 2012); 

(d) at Ghurayyah in Tabuk Province, Saudi Arabia; and 

(e) at Khaldzan-Buregtey in Hovd Aymag, Mongolia.

Peralkaline igneous rock types in REE-bearing systems 

can vary considerably from deposit to deposit; thus, it is 

difficult to generalize which peralkaline lithologies are 

consistently associated with REE deposits. The advanced 

exploration projects (current as of 2012), which are listed 

in table O4 and the locations of which are shown in fig. O4, 

indicate significant REE mineralization of 0.3 to 2.6 percent 

total REO in alkali granites, nepheline syenite, syenite, 

and trachyte. Peralkaline igneous rocks are typically more 

enriched in REEs, especially HREEs, than their silica-rich 

igneous counterparts. Still, the processes of REE transport 

and enrichment in peralkaline systems remain poorly 

understood (Chakhmouradian and Zaitsev, 2012). The recent 

increase of exploration and drilling into peralkaline systems 

(table O4) is expected to provide a better characterization of 

these deposits.

The uncommon mineralogy of peralkaline-related REE 

deposits is summarized in the Mineralogy section of this 

chapter and described by Chakhmouradian and Zaitsev (2012). 

The mineralogies encountered in REE ores in peralkaline 

igneous systems present challenges in ore processing. Routine 

metallurgical methods for processing these types of ores do 

not exist; procedures are developed on a deposit-by-deposit 

basis. The ore processing must be efficient and economic, 

employing methods that effectively liberate the diverse REE 

minerals, then sequentially separate and concentrate the individual REEs. A peralkaline igneous deposit can contain more 

than 20 REE-bearing minerals. Additionally, the intergrown 

REE minerals can occur in a variety of compositions, such as 

carbonates, oxides, phosphates, and silicates.

Magmatic magnetite-hematite bodies. Magmatic 

magnetite-hematite bodies can contain REE-bearing minerals, 

with the potential to recover the REEs as a byproduct during 

iron mining. Examples are deposits in the Pea Ridge iron 

district in southeastern Missouri and the Mineville iron district 

in upstate New York (Long and others, 2010).

The Pea Ridge magnetite-hematite orebody is hosted by 

Mesoproterozoic volcanic rocks of the St. Francois terrane of 

southeastern Missouri. The magnetite-rich body is interpreted 

as a high-temperature, magmatic-hydrothermal deposit (Sidder 

and others, 1993) in ash-flow tuffs and lavas, which may have 

formed in the root of a volcanic caldera (Nuelle and others, 

1991). Four REE-bearing breccia pipes steeply crosscut the 

magnetite-hematite orebody and the altered rhyolite host rock. 

Exposed portions of the breccia pipes are as much as 60 m 

(197 feet [ft]) in length and as much as 15 m (49 ft) in width; 

the pipes extend below the mined levels to an undetermined 

depth (Seeger and others, 2001). Identified REE-bearing 

minerals in the breccia pipes are monazite, xenotime, and 

minor amounts of bastnaesite and britholite. According to 

Seeger and others (2001, p. 2), “total REO concentrations of 

grab samples [of breccia] range from about 2.5 to 19 weight 

percent.” Mine tailings on the surface contain additional 

lanthanide resources, primarily in fine-grained REE-bearing 

minerals, chiefly monazite and xenotime, which form inclusions within apatite (Vierrether and Cornell, 1993).
O14 Critical Mineral Resources of the United States—Rare-Earth Elements

Table O4. Advanced rare-earth-element (REE) exploration projects and the reported estimates of their REE resources, by deposit type.

—Continued

[Mt, million metric tons; t, metric ton; %, percent; TREO, total rare-earth oxide. Chemical elements: Ag, silver; Au, gold; Cu, copper; Dy, dysprosium; 

Mo, molybdenum; Nb, niobium; REE, rare-earth element; U, uranium; Zr, zirconium]

Deposit Location Reported resources Reference(s)

Carbonatites

Araxá Minas Gerais, 

Brazil

Measured and indicated resources=6.34 Mt at 5.01% TREO; 

inferred resources=21.94 Mt at 3.99% TREO

MBAC Fertilizer Corp. 

(2012)

Ashram project

(formerly Eldor 

project)

Quebec, Canada Measured resources=1.59 Mt at 1.77% TREO;

indicated resources=27.67 Mt at 1.9% TREO;

inferred resources=219.8 Mt at 1.88% TREO

Commerce Resources Corp. 

(2016)

Bear Lodge Wyoming,

United States

Measured and indicated resources = 16.3 Mt at 3.05% TREO;

inferred resources = 28.9 Mt at 2.58% TREO

Dahlberg (2014, p. 1-5, 1-8)

Cummins Range Western Australia, 

Australia

Inferred resources=4.9 Mt at 1.74% TREO Hatch (2015)

Elk Creek Nebraska, 

United States

Indicated resources = 80.5 Mt at 0.71% Nb oxide; 

inferred resources = 99.6 Mt at 0.56% Nb oxide; 

REE resources exist, but are not estimated

Pittuck and others (2015, 

p. 150)

Glenover project South Africa Indicated resources=16.78 Mt at 1.45% TREO;

inferred resources=12.14 Mt at 0.98% TREO

Galileo Resources PLC 

(2015)

Lavergne-Springer 

property

Ontario, Canada Indicated resources=4.2 Mt at 1.14% TREO;

inferred resources=12.7 Mt at 1.17% TREO

Hatch (2015)

Lofdal project Namibia Indicated resources=0.90 Mt at 0.62% TREO;

inferred resources=0.75 Mt at 0.56% TREO

Namibia Rare Earths Inc. 

(2012)

Montviel project Quebec, Canada Indicated resources=82.4 Mt at 1.51% TREO;

inferred resources=184.2 Mt at 1.43% TREO

GéoMégA Inc. (2015)

Mrima Hill Kenya Measured and indicated resources=48.7 Mt at 4.40% TREO; 

inferred resources=110.7 Mt at 3.61% TREO

Pacific Wildcat Resources 

Corp. (2013)

Ngualla Tanzania Measured resources=81 Mt at 2.66% TREO;

indicated resources=94 Mt at 2.02% TREO;

inferred resources=20 Mt at 1.83% TREO

Peak Resources Ltd. (2013)

Nolans Bore Northern Territory, 

Australia

Measured resources=4.3 Mt at 3.3% TREO; 

indicated resources=21 Mt at 2.6% TREO;

inferred resources=22 Mt at 2.4% TREO

Arafura Resources Ltd., 

(2014)

Sarfartoq Greenland Indicated resources=5.884 Mt at 1.77% TREO;

inferred resources=2.459 Mt at 1.59% TREO

Hudson Resources Inc. 

(2012)

Songwe Hill Malawi Indicated resources=6.2 Mt at 2.05% TREO;

inferred resources=5.1 Mt at 1.83% TREO

Mkango Resources Ltd. 

(2012)

Wigu Hill Tanzania Inferred resources=3.3 Mt at 2.6% TREO Montero Mining & 

Exploration Ltd. (2011)

Zandkopsdrift South Africa Measured resources=23.0 Mt at 2.07% TREO; 

indicated resources=22.7 Mt at 1.73% TREO;

inferred resources=1.1 Mt at 1.52% TREO

Frontier Rare Earths Ltd. 

(2014)

Peralkaline igneous

Bokan Mountain Alaska, 

United States

Indicated resources for the Dotson Ridge 

deposit=4.79 Mt at 0.602% TREO; 

inferred resources=1.05 Mt at 0.603% TREO

Ucore Rare Metals, Inc. 

(2015)

Clay-Howells project Ontario, Canada Inferred resources=8.5 Mt at 0.73% TREO Hatch (2015)

Dubbo Zirconia 

project

New South Wales, 

Australia

Measured resources=35.7 Mt at 0.75% TREO;

inferred resources=37.5 Mt at 0.75% TREO

Alkane Resources Ltd. 

(2015)

Foxtrot project Newfoundland and 

Labrador, Canada

Indicated resources=3.41 Mt at 1.07% TREO;

inferred resources=5.85 Mt at 0.96% TREO

Srivastava and others 

(2012)

Hastings project Western Australia, 

Australia

Indicated resources=27.0 Mt at 0.2103% TREO; 

additionally, 0.8913% Zr oxide and 0.3545% Nb oxide;

inferred resources=9.1 Mt at 0.21% TREO

Hastings Rare Metals Ltd. 

(2015)
Geology O15

Table O4. Advanced rare-earth-element (REE) exploration projects and the reported estimates of their REE resources, by deposit type.

—Continued

[Mt, million metric tons; t, metric ton; %, percent; TREO, total rare-earth oxide. Chemical elements: Ag, silver; Au, gold; Cu, copper; Dy, dysprosium; 

Mo, molybdenum; Nb, niobium; REE, rare-earth element; U, uranium; Zr, zirconium]

Deposit Location Reported resources Reference(s)

Peralkaline igneous—Continued

Hoidas Lake deposit

Ilimaussaq complex, 

Kvenfjeld deposit

Ilimaussaq complex, 

Sørensen deposit

Ilimaussaq complex, 

Zone 3 deposit

Kipawa

Kutessay II project

Nechalacho project 

(Thor Lake)

Norra Kärr

Round Top

Strange Lake

TRE project

Two Tom

Saskatchewan, 

Canada

Greenland

Greenland

Greenland

Quebec, Canada

Kyrgyzstan

Northwest 

Territories, 

Canada

Southern Sweden

Texas, United States

Quebec, and Newfoundland and 

Labrador, Canada

Northern 

Madagascar

Newfoundland and 

Labrador, Canada

Measured resources = 0.9638 Mt at 2.568% TREO;

indicated resources = 1.6 Mt at 2.349% TREO;

inferred resources = 0.2866 Mt at 2.139% TREO

Indicated resources = 437 Mt at 1.093% TREO;

inferred resources = 182 Mt at 0.9763% TREO

Inferred resources = 242 Mt at 1.1022% TREO

Inferred resources = 95 Mt at 1.1609% TREO

Indicated resources = 12.472 Mt at 0.512% TREO;

inferred resources = 3.842 Mt at 0.463% TREO

Measured resources = 13.548 Mt at 0.27% TREO;

indicated resources = 2.717 Mt at 0.238% TREO;

inferred resources = 1.746 Mt at 0.204% TREO

Measured resources = 12.56 Mt at 1.714% TREO;

indicated resources = 96.54 Mt at 1.57% TREO;

inferred resources = 160.25 Mt at 1.38% TREO

Inferred resources = 60.5 Mt at 0.54% TREO

Measured and indicated resources = 480 Mt at 0.063% TREO; 

inferred resources = 342 Mt at 0.063% TREO

Indicated resources = 278.12 Mt at 0.93% TREO;

inferred resources = 214.4 Mt at 0.85% TREO

Deposits include upper weathered bedrock of peralkaline 

intrusive complex and overlying argillaceous laterites 

(saprolite); inferred resources = 130 Mt at 0.08% TREO

Inferred resources = 41 Mt at 1.18% TREO

Star Minerals Group Ltd. 

(2014)

Greenland Minerals and 

Energy Ltd. (2012)

Greenland Minerals and 

Energy Ltd. (2012)

Greenland Minerals and 

Energy Ltd. (2012)

Saucier and others (2012, 

p. 75)

Stans Energy Corp. (2015)

Avalon Rare Metals Inc. 

(2015)

Tasman Metals Ltd. (2012)

Hulse and others (2014, p. 

152)

Quest Rare Minerals Ltd. 

(2012)

Tantalus Rare Earths AG 

(2013)

Daigle (2012)

Polymetallic deposits

Eco Ridge Mine

Milo project

Plant in Stepnogorsk

Steenkampskraal

Ontario, Canada

Queensland, 

Australia

Northern 

Kazakhstan

South Africa

Indicated resources = 22.743 Mt at 0.1606% TREO;

inferred resources = 36.560 Mt at 0.1554% TREO

Inferred resources = 187 Mt at 0.06% TREO;

deposit contains REEs, Cu, Mo, Au, Ag, and U

Resources and grade are not reported. Plant is being 

designed to extract Dy from uranium tailings

Measured resources = 85,000 t at 19.5% TREO;

indicated resources = 474,000 t at 14.1% TREO;

inferred resources = 60,000 t at 10.5% TREO

Pele Mountain Resources 

(2013)

GBM Resources Ltd. 

(2012)

Blank (2012)

Great Western Minerals 

Group Ltd. (2014)

Heavy-mineral sand deposits

Charley Creek 

deposits

Canakli deposit

Kerala deposits

Northern Territory, 

Australia

Turkey

India

Alluvial placers with U resources along with monazite and xenotime; inferred total resources of 805.3 Mt at 0.03% TREO

Inferred resource = 494 Mt at 0.07% TREO

Total resources of 1.51 Mt of monazite reported in heavymineral sand deposits in the State of Kerala; resource grade 

is not reported

Hatch (2015); Pancontinental Uranium Corp. (2012)

AMR Mineral Metal Inc. 

(2015)

Government of India, 

Department of Atomic 

Energy (2011)
O16 Critical Mineral Resources of the United States—Rare-Earth Elements

Thorium and REEs reside within apatite in iron bodies 

once mined in the Mineville iron district, New York, which is 

located in the northeastern part of the Adirondack Mountains. 

The orebodies are magnetite deposits that are intricately folded 

and faulted within a complex suite of Precambrian metamorphic and igneous rocks. The host rocks have both mafic and 

felsic compositions that include, in general order of abundance, gabbro, diorite, augite syenites, and granite (Staatz and 

others, 1980). The iron deposits are mainly magnetite, martite, 

and apatite, with accessory augite, hornblende, albite, quartz, 

pyrite, and tourmaline (McKeown and Klemic, 1956). Iron 

deposits in the Mineville–Port Henry area contain apatite and 

are likewise rich in phosphorous, REEs, and thorium because 

these elements are concentrated within the apatite grains. 

In addition, the magnetite (the primary iron ore mineral) is 

intergrown with 1- to 3-millimeter (0.04- to 0.12-inch)-long 

rice-shaped grains of apatite. Massive piles of tailings plus 

unmined parts of magnetite bodies in the Mineville district 

each contain REE-bearing apatite-rich rock. McKeown 

and Klemic (1956) reported an average REO content of 

11.14 percent in 14 samples of apatite separated from the 

Old Bed, Joker, and Smith orebodies in this district.

Iron oxide-copper-gold (IOCG) deposits. IOCG deposits 

are closely related to magmatic iron deposits, described just 

above, and one could argue that the Pea Ridge and Mineville 

iron deposits represent variations within the IOCG deposit 

group. Strictly, IOCG deposits are defined as magmatichydrothermal iron deposits that host economic concentrations 

of copper and gold. This group of deposits was first described 

and defined after the 1975 discovery of the giant Olympic 

Dam copper-uranium-gold deposit in South Australia, 

Australia, which is one of the largest ore deposits in the world. 

An overview of the IOCG deposit type along with examples 

of significant IOCG deposits worldwide is provided by Groves 

and others (2010).The IOCG deposits can be very large and 

surrounded by alteration zones that can extend for kilometers 

in width. As at Olympic Dam, IOCG deposits can host REE 

mineralization, occurring both in iron oxides and in small 

carbonatites within the intrusive complex. Although an REE 

resource of substantial size has been identified at Olympic 

Dam, the mine operator has not yet recovered REEs.

Xenotime-monazite accumulations in mafic gneiss.

Xenotime-monazite accumulations in mafic gneiss have not 

been historically important sources of REEs. Because deposits 

of this type tend to be preferentially enriched in the HREEs, 

however, they may be more economic in the future.

Example deposits occur in the Music Valley area of 

the northern part of Joshua Tree National Park in southern 

California. Pod-like xenotime deposits of the Music Valley 

area lie within the Pinto Gneiss of probable Precambrian 

age. The Pinto Gneiss consists of roughly equal amounts of 

quartz and plagioclase feldspar, and it averages approximately 

35 percent biotite (Evans, 1964). Accessory minerals present 

in trace amounts in the gneiss include, in general order of 

abundance, sericite, apatite, magnetite, zircon, and sphene, and 

locally monazite, actinolite, orthoclase, microcline, perthite, 

and muscovite. Biotite-rich zones in the gneiss can contain 

abundant amounts of orange xenotime grains; xenotime 

commonly forms 10 to 15 percent of the biotite zones and 

locally as much as 35 percent (Evans, 1964). Owing to the 

xenotime and accompanying monazite, the biotite-rich zones 

can contain more than 8 percent yttrium (Evans, 1964) and 

elevated percentages of HREEs (fig. O5).

Ion-adsorption clay deposits. Ion-adsorption clay 

deposits in southern China are the world’s primary sources 

of the HREEs. This deposit type is often informally referred 

to as “south China clays” (fig. O4). Despite their economic 

importance, little geologic research has been published that 

describes the south China clay deposits in detail; a study by 

Bao and Zhao (2008) is an exception, as well as discussions in 

Chi and Tian (2008) and Kynicky and others (2012).

Rare-earth-element-bearing ion-adsorption clay deposits 

form in tropical regions with moderate to high rainfall through 

the following general processes:

1. The REEs are leached by groundwater from granites 

(the bedrock),

2. Thick zones of laterite soils develop above the granites; 

this intensely weathered zone contains an abundance 

of clays, and

3. The mobilized REEs become weakly fixed (ionadsorption) onto the clays (kaolinite and halloysite) 

in the soils.

In deposits of this type in southern China, REE concentrations range from approximately 300 parts per million 

(ppm) to 2,000 ppm and appear to vary with the parent rocks 

(Bao and Zhao, 2008). These modest REE concentrations are 

economic for the south China deposits because (a) the REEs 

can be easily extracted from the clays with weak acids; (b) the 

deposits are often enriched in the high-value HREEs; (c) the 

area has low labor costs; and (d) there has been a localized 

lack of environmental protection (Liang and Others, 2014).

The highest REE concentrations usually occur in the 

B horizon (the subsoil) of the weathered zone. In addition 

to the highest REE content, the B horizon also contains the 

highest clay content. The overlying A horizon (the uppermost 

zone) is richest in organic matter, and contains much clay. 

The underlying C horizon consists of clays and pieces of the 

weathered granite.

Sanematsu and others (2009) conducted geochemical 

and sequential leaching experiments of ion-adsorption clays 

from deposits in Laos. They concluded that “the enrichment 

of REE[s] is attributed to the occurrence of ion-exchangeable 

clay minerals and REE phosphates in the weathered crusts 

and that HREE[s] are [more] selectively adsorbed on the clay 

minerals than are LREE[s] by weathering” (Sanematsu and 

others, 2009, p. 527). Laterite clay deposits of the Longnan 

district in Jiangxi Province in southern China produce HREErich material, whereas clay ores from the Xunwu district, also 

in Jiangxi Province, produce material that is rich in LREEs 

(O’Driscoll, 2003).
Resources and Production O17

Laterite clay deposits that are generally similar to the 

deposits in southern China are under development in northern 

Vietnam at the Dong Pao Mine in Lai Chau Province (fig. O4). 

In Madagascar, laterite clays formed by tropical weathering 

above REE-enriched dikes and sills are being explored at the 

TRE project (table O4; fig. O4; Tantalus Rare Earths AG, 2013).

An announced pilot project in Jamaica is intended to 

test the recovery of REEs from tailings of red mud produced 

from bauxite mining. The red mud reportedly contains 

0.23 to 0.38 weight percent REEs (Wagh and Pinnock, 1987) 

and has higher HREE concentrations than do other representative Chinese clay samples (Mariano and Mariano, 2012).

Monazite-xenotime-bearing placer deposits. Monazitexenotime-bearing placer deposits were important sources of 

REEs up to the mid-1960s, but they decreased in prominence 

owing to the discovery of the Mountain Pass deposit in 

California. Placer deposits may be significant sources of REEs 

again in the near future. Presently, monazite is extracted from 

coastal sands at (a) the Buena Norte mining district on the 

east coast of Brazil (table O3); (b) along the southwestern 

coast of India in the State of Kerala (table O4); and (c) along 

the coastline in the State of Odisha in northeastern India 

(fig. O4; Government of India, Department of Atomic Energy, 

2011). The monazite is recovered as a byproduct during the 

extraction of ilmenite (FeTiO3 ), leucoxene (altered ilmenite), 

and rutile (TiO2) from beach sands. The ilmenite, leucoxene, 

and rutile—the principal products of value—are chemically 

processed to remove titanium dioxide, which is used primarily 

as a pigment in paints.

Two broad settings for monazite-bearing placer deposits 

are (a) deposits of streams and rivers (fluvial deposits), and 

(b) coastal and nearshore deposits of sand and silt. In both 

deposit types, heavy dense minerals (including monazite) are 

physically sorted and deposited together by the combined actions 

of water movement and gravity. The resulting deposits of sediments, which are composed of sand and silt with thin layers of 

heavy minerals, are often referred to as “heavy-mineral sands.”

“Heavy minerals” are commonly defined as dense 

minerals with a specific gravity greater than about 2.9. These 

minerals are generally resistant to chemical weathering and 

physical degradation and thus survive well in fluvial environments. Heavy minerals are eroded from bedrock sources, 

carried by streams or rivers until they are deposited in the 

stream channel system or into a coastal plain setting. In the 

stream, the heavy minerals are carried within a slurry mixture 

of sand and silt. Because of their high density, the heavy 

minerals can be well sorted by fluvial processes and preferentially deposited into layers through the combined effects of 

gravity and decreases in stream energy. The Charley Creek 

deposits in Northern Territory, Australia (table O4; fig. O4) 

are an example of REE-bearing alluvial placer deposits. 

At Charley Creek, the REEs occur in the heavy minerals 

monazite ((REE,Th)PO4 ) and xenotime (YPO4 ).

Carrying heavy minerals, sediments are brought to 

coastal areas by streams and rivers. These sediments are 

reworked in nearshore (coastal) areas by the actions of waves 

and tides, which sort the heavy minerals into discreet layers. In 

these deposits, dark layers of heavy minerals are often referred 

to as “black sands.” A notable example of such deposits is the 

detrital monazite deposits in the coastal regions of southern 

India, which are found in piedmont lakes, shallow seas, sand 

bars across the mouths of rivers, deltas, beaches, and sand 

dunes behind the beaches (Bhola and others, 1958).

In addition to modern coastal regions, heavy-mineral 

deposits occur in ancient coastal plain environments, which are 

other potential sources of monazite. Carpenter and Carpenter 

(1991) describe heavy-mineral deposits formed by Pliocene 

transgressive-regressive events in the Upper Coastal Plain 

region of eastern North Carolina and eastern Virginia. In 2015, 

Iluka Resources Ltd. was mining and processing Pliocene-age 

heavy-mineral sands in the coastal plain of Virginia. Iluka has 

also worked similar ancient nearshore heavy-mineral deposits 

in Australia since the 1950s. Iluka’s heavy-mineral sands 

operations target the titanium minerals (ilmenite, leucoxene, 

rutile) and zircon; they do not recover the monazite.

Heavy minerals in the fluvial and coastal deposit 

types may include, in order of general abundance, ilmenite, 

leucoxene, rutile, magnetite, zircon, staurolite, kyanite, sillimanite, tourmaline, garnet, epidote, hornblende, spinel, iron 

oxides, sulfides, monazite, and xenotime. The typical minerals 

of value are ilmenite, leucoxene, and rutile for their titanium 

and titanium dioxide; zircon for its refractory uses (such as 

specialty ceramics); and in some cases, monazite, for its REEs 

and thorium. The heavy minerals as a suite typically make up 

no more than 15 percent of a deposit, and usually much less; 

quartz and clay minerals form the bulk of the sediment. As an 

example, in the Pliocene-age Upper Coastal Plain deposits of 

North Carolina and Virginia, the heavy-mineral suite consists 

of, on average, about 9 percent of the sediments and monazite 

occurs in trace amounts (0 to 0.9 weight percent of the heavy 

minerals) (Berquist, 1987; Carpenter and Carpenter, 1991).

The occurrence of REEs in fluvial placer or nearshore 

placer deposits depends on the presence of monazite and (or) 

xenotime in the bedrock sources upstream. Thus, exploration 

for monazite-bearing placer deposits must consider the mineralogy of the bedrocks that are the source of the sediments.

Resources and Production

Identified Resources

Tables O3 and O4 summarize the estimated REE 

resources reported for the world’s most extensively explored 

deposits. Exploration for additional REE deposits is ongoing; 

therefore, the world’s identified REE resources very likely will 

increase. The resource estimates shown in tables O3 and O4 

indicate that large resources of REEs exist. If developed, these 

deposits may supply the world’s REE needs for many decades. 

The process of bringing a particular deposit from the phase of 

orebody delineation to eventual mining and REE processing is a 

complex, multistage, multiyear process, however, as explained 
O18 Critical Mineral Resources of the United States—Rare-Earth Elements

in Long and others (2010). The pace or ultimate success of the 

REE exploration projects in table O4 is difficult to predict.

Berger and others (2009) compiled published tonnage 

and grade data for 58 REE- and niobium-bearing carbonatite 

deposits worldwide; the carbonatites included in the dataset 

are mostly well explored and either partially mined or known 

to contain these resources. For their carbonatite dataset, they 

calculated a median tonnage of 49 million metric tons and median 

grades of 0.61 percent total REO and 0.23 percent niobium oxide.

An earlier inventory of global REE resources (Jackson 

and Christiansen, 1993) reported that China hosts the largest 

resources of REEs, followed by Namibia, the United States, 

Australia, and India. Jackson and Christiansen (1993, p. 43) 

stated that “Globally, economically recoverable REO ore 

resources, based on 106 properties, amount to 20.6 billion 

metric tons, of which 85 percent is in placers and 15 percent 

is in hard-rock deposits.” The REOs in the ore deposits occur 

dominantly within the hard-rock deposits, however; according 

to their study, 7 percent of the REOs occur in placers and 

93 percent are in hard-rock deposits.

China’s Bayan Obo iron-carbonatite deposit in Nei 

Mongol Autonomous Region has been the world’s principal 

source of LREEs since the mid to late 1980s (Long and others, 

2010; U.S. Department of Energy, 2010, 2011). Bayan Obo is 

thought to be the world’s largest REE deposit—it is estimated 

to contain about 800 million metric tons of ore material at a 

grade of 6 percent REO content (Berger and others, 2009).

Mining by Molycorp, Inc. in Mountain Pass, Calif., has 

exploited the Sulphide Queen carbonatite (Olson and others, 

1954; Castor, 2008), which is the largest high-grade, LREE 

deposit in the United States (figs. O3, O4, and O5). This 

carbonatite reportedly contains 16.7 million metric tons of 

proven and probable reserves at a grade of 7.98 percent total 

REO (Molycorp, Inc., 2012).

India reports large resources of REEs, mostly in monazite 

within coastal sediments along its southern coastlines. According 

to India’s Atomic Mineral Directorate for Exploration & 

Research, “the reserves of monazite in India is about 10.70 

million tonnes which translates to approx. 5 million tonnes of 

rare earth oxide” (Government of India, Department of Atomic 

Energy, 2011, p. 2). Also, the Directorate estimates that 

about 2,000 metric tons of xenotime-bearing heavy-mineral 

concentrate, which contains about 2 percent xenotime, occurs 

in riverine placer deposits in the States of Chhattīsgarh and 

Jharkhand. With national sponsorship, the coastal placers 

are mined for their titanium minerals (rutile, ilmenite) and 

monazite. IREL, which is a public sector entity under the 

Department of Atomic Energy, is constructing a monazite 

processing plant in the State of Odisha (Government of India, 

Department of Atomic Energy, 2011).

Undiscovered Resources

The increase in REE prices has led to an increase in 

exploration activity worldwide; as a result, many more 

REE resources have been discovered or, in many cases, 

reinvestigated and better defined. Because REE enrichments 

can be found in a wide variety of rock types and environments, the probability for the occurrence of substantial 

undiscovered resources is high. With engineering advances 

in ore processing, additional deposits may become economic. 

The following are examples of some of the deposit types that 

could be potential future sources of REEs.

• Although the ion-adsorption clay deposits of southern 

China now serve as the world’s primary source of 

the HREEs, these deposits appear not to be unique to 

southern China. Clay deposits with similar geologic 

settings and characteristics have been explored 

elsewhere, although none has yet been brought 

into production on a significant scale. The Dong 

Pao deposit in northern Vietnam, for example, is 

composed of laterite clays that overlie syenite intrusions. The Dong Pao Mine reportedly is in a late stage 

of development.

• Elevated concentrations of REEs in sea-floor sediments sampled by deepwater drill coring at numerous 

sites on the floors of the eastern South Pacific Ocean 

and central North Pacific Ocean were reported by 

Japanese scientists in 2011 (Kato and others, 2011). 

The scientists collected more than 2,000 sea-floor sediments, sampled from about 1 m into the sea floor, from 

78 sites across the Pacific Ocean. As one example of 

their results, Kato and others (2011) reported that muds 

in the eastern South Pacific Ocean had REE contents 

comparable to or higher than those of the south China 

ion-adsorption clay deposits; also, that the HREE 

values were nearly twice as much as the China clays.

• Some modern and ancient coastal deposits can host 

large undeveloped, poorly studied deposits of monazite. Coastal deposits offer several advantages, such 

as that (a) these deposits are unconsolidated or weakly 

consolidated and thus easy to excavate; (b) the deposits 

are relatively easy to process in order to separate out 

the heavy minerals; and (c) multiple minerals can be 

extracted as the primary products of value, such as 

ilmenite and rutile for titanium and titanium dioxide, 

zircon for refractories, and garnet for abrasives. As one 

example, large undeveloped deposits of coastal sediments, ranging from Cretaceous to recent in age, occur 

from eastern Virginia to Florida in the Southeastern 

United States. Many sand and silt units in this coastal 

plain region contain monazite (Staatz and others, 1980; 

Carpenter and Carpenter, 1991).

Exploration for New Deposits

The search for REEs covers a wide variety of geologic 

provinces; noteworthy deposits of REE-bearing minerals have 

been found in alkaline igneous rocks, carbonatites, layers in 
Exploration for New Deposits O19

mafic metamorphic rocks, laterite clay deposits developed 

upon weathered granitic plutons, and coastal and fluvial 

sediments. Some of these types of settings are described 

briefly below.

Carbonatites and peralkaline igneous complexes. 

According to Woolley and Kjarsgaard (2008), there are 

527 known carbonatites in the world, and about 40 percent are 

located in the African continent. The majority of the known 

carbonatites are located in relatively tectonically stable regions 

within crustal plates, and they are often positioned on deepseated structural zones of regional extension (intra-continental 

rift zones). Less commonly, some carbonatites are found 

near plate margins, and their formation is possibly linked to 

processes of crustal plate separation or orogeny.

By definition, carbonatites are an igneous rock composed 

of at least 50 percent carbonate minerals, which sets them 

apart from other intrusions in igneous terrains. The gangue 

carbonate minerals, such as calcite and dolomite, can be 

identified with weak acid, as with any carbonate rock. Most 

carbonate minerals in carbonatites are generally subdued in 

outcrop, and the principal REE ore minerals are usually not 

colorful. In regions with deep, intense weathering, such as 

laterite weathering of the upper parts of carbonatites in Brazil 

and Australia, the laterites contain resistate ore minerals 

(minerals that are resistant to weathering), such as pyrochlore 

and monazite.

Fenite is a quartz-alkaline feldspar-rich rock that has 

been altered by alkaline (sodium- and [or] potassium-rich) 

hydrothermal fluids at the contact of a carbonatite intrusion 

complex. Fenitization is the alkaline mineralization formed by 

alkali metasomatism during the emplacement of an alkaline 

intrusion or carbonatite. This process can form alteration 

aureoles (fenite) of variable widths and intensity that are 

developed in the intrusion and the surrounding country rock. 

The presence of a fenite halo indicates that alkali hydrothermal processes took place there and is generally a favorable 

indication of REE enrichment.

Brief descriptions of alkaline provinces and individual 

alkaline complexes are available for North America and 

South America (Woolley, 1987), the former Soviet Union 

(Kogarko and others, 1995), and Africa (Woolley, 2001). 

These publications are useful references for those conducting 

REE exploration in these regions.

In an alkaline igneous complex, rock types generally 

progress from oldest to youngest as ultramafic to mafic 

lithologies (pyroxenite, biotite-shonkinite, peridotite, and 

jacupirangite, as examples), to felsic compositions (syenite, 

alkaline granites, ijolite, trachyte, and phonolite), to 

carbonatites. Coincidences of these alkaline rock types can 

indicate evolved alkaline systems. Carbonatite formation 

is a late (usually last) phase of the alkaline complex, and 

the carbonatite is often centrally located in the complex. 

Carbonatite dikes, if present, are also late-phase formations, 

and they may extend for kilometers outward from the 

complex. Dikes or veins of carbonatite can indicate that a 

larger carbonatite mass lies below or in the area.

In peralkaline intrusions, some REE deposits contain 

relatively rare but useful indicator REEs and alkali minerals, 

such as acmite (brown/green), alkali amphiboles (blue/black), 

eudialyte (pink/red), gadolinite (green/brown-black), rinkolite 

(red brown/yellow brown), and sodalite (dark blue).

Because of the presence of associated thorium-uraniumbearing minerals in alkaline complexes and carbonatites, 

radioactivity is a common characteristic of REE mineralization. Most of the carbonatites and alkaline-intrusion-hosted 

REE deposits in the United States were discovered by uranium 

prospectors using handheld radiation detectors, such as Geiger 

counters and scintillation counters (scintillometers). In most 

of these deposit types, the radiation is primarily a result of the 

presence of thorium.

The most diagnostic geochemical elements for 

carbonatites and peralkaline-hosted REE deposits are the 

high-field-strength elements, which include hafnium, niobium, 

REEs+yttrium, scandium, tantalum, thorium, titanium, 

uranium, and zirconium. Large-ion lithophile elements—

barium, cesium, potassium, rubidium, and strontium—

are also enriched in carbonatites and alkaline intrusions. 

Associated elements can include chlorine, fluorine, 

manganese, and phosphorus.

Geophysical methods can be used to detect and map 

an ore-bearing mass that may not be exposed. Ground and 

airborne geophysical techniques have been used to identify 

and define the size of buried carbonatites and mineralized 

alkaline intrusions. Geophysical data can be collected and 

analyzed at a wide range of scales, thereby contributing to 

exploration at a variety of scales, from regional features to 

detailed characterization of an individual deposit.

Modern airborne geophysical technologies offer the 

capability to collect gamma-ray, magnetic, and gravity data 

concurrently. Electromagnetic data can also be acquired 

this way. The combination and integration of these methods 

provides significant insight into a number of physical 

properties, which, in combination, can significantly enhance 

an exploration program for intrusion-hosted REE deposits. 

Airborne geophysical surveys that are flown most often over 

prospective REE deposits are magnetic, gamma-ray, and, 

most recently, gravity and gravity gradiometry. The choice of 

airborne geophysical tool(s) and platform (helicopter or fixed 

wing) used to explore for an intrusion-hosted REE deposit 

depends on a number of factors, such as accessibility to the 

deposit, ruggedness of the terrain, and whether the deposit 

is entirely or partly exposed or is concealed by rock, glacial 

deposits, vegetation, or water and (or) ice.

Gravity and magnetic surveying can be especially useful 

to map partially or entirely concealed carbonatites and alkaline 

intrusions. Although these methods do not provide direct 

detection of REE mineralization, they can provide necessary 

geologic controls on the depth, geometry, and extent of the 

host intrusion(s). The success of these methods, as with all 

geophysical mapping tools, depends on there being a physical 

property contrast (density and magnetization) of the intrusion 

with the surrounding rocks.
O20 Critical Mineral Resources of the United States—Rare-Earth Elements

Magmatic magnetite-hematite deposits and iron oxidecopper-gold (IOCG) deposits. As noted earlier, IOCG deposits 

are typically giant deposits with broad alteration zones that 

can extend for several kilometers outward from the main 

deposit (Groves and others, 2010). Iron-dominated deposits, 

typified by magmatic iron and IOCG deposits, produce 

substantial magnetic and gravity anomalies, which are usually 

clearly evident in geophysical surveys at regional and local 

scales. The techniques traditionally employed in the exploration for precious and base-metal deposits can be likewise used 

to discover magmatic iron and IOCG deposits.

Xenotime-monazite deposits in metamorphic rocks. 

High-grade concentrations of xenotime and monazite have 

been found locally in some mafic gneiss, paragneiss, and 

migmatites (Mariano and Mariano, 2012). Thus far, none of 

the discovered occurrences of this type have been deemed 

large enough to be economic; however, their inherent enrichment in HREEs continues to attract interest.

Local exploration for these deposits can utilize the 

durability of xenotime and monazite; they are heavy dense 

minerals that are resistant to mechanical weathering and, as a 

result, they accumulate with the heavy-mineral suite in stream 

sediments. Monazite- and xenotime-bearing sediments are 

typically elevated in the LREEs, yttrium, and thorium.

Radioactivity owing to monazite may be detectable using 

high-resolution airborne gamma-ray surveys with closely 

spaced flight lines if the host rocks are exposed at or very near 

the surface. Ground surveys with a handheld radiation meter 

can identify monazite-rich zones.

Ion-adsorption clay deposits. On the most basic level, 

favorable regions for ion-adsorption clay deposits are those 

where tropical weathering horizons (laterites) formed atop 

alkaline igneous masses. In southern China, underlying 

bedrock includes altered “two mica” granites. In addition to 

the modern tropics, paleo-laterites coincident with alkaline 

intrusive masses can also be considered.

In many ways, discovery of these types of REE deposits 

have proven the most elusive thus far, outside of southern 

China. These clay deposits lack radioactivity (Bao and Zhao, 

2008) and have relatively low concentrations of REEs. 

Sampling studies of prospective laterites typically extract 

subsurface samples of the weathered zone at selected soil 

horizons; the samples are analyzed for their REE content 

and REE leachability.

Monazite-xenotime-bearing fluvial placer deposits.

Many of the REEs and associated high-field-strength elements 

occur in heavy, dense minerals that are resistant to weathering 

(resistates); thus, heavy-mineral concentrates from stream 

sediments are a useful sampling technique for these deposit 

types. Resistate indicator minerals in the heavy concentrate 

portion of stream sediments can include, in order of general 

abundance, monazite, pyrochlore, xenotime, allanite, zircon, 

kainosite, thorite and uranothorianite, and euxenite (Barker 

and Van Gosen, 2012). Using airborne gamma-ray survey data, 

radioactivity caused by the presence of thorium (mainly in 

monazite) may outline fluvial networks that contain monazite.

Monazite-bearing coastal plain deposits. Deposits of 

sand and silt that contain monazite occur in many coastal 

environments. These sediments are carried by rivers and 

streams to coastal settings, including the shore. The sediments 

originate from the erosion of metamorphic and (or) igneous 

bedrock located up gradient of the coast. The monazite in 

these sediments occurs in close association with other heavy 

minerals, which are most commonly ilmenite, leucoxene, rutile, 

and zircon. These heavy minerals could thus be considered 

indicator minerals in these types of deposits. Economic “black 

sand” deposits in coastal settings (modern or paleo-settings) 

typically contain 5 percent or greater heavy-mineral content.

Exploration for coastal heavy-mineral sand deposits 

involves primarily any or all of the following:

• Locating ancient to modern coastal sedimentary 

deposits that were and (or) are sourced by terrains 

containing abundant metamorphic rocks and (or) 

igneous rocks. Favorable target areas are those that 

experienced transgressive-regressive marine cycles, 

as well as areas at the intersection of ancient 

shorelines and major paleorivers.

• Sampling sediments at the surface or (more thoroughly) vertically by using truck-mounted power 

auger drills.

• Simple panning on site to estimate the approximate 

heavy-mineral content of a sample. A more-exacting 

determination of heavy-mineral content is made 

by using multiple steps, including heavy liquid 

separation, grain counts of heavy-mineral concentrates, 

and thorough analyses of heavy-mineral concentrates 

using spirals, solutions, magnetic separators, and 

electrostatic separators, as described by Carpenter 

and Carpenter (1991).

Environmental Considerations

Rare-earth elements are currently produced from just a 

few mines globally, and, as a result, there are no abandoned 

REE mines available for study. Previous production from the 

Mountain Pass Mine in California started before the era of 

major environmental regulatory oversight; as a result, much 

of the environmental information associated with new mines 

is lacking. Further, the Bayan Obo Mine in China, which is 

the main current source of REEs, also has limited environmental information published. Two advanced exploration 

projects—the Nechalacho alkaline-intrusion-related deposit 

(near Thor Lake, Northwest Territories, Canada) and the Bear 

Lodge carbonatite deposit (Bear Lodge Mountains, northeastern Wyoming)—represent new sources of environmental 

data as a result of ongoing activities at these sites. At present, 

however, information regarding the environmental aspects of 

REE mining is limited. Toxicological data about the effects of 

REEs on aquatic, animal, or human health are also limited.
Environmental Considerations O21

Sources and Fate in the Environment

The concentrations of REEs in environmental media 

are influenced by their normal tendency to occur in diluted 

concentrations in crustal rocks (rather than in concentrated 

deposits) and their limited solubility in most groundwaters 

and surface waters. A significant amount of information 

is available on their natural abundances and distributions 

in environmental media, such as seawater (Elderfield and 

others, 1988), river water (Goldstein and Jacobsen, 1988; 

Sholkovitz, 1995), groundwater (de Boer and others, 1996), 

soil, and plants (Tyler, 2004), because of their value in tracing 

hydrochemical and geochemical processes. The average upper 

crustal abundance of all REEs combined (lanthanum through 

lutetium) is 146.4 ppm; the average abundance of yttrium 

is 22 ppm. The mean REE concentration of soils is similar 

to the average upper crustal abundance of these elements 

(Smith and others, 2013; Tyler, 2004). In humid settings with 

extreme weathering, however, REE concentrations in soil 

tend to be higher than those in the underlying rock because of 

leaching of major elements. Plant uptake of REEs is generally 

low because of the low solubility of REEs, and uptake does 

correlate with soil acidity (Tyler, 2004). The solubility of 

REEs in typical surface water or groundwater compositions 

is low (Wood, 1990). In seawater, the average dissolved 

concentration of combined REEs is 0.15 part per billion (ppb) 

(Elderfield and others, 1988). In major rivers, the REE concentrations average 1.6 ppb (Sholkovitz, 1995). Under extreme 

acid-mine drainage conditions (pH 1.1)—conditions that are 

not related to REE mining—dissolved REE concentrations can 

easily reach 2,500 ppb (Seal and others, 2008).

Information on pre-mining environmental concentrations 

of REEs is limited, in part because of their low concentrations. 

Data from solid media (rock, soil, and sediment) are more 

common because REEs are routinely analyzed in solid 

media. Analysis of REEs in water samples is less common. 

Environmental baseline studies have been conducted at the 

Bear Lodge deposit (Dahlberg, 2014) and at the Nechalacho 

deposit (Avalon Rare Metals Inc., 2011); both are currently 

advanced exploration projects. Baseline environmental data 

from all other REE deposit types have not been identified.

At Nechalacho, the surface water sample sites encompass 

an area that extends up to 6 kilometers from the deposit, 

but the groundwater samples are from within the deposit as 

reported by Avalon Rare Metals Inc. (2011). The pH of surface 

water and groundwater is neutral to slightly alkaline, ranging 

between 6.9 and 8.6. The cationic species in groundwater 

and surface water around the deposit are dominated by 

magnesium, followed by calcium and sodium. Water hardness 

values range between 71 and 470 ppm calcium carbonate 

(CaCO3

) for both surface water and groundwater. As reflected 

by alkalinity values that range between 74 and 442 ppm 

CaCO3, bicarbonate is the most important anionic species, 

followed by sulfate, chloride, and fluoride. Fluoride concentrations range from 0.06 to 4.4 ppm, which locally exceeds the 

U.S. Environmental Protection Agency secondary maximum 

contaminant limit of 2 ppm. Groundwater tends to have higher 

concentrations of anions than does surface water.

Trace elements are generally low in the vicinity of the 

Nechalacho deposit but tend to be slightly higher in groundwater compared with surface water. Baseline dissolved iron 

concentrations in surface water and groundwater are noteworthy and range from less than 30 to more than 10,000 ppb. 

Baseline dissolved copper concentrations range from 

0.2 to 48 ppb. Dissolved aluminum concentrations generally 

range from 0.5 to 108 ppb, although one outlier sample of 

groundwater measured 15,300 ppb. Dissolved manganese 

ranges from 0.2 to 544 ppb. Baseline dissolved nickel and 

molybdenum concentrations range from 0.25 to 6.6 ppb and 

from 0.2 to 63 ppb, respectively. Dissolved concentrations 

of uranium range from less than 0.05 to 20 ppb. Dissolved 

concentrations of arsenic and antimony are less than 

5 and 0.3 ppb, respectively.

Baseline surface water and groundwater chemistry data 

associated with the Bear Lodge deposit (Dahlberg, 2014), 

are grossly similar to those from Nechalacho. The pH of 

waters at the Bear Lodge deposit spans a slightly larger range 

(6.0 to 8.4). Water hardness ranges from 10 to 380 ppm 

CaCO3. Bicarbonate is the dominant cation as reflected by 

the alkalinity values of between 10 and 300 ppm CaCO3, and 

fluoride is present in concentrations of less than 1 ppm. Trace 

element concentrations at the Bear Lodge deposit are similar 

to those at Nechalacho.

Mine Waste Characteristics

The mineralogy of the dominant minerals from 

carbonatite and alkaline intrusion REE deposits influences 

the environmental character of their mine waste. The paucity 

of sulfide minerals, including pyrite, minimizes or eliminates 

concerns about acid-mine drainage for carbonatite-hosted REE 

deposits and alkaline-intrusion-related REE deposits. In fact, 

tailings from metallurgical testing at the Nechalacho deposit 

contain less than 0.1 weight percent total sulfur (Avalon 

Rare Metals Inc., 2011). The low acid-generating potential of 

these deposits is further offset by their high acid-neutralizing 

potential, particularly for carbonatite deposits, which are 

dominated by carbonate minerals.

In terms of trace elements, the REE ore minerals factor 

most prominently in the environmental character of the ores 

and mine wastes. The main ore minerals currently processed 

for REEs—bastnaesite (LnCO3

F), monazite ((Th,Ln)PO4 ), and 

xenotime ((Y,Ln)PO4) (where “Ln” [for lanthanide] indicates 

the place of REEs in the mineral formulas)—all contain 

appreciable amounts of uranium and thorium. Bastnaesite 

reportedly contains between 0 and 0.3 weight percent thorium 

dioxide (ThO2) and 0.09 weight percent uranium dioxide 

(UO2 ); monazite, between zero and 20 weight percent ThO2

and zero and 16 weight percent UO2; and xenotime, between 

zero and 5 weight percent UO2 (Long and others, 2010). The 

thorium and uranium contents of the ore minerals and mine 
O22 Critical Mineral Resources of the United States—Rare-Earth Elements

waste represent one of the biggest environmental challenges 

that must be managed during mining, ore processing, and 

mine closure. The geochemistry of thorium and uranium 

influence their interactions with the environment and resulting 

environmental effects. Thorium is highly insoluble under most 

conditions. Therefore, for thorium and its radioactive decay 

products, solid-phase pathways, such as windblown dust 

from tailings piles and erosion of tailings into water bodies, 

are of greatest priority for mine planning. Under oxidizing 

conditions, uranium is highly soluble; therefore, aqueous 

pathways leading to surface water or groundwater are priority 

issues for mine planning.

According to the metallurgical testing done, the 

geochemistry of the tailings at Nechalacho (Avalon Rare 

Metals Inc., 2011) is anomalous for several elements 

compared with the guidelines for soil (U.S. Environmental 

Protection Agency, 2015) and sediment (MacDonald 

and others, 2000). The chromium concentration of the 

metallurgical reject material (640 ppm) exceeds the 

residential soil guideline (280 ppm). Cadmium (6 ppm), 

chromium (640 ppm), and nickel (330 ppm) concentrations 

of the reject material exceed sediment guidelines for “probable effects concentrations” for these elements (4.98 ppm, 

111 ppm, and 48.6 ppm, respectively). The “probable effects 

concentration” in sediment is the concentration above which 

harmful effects on freshwater ecosystems are likely to be 

observed (MacDonald and others, 2000).

Human Health Concerns

Human health risks associated with mining are typically 

associated either with contamination of potential sources of 

drinking water or ingestion of fine particulate mine waste

(tailings). The human health effects of REEs have been 

sparsely studied since the work of Haley (1965). Ingestion and 

inhalation studies on cerium oxide and cerium compounds 

indicate minimal effects in laboratory studies on rats 

(U.S. Environmental Protection Agency, 2009). Exposure 

to cerium oxides and cerium compounds is unlikely during 

routine mining, although the potential for exposure is present 

during ore processing and refining. The United States lacks 

drinking water standards for the REEs. The limited data on 

the toxicity of REEs in drinking water has been reviewed by 

de Boer and others (1996).

For other trace elements, some baseline groundwater 

and surface waters in the vicinity of the Nechalacho 

deposit exceed the secondary maximum contaminant limit 

for fluoride. At the Bear Lodge deposit, a single baseline 

groundwater sample (13 ppb arsenic) is slightly above the 

U.S. Environmental Protection Agency drinking water 

standard for arsenic concentration (10 ppb). Several baseline 

surface water samples from the Bear Lodge Mountains as well 

as a single humidity-cell leachate sample from Nechalacho 

exceed the combined radium radioisotope (226Ra+ 228Ra) 

drinking water standard.

Ecological Health Concerns

The primary pathways for REE mining to affect aquatic 

ecosystems are through water or sediment contamination. 

Knowledge regarding the toxicity of REEs to aquatic organisms through either pathway is limited. Environmental guidelines for REEs in surface water and sediment are lacking in 

the United States. Guidelines for surface water and sediment 

for both fresh and salt water have been recommended in the 

Netherlands, however (Sneller and others, 2000). Several 

studies have documented that the toxicity and biologic uptake 

of REEs in surface water vary as a function of water hardness 

and alkalinity (Barry and Meehan, 2000; Moermond and 

others, 2001). Further, Weltje and others (2004) found that the 

toxicity of lutetium was proportional to the concentration of 

the free ion (Lu+3) in microbial bioassay studies. Therefore, 

REE complexation with carbonate, sulfate, or other ligands 

in natural systems may significantly influence their toxicity 

(Wood, 1990). Published aqueous REE data are lacking for 

the Nechalacho and the Bear Lodge deposits.

Insights into potential ecological risks associated with 

REE deposits in North America and worldwide are limited to 

data associated with the baseline environmental characterization of the Nechalacho (Avalon Rare Metals Inc., 2011) and 

Bear Lodge (Dahlberg, 2014) deposits. Uranium is the only 

trace element in baseline surface water samples near both 

deposits. Humidity-cell leachates from Nechalacho exceed the 

chronic surface water benchmark for uranium (1.87 ppb) but 

not the acute benchmark (33.5 ppb) (Suter, 1996). Dissolved 

thorium concentrations are approximately one to two orders 

of magnitude lower than dissolved uranium concentrations 

at Nechalacho. Therefore, the risk of mine drainage issues 

associated with tailings from either alkaline-intrusion-related 

or carbonatite-hosted deposits is likely minimal.

Accidental release of mill tailings, either during or after 

closure, can pose a risk to aquatic organisms. This risk can be 

assessed by the comparison of the geochemical composition 

of mill tailings with relevant sediment-quality guidelines. 

In general, most concentrations of trace elements are low. 

The concentrations of cadmium, chromium, and nickel in a 

tailings sample from Nechalacho exceeded both the threshold 

effects concentration and the probable effects concentration 

(MacDonald and others, 2000). Concentrations of the REEs 

in the tailings sample exceeded their corresponding negligible 

concentrations, but not their maximum permissible concentrations (Sneller and others, 2000). Therefore, only a few 

elements are marginally of concern in the event of accidental 

release of tailings.

Carbon Footprint

There are several potential links between REEs and 

greenhouse gas emissions. One of the main ore minerals, 

bastnaesite (Ln(CO3)F), contains carbonate, which is liberated 

during ore processing. Because of the low grades of economic 
Problems and Future Research O23

deposits (<10 percent total REOs), however, this carbon flux is 

not likely to be significant relative to other sources associated 

with mining. In addition, one of the most important uses of 

REEs is in permanent magnets used in “green” technologies, 

such as electric cars, hybrid cars, and wind turbines.

Mine Closure

Mine closure procedures depend on the mining method 

employed. Mountain Pass in California, Bayan Obo in 

China, and Mount Weld in Australia are all mined as open 

pits. At closure, the pits will likely fill with water, depending 

on the elevation of the local water table. Waste rock piles 

and mill tailing piles will also remain on site. Both types of 

piles may be regraded and revegetated. Recent operations 

at Mountain Pass recycled all water to avoid the need for 

evaporation ponds (Molycorp, Inc., 2015), which were used 

during previous mining at the site, to manage process water. 

Mill tailings disposal included paste disposal, eliminating 

the need for a tailings dam and its long-term maintenance. 

(Paste disposal produces tailings that have been significantly 

dewatered, and, as a result, the tailings do not have a critical 

flow velocity when pumped into the disposal area.)

Problems and Future Research

The fascinating new era for the REEs involves many 

challenges that remain to be fully addressed, including those 

related to economics (supply and demand and price stability 

issues), technology, and scientific research. In recent years, 

considerable news coverage and discussion have focused 

on the global supply and demand concerns for the REEs, 

underscored by their importance to modern and evolving 

technologies. In the foreseeable future, will processed 

REE compounds be readily available and affordable and in 

adequate supply to meet the needs of technology applications? 

If not, will suitable substitutes for the REEs be developed? 

In the fields of geologic research, although REE deposits 

have been studied for more than a century, many questions 

remain about their genesis. The questions that follow should 

be viewed as a partial list of the many issues likely to involve 

the REEs during the 21st century. These questions also 

represent opportunities where significant new contributions 

can be made.

• Will the addition of new LREE mines—such as 

the Mount Weld Mine (Australia)—prove to be 

adequate to offset reductions in the exports of REEs 

from China?

• Do substantial sources of HREEs occur outside 

of China? Are the ion-adsorption clay deposits of 

southern China unique? These deposits serve as the 

world’s principal source of the HREEs and, as of 

2015, other HREE deposits had not been brought 

into production. Can other clay deposits similarly 

enriched in the HREEs be brought into production 

outside of southern China? Will other types of HREE 

deposits eventually be brought into production?

• What exploration guidelines and techniques can 

be applied or developed that lead to discoveries of 

high-grade deposits of REEs, particularly hidden 

(concealed) deposits?

• Can advances in ore processing provide an economic 

advantage to some deposits? For mines that have 

REEs in the ores that host another primary mineral 

commodity, can REEs be developed as a byproduct?

• Can REEs be efficiently and economically recovered 

and recycled from disposed electronics, batteries, and 

other devices?

• Can REEs be replaced in some current applications 

by other metals that are less costly and more readily 

available? If so, will these products perform as well 

as those that use REEs?

• Can the complex ores of peralkaline deposits be 

processed by methods that are efficient and economical? This aspect is especially relevant because these 

deposit types often have a general enrichment in the 

highly sought HREEs.

• Why do REEs occur only in modest concentrations in Earth’s upper mantle, yet they can be highly 

concentrated in carbonate and alkaline magmas that 

ascend to the crust, thereby forming ore deposits? 

The systematic geologic and chemical processes that 

explain these observations are not well understood, 

especially in comparison to many other ore-forming 

systems. Exploration for REEs throughout the world 

in recent years has provided considerable drilling 

and other data, which are likely to provide an abundance of materials and information useful for detailed 

geologic research.

Case histories for the environmental behavior of REE 

deposits, mines, and their mine wastes are extremely limited 

because there are so few past and current mines. Further, 

most past and current mines are located in China, where little 

environmental information is available. Therefore, insights 

into future environmental characteristics of REE mine wastes 

must rely on predictive models.

Little is known about the aquatic toxicity of REEs 

(Sneller and others, 2000). Although their low solubility 

likely limits their toxicity, a few studies suggest that their 

aquatic toxicity varies as a function of water chemistry 

(Barry and Meehan, 2000; Moermond and others, 2001; 

Weltje and others, 2004). More study is needed to evaluate 

the potential environmental risks associated with REEs in 

hydrologic systems.
O24 Critical Mineral Resources of the United States—Rare-Earth Elements

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Rhenium

By David A. John, Robert R. Seal II, and Désirée E. Polyak

Chapter P of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–P

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter P

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................P1

Introduction..................................................................................................................................................P1

Uses, Demand, and Availability of Supply .....................................................................................P1

Substitutes for Rhenium ...................................................................................................................P3

Geology.........................................................................................................................................................P3

Geochemistry......................................................................................................................................P3

Mineralogy..........................................................................................................................................P4

Deposit Types .....................................................................................................................................P4

Rhenium in Porphyry Copper Deposits ...............................................................................P10

Rhenium in Sediment-Hosted Strata-Bound Copper Deposits .......................................P12

Occurrences in Sandstone-Type Strata-Bound Copper Deposits ........................P14

Occurrences in Reduced-Facies-Type (Kupferschiefer) Strata-Bound 

Copper Deposits ...............................................................................................P14

Rhenium in Sandstone-Type Uranium Deposits ................................................................P15

Rhenium in the Merlin Deposit .............................................................................................P17

Resources and Production ......................................................................................................................P18

Identified Resources .......................................................................................................................P18

Production.........................................................................................................................................P18

Undiscovered Resources ...............................................................................................................P18

Exploration for New Deposits .................................................................................................................P19

Environmental Considerations ................................................................................................................P19

Sources and Fate in the Environment ..........................................................................................P20

Mine Waste Characteristics ..........................................................................................................P21

Human Health Concerns .................................................................................................................P23

Ecological Health Concerns ...........................................................................................................P24

Carbon Footprint...............................................................................................................................P25

Mine Closure.....................................................................................................................................P25

Problems and Future Research ..............................................................................................................P25

Acknowledgments....................................................................................................................................P25

References Cited.......................................................................................................................................P26

Appendix P1. Rhenium Data Sources and Limitations of Data Used in Rhenium 

Resource Estimates ....................................................................................................................P48

Suggested citation:

John, D.A., Seal, R.R., II, and Polyak, D.E., 2017, Rhenium, chap. P of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and 

Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and prospects 

for future supply: U.S. Geological Survey Professional Paper 1802, p. P1– P49, http://dx.doi.org/10.3133/pp1802P.
Rhenium

By David A. John, Robert R. Seal II, and Désirée E. Polyak

Abstract

Rhenium is one of the rarest elements in Earth’s 

continental crust; its estimated average crustal abundance is 

less than 1 part per billion. Rhenium is a metal that has an 

extremely high melting point and a heat-stable crystalline 

structure. More than 80 percent of the rhenium consumed in 

the world is used in high-temperature superalloys, especially 

those used to make turbine blades for jet aircraft engines. 

Rhenium’s other major application is in platinum-rhenium 

catalysts used in petroleum refining.

Rhenium rarely occurs as a native element or as its own 

sulfide mineral; most rhenium is present as a substitute for 

molybdenum in molybdenite. Annual world mine production 

of rhenium is about 50 metric tons. Nearly all primary 

rhenium production (that is, rhenium produced by mining 

rather than through recycling) is as a byproduct of copper 

mining, and about 80 percent of the rhenium obtained through 

mining is recovered from the flue dust produced during the 

roasting of molybdenite concentrates from porphyry copper 

deposits. Molybdenite in porphyry copper deposits can contain 

hundreds to several thousand grams per metric ton of rhenium, 

although the estimated rhenium grades of these deposits range 

from less than 0.1 gram per metric ton to about 0.6 gram per 

metric ton.

Continental-arc porphyry copper-molybdenum-gold 

deposits supply most of the world’s rhenium production 

and have large inferred rhenium resources. Porphyry copper 

mines in Chile account for about 55 percent of the world’s 

mine production of rhenium; rhenium is also recovered from 

porphyry copper deposits in the United States, Armenia, 

Kazakhstan, Mexico, Peru, Russia, and Uzbekistan. 

Sediment-hosted strata-bound copper deposits in Kazakhstan 

(of the sandstone type) and in Poland (of the reduced-facies, 

or Kupferschiefer, type) account for most other rhenium 

produced by mining. These types of deposits also have large 

amounts of identified rhenium resources. The future supply 

of rhenium is likely to depend largely on the capacity of the 

specialized processing facilities needed to recover rhenium 

from molybdenite concentrates.

The environmental consequences of rhenium recovery are 

closely linked to the consequences of mining large porphyry 

copper and strata-bound copper deposits; no additional 

environmental impact from recovery of rhenium from these 

deposits has been identified. No information is available 

regarding the potential toxic effects of rhenium on humans, 

partly because of the low natural abundance of rhenium.

Introduction

Rhenium (Re) is one of the rarest elements in Earth’s 

continental crust; its average crustal abundance is about 0.4 part 

per billion (ppb), although estimates range from 0.2 to 2 ppb 

(Taylor and McLennan, 1995; Rudnick and Gao, 2003; Sun and 

others, 2003). This silvery-white metallic element’s extremely 

high melting point (3,180 degrees Celsius [°C]) and heat-stable 

crystalline structure make it an excellent refractory metal (fig. P1).

Rhenium was the last stable, naturally occurring element 

discovered, although its existence was long predicted from 

the periodic table. The element was discovered in 1925 by 

German chemists Walter Noddack, Ida Tacke, and Otto Berg, 

who concentrated rhenium from gadolinite ore that contained 

about 10 parts per million (ppm) rhenium (Emsley, 2001). The 

German chemists subsequently separated 1 gram of rhenium 

from 660 kilograms (kg) of molybdenum ore.

Uses, Demand, and Availability of Supply

Since the late 1980s, the two most important uses of 

rhenium have been in high-temperature superalloys and in 

platinum-rhenium catalysts (fig. P2). The high-temperature 

alloys in which rhenium is used include several nickel-base 

superalloys that are used mainly in the manufacture of 

turbine blades for jet aircraft engines and in power-generation 

applications. The high-temperature properties of rhenium 

allow turbine engines to be designed with closer tolerances, 

thus enabling increased thrust and higher operating efficiency. 

Because the life cycle of turbine blades in jet engines is only 

about 10 years, significant quantities of used blades (which 
P2 Critical Mineral Resources of the United States—Rhenium

A

C D

B

Rhenium figure 1.

Figure P1. Photographs of rhenium and rhenium compounds. A, A single crystal of high-purity (99.999%) rhenium (left), a remelted 

rhenium bar (center), and a 1-cubic-centimeter rhenium cube (right). Photograph courtesy of Alchemist-hp/CC-BY-NC-ND-3.0 

(http://commons.wikimedia.org/wiki/File:Rhenium_single_crystal_bar_and_1cm3_cube.jpg). B, Rhenium powder. Photograph courtesy 

of H.C. Starck. C, Ammonium perrhenate (NH4ReO4) powder, which is the most widely marketed form of rhenium. Photograph courtesy 

of Stanford Advanced Materials (http://www.samaterials.com/847/ammonium-perrhenate.jpg). D, Rheniite (ReS2) crystals from hightemperature fumaroles on the Kudryavyy volcano, which is located on Iturup Island in Sakhalinskaya Oblast’, Russia. Photograph 

courtesy of The Arkenstone, iRocks.com (http://www.irocks.com/minerals/specimen/42387).

typically contain 3 percent rhenium, although some alloys 

contain 6 percent rhenium) are accumulating. Technology 

is continuing to be developed to allow recycling of these 

blades to recover the rhenium, which can then be used in the 

manufacture of new, second-generation blades. Recycling of 

rhenium from used turbine blades could potentially reduce the 

requirement for primary rhenium (that is, rhenium produced 

by mining rather than through recycling) by about 50 percent. 

Most rhenium recycling is currently performed in the 

United States and Germany.

World consumption of rhenium from primary sources was 

estimated to be between 50 and 54 metric tons in 2012 (Polyak, 

2014). Approximately 83 percent (45 metric tons) of rhenium 

from primary sources was used in making superalloys, mainly 

for the aerospace industry (fig. P2). New types of catalysts in 

the petrochemical industry accounted for about an additional 

10 percent (5 metric tons). The remaining 7 percent (4 metric 

tons) of world consumption in 2012—primarily in the form of 

tungsten-rhenium and molybdenum-rhenium alloys—was used 

in a variety of other applications, including in the manufacture 

of electrical contact points, flashbulbs, heating elements, 

vacuum tubes, and X-ray tubes and targets. An additional 15 

metric tons of rhenium is recycled each year from spent catalysts and is not included in the world consumption total shown 

in figure P2 (Lipmann Walton & Co. Ltd., 2010; Minor Metals 

Trade Association, 2012).

Platinum-rhenium catalysts are used to produce highoctane lead-free gasoline. In the early 1970s, Chevron Corp. 

of the United States developed a series of platinum-rhenium 

catalysts that do not react with sulfur. As a response to the 

Organization of the Petroleum Exporting Countries (OPEC) 

oil embargo in the 1970s and the oil crisis of the early 1980s, 

a second generation of more-efficient catalysts was developed 

with double the platinum-rhenium content, which boosted 
Geology P3

Superalloys

83%

Catalysts

10%

Other

7%

Rhenium fig 2. 

Figure P2. Pie chart showing major end uses of 

primary rhenium as a percentage of world consumption 

in 2012. Total world production for the year was between 

50 and 54 metric tons. Approximately 83 percent of 

the rhenium from primary sources was used to make 

superalloys, mainly for the aerospace industry.

the refinery efficiency and the gasoline’s octane levels. 

Ammonium perrhenate (NH4ReO4) is the most common 

form of rhenium used by catalyst manufacturers (fig. P1C). 

Petroleum-reforming platinum-rhenium catalysts are used 

mainly in North America and Europe, where demand for 

gasoline and diesel fuel is highest. These are also the locations 

of the major rhenium-platinum catalyst manufacturers.

The recovery of rhenium from platinum-rhenium catalysts 

is a mature business. A closed-loop recycling system operated 

by catalyst manufacturers, refiners, and secondary rhenium 

producers ensures that consumption of newly mined rhenium 

is kept to a minimum. The catalysts are recycled mainly to 

recover the platinum content, so much of the rhenium is 

recycled only because of the platinum content of the catalysts.

Rhenium in molybdenite concentrates derived from 

porphyry copper ores is recovered at some mines as a 

byproduct during the molybdenite roasting process. In the 

United States in 2012, three molybdenum mines had roasters 

associated with mines, but only one of the roasters was 

equipped to recover rhenium. This roaster, which is located 

at the Sierrita facility in Arizona, is operated by FreeportMcMoran Copper & Gold Inc. of the United States. This 

facility processes all the byproduct molybdenite concentrates 

from Freeport-McMoran mines and processes molybdenite 

concentrates on a toll basis for third parties. In contrast, 

Kennecott Utah Copper Corp. (which is owned by Rio Tinto 

plc of the United Kingdom) sends byproduct molybdenite 

concentrates from the Bingham Canyon Mine in Utah to 

Molymex S.A. de C.V.’s plant in Mexico for roasting; the 

recovered rhenium concentrate is then sent to Molibdenos y 

Metales S.A. (Molymet) in Chile for purification.

Substitutes for Rhenium

Substitutes for rhenium in platinum-rhenium catalysts 

are being evaluated on an ongoing basis. Iridium and tin have 

achieved commercial success in one application. Other metals 

being evaluated as possible substitutes for rhenium in catalysts 

include gallium, germanium, indium, selenium, silicon, 

tungsten, and vanadium. The use of these and other metals 

in catalysts might someday decrease rhenium’s share of the 

existing catalyst market; however, rhenium-bearing catalysts 

that are being considered for use in several proposed gas-toliquid projects likely would offset this decrease. Materials that 

can substitute for rhenium in various end uses are cobalt and 

tungsten for coatings on copper X-ray targets, rhodium and 

rhodium-iridium for high-temperature thermocouples, tungsten 

and platinum-ruthenium for coatings on electrical contacts, 

and tungsten and tantalum for electron emitters (Polyak, 

2013). Aerospace superalloy producers continue to research 

new alloys with smaller quantities of rhenium; however, it has 

proven difficult to develop alloys with lower rhenium content 

for use in jet engines without a loss in performance.

Geology

Geochemistry

Rhenium has an atomic number of 75 and an atomic mass 

of 186.2 grams per mole. In the periodic table of elements, 

rhenium is a third row, heavy transition metal located in 

column 7B below manganese and technetium, adjacent to 

tungsten and the platinum-group elements (PGEs), and 

diagonally below molybdenum.

Rhenium has a density of 21.02 grams per cubic centimeter, making it the fourth densest element (iridium, osmium, 

and platinum have higher densities). It has a hexagonal 

close-packed crystal structure. The 3,180 °C melting point of 

rhenium is the third highest melting point of all the elements; 

only carbon (3,500 °C) and tungsten (3,422 °C) have higher 

melting points.

Rhenium has the widest range of valences of any 

element—nine in total, ranging from –1 to +7— although +7, 

+6, +4, and +2 are the most common ions. Rhenium forms 

three stable oxides—rhenium heptoxide (Re2

O7), rhenium 

trioxide (ReO3), and rhenium dioxide (ReO2 ) —of which 

rhenium heptoxide is the most common. Rhenium heptoxide 

is a bright yellow volatile solid that dissolves in water to 

form rhenic acid (HReO4). Rhenium has a marked affinity 

for sulfur and occurs mostly in nature as a substitute (solid 

solution) for molybdenum in molybdenite (MoS2). Unusual 

features of rhenium compounds include the unusually high 

volatility of rhenium heptoxide and the high solubility of 

rhenium heptoxide in water. These properties facilitate 

rhenium’s recovery from flue dusts produced by roasting of 

molybdenite concentrates.
P4 Critical Mineral Resources of the United States—Rhenium

The two dominant naturally occurring rhenium isotopes 

are 185Re, which is stable and accounts for 37.4 percent of 

naturally occurring rhenium, and 187Re, which is radiogenic 

and accounts for 62.6 percent of naturally occurring rhenium. 

Twenty-six other radioactive isotopes of rhenium have been 

recognized. 187Re decays to 187Os by beta decay and has a halflife of about 4.1×1010 years. The rhenium-osmium isotopic 

system is used to date sulfide minerals (most commonly 

molybdenite) in mineral deposits (McCandless and Ruiz, 

1993; Stein and others, 2001).

Mineralogy

Rhenium rarely occurs as a native element or as its 

own sulfide mineral. Trace amounts of native rhenium (Re0 ) 

were recently discovered in ultramafic rocks in Ukraine 

(Bobrov and others, 2008). Microscopic crystals of rheniite 

(ReS2 ) were first reported in 1986 in volcanic fumaroles 

on Mount Usu, Japan (Bernard and Dumortier, 1986), and 

megascopic crystals of rheniite were discovered in 1992 

in high-temperature fumaroles on the Kudryavyy (also 

spelled Kudriavy) volcano on Iturup Island in Sakhalinskaya 

Oblast’, Russia (fig. P1D; Korzhinsky and others, 1994). 

Rheniite also has been reported in the Pagoni Rachi porphyry 

copper-molybdenum prospect in northern Greece (Voudouris 

and others, 2009). Microscopic (≤ 75 micrometers [μm]) 

crystals of rhenium-rich tarkianite ((Cu,Fe)(Re,Mo)4S8) 

have been discovered in sulfide concentrates from the Hitura 

nickel-copper-PGE mine at Nivala, Finland (Kojonen and 

others, 2004). None of these rhenium-rich minerals are 

economic sources of rhenium, however.

Dzhezkazganite forms microscopic collomorphic 

aggregates and veinlets in bornite, chalcocite, and galena 

and replaces bornite in sandstone-hosted strata-bound copper 

deposits at the Dzhezkazgan Mine in Qaraghandy, Kazakhstan 

(Poplavko and others, 1962; Abisheva and others, 2001). 

Dzhezkazganite was first inferred as a copper-rhenium 

sulfide (CuReS4 ) mineral (Poplavko and others, 1962), but 

subsequent electron microprobe analyses suggest that it has a 

chemical formula of ReMoCu2

PbS6 (Genkin and others, 1994). 

It is not a recognized mineral species by the International 

Mineralogical Association, however, because of the lack 

of structural data and uncertainty in its chemical formula 

(Fleischer, 1963).

Molybdenite is the principal source of rhenium. The 

rhenium content of molybdenite varies widely from less than 

1 ppm to several weight percent, but molybdenite in porphyry 

copper deposits typically contains about 100 to 3,000 ppm 

rhenium (tables P1 and P2 at back of chapter; Fleischer, 1959; 

Giles and Schilling, 1972; Newberry, 1979; Berzina and 

others, 2005; Sinclair and others, 2009; John and Taylor, 2016). 

In contrast, the rhenium content of molybdenite in porphyry 

molybdenum deposits generally is much lower, in many cases 

less than 20 ppm (table P2). Rhenium contents of molybdenite 

in other types of porphyry deposits and in quartz veins 

typically also are low (<10 to about 200 ppm) (for example, 

Terada and others, 1971; Giles and Schilling, 1972; Ishihara, 

1988; Sinclair and others, 2009; Millensifer and others, 2014). 

Molybdenite in Australia’s Merlin molybdenum-rhenium 

deposit (discussed below), however, averages about 1,000 ppm 

rhenium (Brown and others, 2010).

The rhenium content of molybdenite also varies widely 

within some porphyry copper deposits (for example, it ranges 

from 130 to 2,000 ppm at Bingham, Utah) (Giles and Schilling, 

1972). The variable rhenium content, both within deposits and 

between different types of porphyry deposits, is not completely 

understood. Variations in the rhenium content of molybdenite 

within deposits may represent multiple generations of molybdenite that have different rhenium contents (for example, the 

Sar Cheshmeh Mine in Kermān Province, Iran; Aminzadeh 

and others, 2011). Rhenium concentration differences reflect a 

combination of factors, including changes in temperature and 

pressure, fluid composition (especially pH and sulfur content), 

oxidation state, and (or) transport mechanism (that is, brine, 

moderate-density liquid, or low-density vapor) during the 

complex magmatic-hydrothermal history of these deposits 

(John and Taylor, 2016).

Deposit Types

Nearly all primary rhenium production (rhenium produced 

by mining rather than through recycling) is as a byproduct of 

copper mining. Rhenium resources are dominantly contained 

in porphyry copper-molybdenum-gold deposits, which supply 

about 80 percent of the rhenium produced by mining (Polyak, 

2013). Rhenium in porphyry copper deposits is contained 

primarily as ReS2 in solid solution in molybdenite (Fleischer, 

1959). Most of the remaining rhenium production is as a 

byproduct of mining sediment-hosted strata-bound copper 

deposits, both the sandstone (red bed) types in Kazakhstan and 

the reduced-facies (Kupferschiefer) types in Poland (Hitzman 

and others, 2005; Zientek and others, 2013). Smaller amounts 

of rhenium are recovered from the processing of roll-front-type 

sandstone uranium ore in Kazakhstan (Dahlkamp, 2009a). The 

Merlin molybdenum-rhenium zone of the Mount Dore copper 

deposit in Queensland, Australia, is a potential source of highgrade rhenium ore and is the only deposit known in which 

rhenium could be a primary commodity (Brown and others, 

2010). The world’s major rhenium-bearing deposits that are 

described in the text and (or) listed in table P1 are shown on 

the map in figure P3.
Geology P5

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Rhenium, Fig 3.

Bagdad

Bingham

Butte

Morenci

Santa Rita

Sierrita-Esperanza

Pebble

Chuquicamata-Radomiro TomicLa EscondidaEl SalvadorLos PelambresLos Bronces-Rio BlancoEl Teniente

La Caridad

Cerro Verde

Toquepala

Kadjaran Kalmakyr

Zuun Mod Molybdenum

Lubin-Sieroszowice

Mansfeld-Sangerhausen

Dzhezkazgan Zhaman- Aibat

Suluchekinskoye

Sugraly

Merlin

Rhenium-bearing deposits, by type

Undefined Mo-ReRoll-front sandstone UReduced-facies strata-bound CuSandstone strata-bound CuPorphyry Cu-(Mo-Au)

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

EXPLANATION

Figure P3. World map showing locations of major rhenium-bearing deposits listed in table P1, including porphyry copper-molybdenum-gold deposits; sediment-hosted strata-bound copper deposits in Poland (of the reduced-facies, or Kupferschiefer, type) and Kazakhstan (of the sandstone type); roll-front-type sandstone uranium deposits in Kazakhstan and Uzbekistan; and the Merlin (Mount Dore) molybdenum-rhenium deposit in Queensland, Australia. Additional information about the rhenium resources in these deposits is in table P1. Au, gold; Cu, copper; Mo, molybdenum; Re, rhenium; U, uranium
P6 Critical Mineral Resources of the United States—Rhenium

Table P1. Summary of rhenium, copper, and molybdenum grades, deposit tonnage, and amount of contained rhenium in the rheniumbearing deposits shown in figure P3.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compounds: Au, gold; Cu, copper; Mo, molybdenum; MoS2, molybdenite; Os, osmium; Re, rhenium; U, uranium. Units of measure: g/t, gram 

per metric ton; Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; n.a., not applicable to the deposit]

Deposit

name

Country

(State/Province)

Deposit

type

Tectonic 

setting1

Deposit 

tonnage 

(Mt)

Grade

(wt. % Cu)

Grade

(wt. % Mo)

Minimum 

Re in MoS2

(ppm)

Maximum 

Re in MoS2

(ppm)

Estimated average Re in MoS2

(ppm)

Bagdad United States 

(Arizona)

Porphyry Cu-

(Mo-Au)

CA 1,600 0.400 0.01 330 642 460 

Bingham United States 

(Utah)

Porphyry Cu-

(Mo-Au)

PC 3,230 0.882 0.053 130 2,000 250 

Butte United States 

(Montana)

Porphyry Cu-

(Mo-Au)

CA 5,220 0.673 0.028 — — 240 

Morenci United States 

(Arizona)

Porphyry Cu-

(Mo-Au)

CA 6,470 0.524 0.0095 270 640 455 

Pebble United States 

(Alaska)

Porphyry Cu-

(Mo-Au)

PC 5,940 0.592 0.0243 329 2,070 1,100 

Santa Rita United States 

(New Mexico)

Porphyry Cu-

(Mo-Au)

CA 3,030 0.468 0.008 700 1,200 800 

SierritaEsperanzaUnited States 

(Arizona)

Porphyry Cu-

(Mo-Au)

CA 2,262 0.294 0.0292 90 1,800 238 

ChuquicamataRadomiro 

Tomic

Chile Porphyry Cu-

(Mo-Au)

CA 21,277 0.860 0.04 93 262 265 

El Salvador Chile Porphyry Cu-

(Mo-Au)

CA 3,836 0.860 0.022 — — 585 

El Teniente Chile Porphyry Cu-

(Mo-Au)

CA 20,731 0.620 0.019 25 1,154 420 

La Escondida Chile Porphyry Cu-

(Mo-Au)

CA 11,158 0.769 0.0062 95 1,805 886 

Los Bronces–

Rio Blanco

Chile Porphyry Cu-

(Mo-Au)

CA 16,816 0.601 0.02 104 898 265 

Los Pelambres Chile Porphyry Cu-

(Mo-Au)

CA 7,458 0.617 0.015 450 820 600 

La Caridad Mexico (Sonora) Porphyry Cu-

(Mo-Au)

CA 1,800 0.452 0.0247 72 570 570 

Cerro Verde Peru Porphyry Cu-

(Mo-Au)

CA 2,258 0.495 0.01 3,060 3,497 3,280 

Toquepala Peru Porphyry Cu-

(Mo-Au)

CA 2,320 0.550 0.04 387 1,496 600 

Kadjaran 

(Kadzharan)

Armenia Porphyry Cu-

(Mo-Au)

CA 1,700 0.270 0.055 33 2,620 280 
Geology P7

Table P1. Summary of rhenium, copper, and molybdenum grades, deposit tonnage, and amount of contained rhenium in the rheniumbearing deposits shown in figure 3.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compounds: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium; U, uranium. Units of measure: g/t, gram per 2

metric ton; Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; n.a., not applicable to the deposit]

Number 

of Re 

analyses

Selected 

sample 

type2

Grade 

(g/t Re)

Contained 

Re 

(t)

Mo:Re 

ratio Notes Data sources Deposit name

9

43

1

4

8

9

6

9

3

17

7

15

3

3

2

5

238

2

1

1

1

3

3

2

3

3

3

2

2

2

3

2

3

1

0.080

0.221

0.112

0.072

0.446

0.107

0.116

0.18

0.215

0.133

0.092

0.09

0.150

0.235

0.12

0.40

0.257

130

714

585

466

2,650

324

262

3,800

825

2,760

1,030

1,500

1,120

423

300

930

437

1,200

2,400

2,500

1,320

545

746

2,520

2,300

1,020

1,430

672

2,300

1,000

1,050

830

1,000

2,140

None

None

None

None

Re grade from bulk 

molybdenite concentrate samples 

with 52% Mo; 

measured and 

indicated resources

None

None

None

None

None

None

None

None

None

None

None

None

Sutulov, 1974; Nadler, 1997; Barra and 

others, 2003; Singer and others, 2008

Giles and Schilling, 1972; McCandless 

and Ruiz, 1993; Chesley and Ruiz, 

1998; Singer and others, 2008; Austen 

and Ballantyne, 2010; J. Chesley, 

written commun., 2013

Giles and Schilling, 1972; Singer and 

others, 2008

Giles and Schilling, 1972; McCandless and

Ruiz, 1993; Singer and others, 2008

Ghaffari and others, 2011; Lang and 

others, 2013

Giles and Schilling, 1972; Sutulov, 1974; 

Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Singer and others, 2008; 

Barra and others, 2013

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Maksaev and others, 

2004; Singer and others, 2008

Mathur and others, 2001; Singer and others, 2008; Romero and others, 2010

Mathur and others 2001; Singer and others, 2008; Deckart and others, 2013

Mathur and others 2001; Singer and 

others, 2008

Nadler, 1997; Valencia and others, 2005; 

Singer and others, 2008

Mathur and others, 2001; Singer and 

others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Mathur and others, 

2001; Singer and others, 2008

Nadler, 1997; Berzina and others, 2005; 

Singer and others, 2008

Bagdad

Bingham

Butte

Morenci

Pebble

Santa Rita

SierritaEsperanza

ChuquicamataRadomiro 

Tomic

El Salvador

El Teniente

La Escondida

Los BroncesRio Blanco

Los Pelambres

La Caridad

Cerro Verde

Toquepala

Kadjaran 

(Kadzharan)
P8 Critical Mineral Resources of the United States—Rhenium

Table P1. Summary of rhenium, copper, and molybdenum grades, deposit tonnage, and amount of contained rhenium in the rheniumbearing deposits shown in figure P3.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compounds: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium; U, uranium. Units of measure: g/t, gram 2

per metric ton; Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; n.a., not applicable to the deposit]

Deposit 

name

Country 

(State/Province)

Deposit 

type

Tectonic 

setting1

Deposit 

tonnage 

(Mt)

Grade 

(wt. % Cu)

Grade 

(wt. % Mo)

Minimum 

Re in MoS 2

(ppm)

Maximum 

Re in MoS 2

(ppm)

Estimated average Re in MoS 2

(ppm)

Kalmakyr 

(Almalyk)

Zuun Mod 

Molybdenum

LubinSieroszowice

MansfeldSangerhausen

MansfeldSangerhausen

Dzhezkazgan

Zhaman-Aibat 

(or ZhamanAybat) 

(Zhomart)

Merlin

Sugraly

Suluchekinskoye

Uzbekistan

Mongolia

Poland

Germany

Germany

Kazakhstan

Kazakhstan

Australia 

(Queensland)

Uzbekistan

Kazakhstan

Porphyry Cu-

(Mo-Au)

Porphyry Cu-

(Mo-Au)

Reduced-facies 

(Kupferschiefer) 

stratabound Cu

Reduced-facies 

(Kupferschiefer) 

stratabound Cu

Reduced-facies 

(Kupferschiefer) 

stratabound Cu

Sandstone 

stratabound Cu

Sandstone 

stratabound Cu

Undefined 

Mo-Re

Roll-front 

sandstone U

Roll-front 

sandstone U

CA

CA

n.a.

n.a.

n.a.

n.a.

n.a.

n.a.

n.a.

n.a.

2,000

237

1,181

35

155

2,000

161

6.9

30

330

0.380

0.069

1.58

2.41

2.41

1.1

1.69

—

—

—

0.006

0.059

—

0.015

0.015

—

—

1.383

—

—

700

250

—

—

—

—

—

—

—

—

2,000

300

—

—

—

—

—

—

—

—

1,500

275

—

—

—

—

—

—

—

—

1

Tectonic setting of porphyry copper deposits: CA, continental arc; PC, post-collisional.

2

Sample type used in calculating Re grade: 1, molybdenite separate; 2, molybdenite separate used in Re-Os dating; 3, molybdenite mill concentrate.
Geology P9

Table P1. Summary of rhenium, copper, and molybdenum grades, deposit tonnage, and amount of contained rhenium in the rheniumbearing deposits shown in figure 3.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compounds: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium; U, uranium. Units of measure: g/t, gram per 2

metric ton; Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; n.a., not applicable to the deposit]

Number 

of Re 

analyses

Selected 

sample 

type2

Grade 

(g/t Re)

Contained 

Re 

(t)

Mo:Re 

ratio Notes Data sources Deposit name

21

2

—

—

—

—

—

—

—

—

1

1

—

—

—

—

—

—

—

—

0.150

0.270

0.6

21

21

1.5

1.45

22.7

12.5

1.5

300

59

709

740

3,300

3,000

233

157

375

495

400

2,190

—

7.1

7.1

0.5 to 1.2

—

609

—

—

None

Average of range of 

Re contents 

in MoS2

2011 end-of-year 

proven and probable reserves; 

average Re grade of 

ore mined in 2009

Estimated remaining 

resources after 

mining ceased in 

1990

Estimated total 

resources before 

mining, which 

began in 1200 and 

ended in 1990

Ore averages 

1 to 2 ppm Re

None

Average Re grade and 

tonnage of indicated and inferred 

resources

Ore averages 

10 to 15 ppm Re

Ore averages 1 to 

2 ppm Re and 

0.07 to 0.13% U

Sutulov, 1974; Singer and others, 2008; 

Pašava and others, 2010

Clark and Baudry, 2011

Smakowski and others, 2010; 

Bartlett and others, 2013

Jankowski, 1995

Jankowski, 1995

Seyfullin and others, 1974; 

Box and others, 2013

Box and others, 2013

Lycopodium Minerals QLD Pty Ltd., 

2012

Dahlkamp, 2009b

Dahlkamp, 2009a

Kalmakyr 

(Almalyk)

Zuun Mod 

Molybdenum

LubinSieroszowice

MansfeldSangerhausen

MansfeldSangerhausen

Dzhezkazgan

Zhaman-Aibat 

(or Zhaman-Aybat) 

(Zhomart)

Merlin

Sugraly

Suluchekinskoye
P10 Critical Mineral Resources of the United States—Rhenium

Rhenium in Porphyry Copper Deposits

Large continental-margin arc porphyry coppermolybdenum-gold deposits dominate world rhenium resources 

(tables P1 and P2; figs. P3 and P4; John and Taylor, 2016). 

Porphyry copper deposits are large (in most cases, greater 

than 100 million metric tons), low-to–moderate grade 

(in most cases, 0.3 to 2.0 percent copper), disseminated, 

breccia and vein-hosted copper deposits (John and others, 

2010). Porphyry copper deposits are parts of larger porphyry 

copper systems in which large volumes (10 to >100 cubic 

kilometers) of hydrothermally altered rock are centered 

on porphyry stocks and other intrusions. The deposits may 

include associated skarn, carbonate-replacement, sedimenthosted, and high- and intermediate-sulfidation epithermal 

base and precious-metal deposits (Sillitoe, 2010). Porphyry 

copper systems most commonly form above active subduction zones at convergent plate margins and are associated 

with calc-alkaline batholiths and volcanic arcs in both 

continental-margin and island-arc settings (Sillitoe, 1972; 

Richards, 2003). Some other porphyry copper systems form 

in post-collisional and other tectonic settings after subduction 

ends (Richards, 2009; Hou and others, 2011). Copper is the 

dominant metal produced from porphyry copper deposits; 

molybdenum, gold, and lesser amounts of silver, rhenium, and 

PGEs (mostly palladium) are important byproducts in some 

deposits. Hydrothermal alteration of porphyry-copper-related 

intrusive rocks and their wallrocks is widely developed and 

includes alkali-dominated assemblages (potassic, sodic, and 

sodic-calcic), acid (low pH) assemblages (advanced argillic 

and sericitic), and propylitic assemblages. Alteration zoning 

can be highly variable, but acidic alteration is typically distal 

and shallow and, where it is found in the center of a deposit, 

it is late relative to alkali alteration assemblages. The deeper 

parts of porphyry copper systems (typically paleodepths 

of approximately 2 to 10 kilometers [km]) contain central 

porphyry copper ± molybdenum ± gold mineralization 

(fig. P5A) and may contain flanking copper, gold, and (or) zinc 

skarn, silver-lead-zinc carbonate replacement, and gold-silver 

sediment-hosted deposits. The shallow, overlying parts of 

these systems may host high- and intermediate-sulfidation 

epithermal gold-silver ± copper deposits. Copper, gold, 

and molybdenum mineralization is generally associated 

with potassic and sericitic alteration. Rhenium is invariably 

associated with molybdenum mineralization.

Rhenium is produced from several Tertiary porphyry 

copper deposits in the Andes Mountains in South America, 

including the giant Chuquicamata-Radomiro Tomic and 

El Teniente deposits in Chile, and Toquepala in Peru; from 

Late Cretaceous to mid-Tertiary deposits in the Western 

United States; and from deposits in Armenia, Kazakhstan, 

Russia, and Uzbekistan (fig. P3). Rhenium in porphyry copper 

1,000 t Re

10,000 t Re

100 t Re

10 t Re

1 t Re

0.1 t Re

Grade, in grams per metric ton Re

0.01

0.10

1

10

100

1 10 100 1,000 10,000 100,000

Minable material, in million metric tons

Porphyry Cu-(Mo-Au)

Deposit type

EXPLANATION

Arc-related porphyry Mo

AFRG porphyry Mo

Reduced-facies strata-bound Cu 

Sandstone strata-bound Cu

Undefined Mo-Re 

Roll-front sandstone U

MansfeldSangerhausen

post-mining

resources

Merlin

MansfeldSangerhausen

pre-mining

resources

Rhenium figure 4

Dzhezkazgan

Quartz Hill Climax

Lubin

Pebble

Bingham

Los Pelambres

El Teniente

Los Bronces–

 Rio Blanco Escondida

Chuquicamata

Figure P4. Plot of rhenium grade versus deposit tonnage for major rhenium-bearing deposits in the world, 

including those shown in figure P3. The diagonal lines are isolines of contained rhenium, in metric tons. 

Additional resource data for these deposits, and sources, are given in table P1. AFRG, alkali-feldspar rhyolitegranite; t, metric ton. Elements: Au, gold; Cu, copper; Mo, molybdenum; Re, rhenium, U, uranium
Geology P11

6,000

8,000

FEET

4,000

2,000

Sea level

NORTHWEST SOUTHEAST

1000

0 500 1,000 1,500 FEET

200 300 400 500 METERS VERTICAL EXAGGERATION ×2

Rhenium figure 5A. 

~0.15 ppm Au

~0.30 ppm Au

~1.00 ppm Au

>0.02 wt. % Mo

>0.08 wt. % Mo

>0.15 wt. % Cu

>0.35 wt. % Cu

>0.70 wt. % Cu

Copper-molybdenum-gold (Cu-Mo-Au) grades

Rock type

EXPLANATION

High Re

Lower

Re

Lower Re Lower

 Re

Paleozoic sedimentary rocks

Equigranular monzonite

Quartz-monzonite porphyry

Pre-mining topography

Pit outline 2006

UTAH

Bingham

Location of section

Figure P5. Cross sections illustrating rhenium occurrences in major deposit types from which 

rhenium is recovered or potentially recoverable. A, Bingham Canyon porphyry copper-molybdenumgold (Cu-Mo-Au) deposit in Utah showing metal zoning around porphyry intrusions. Molybdenum (in 

molybdenite) is displaced to deeper levels than copper and gold, which are mostly deposited earlier 

than molybdenum. Rhenium (Re) grades apparently are highest in the copper-molybdenum-gold ores in 

the center of the deposit (“High Re”) and lower in the deep molybdenum-only ore (“Lower Re”) (Austen 

and Ballantyne, 2010). Figure modified from Landtwing and others (2010) and Seo and others (2012). 

ppm, part per million; wt.%, weight percent

A
P12 Critical Mineral Resources of the United States—Rhenium

deposits is contained primarily as ReS2 in solid solution in 

molybdenite (Fleischer, 1959) at concentrations ranging 

from less than 10 ppm to about 1 weight percent (table P1). 

Rhenium grades of porphyry copper deposits are low, generally less than 0.5 gram per metric ton (g/t) (table P1; fig. P4; 

Sinclair and others, 2009; John and Taylor, 2016), but the 

large tonnage of ore processed, the presence of sufficient 

molybdenite to make molybdenum recovery economic, and 

the presence of specialty circuits that allow recovery of the 

rhenium contained in molybdenite means that a relatively 

large proportion of rhenium production is from these deposits.

Rhenium contents of molybdenite in porphyry molybdenum deposits are generally much lower than in molybdenite 

in porphyry copper deposits (table P2; Giles and Schilling, 

1972; Sinclair and others, 2009; John and Taylor, 2015). Thus, 

despite their higher molybdenum grades, rhenium grades of 

porphyry molybdenum deposits are generally less than those 

of porphyry copper deposits (table P2; fig. P4). The relatively 

low grades and small sizes of most porphyry molybdenum 

deposits result in small rhenium resources (table P2; fig. P4), 

and rhenium is not recovered from these deposits.

Rhenium is recovered from molybdenite concentrates that 

are separated from copper-iron sulfides by flotation methods. 

During roasting of the molybdenite concentrates to produce 

molybdenum oxide, rhenium is oxidized to Re2O7 and passes 

up the flue stack with the sulfur gases. When the flue dusts 

and gases are scrubbed, rhenium is dissolved in the resulting 

sulfuric acid and is eventually precipitated out as ammonium 

perrhenate (NH4ReO4) (Sutulov, 1974; Nadler, 1997). 

Ammonium perrhenate is a white crystalline powder that 

contains about 69 weight percent rhenium and is the principal 

form in which rhenium is marketed (fig. P1C).

Rhenium in Sediment-Hosted Strata-Bound 

Copper Deposits

Sediment-hosted strata-bound copper deposits are the 

other major primary source of rhenium. Rhenium is recovered 

during processing of copper ores from sandstone-type (red 

bed) copper deposits in Kazakhstan and from reduced-facies 

(Kupferschiefer)-type copper-silver deposits in Poland. 

Sediment-hosted strata-bound copper mineralization consists 

of fine-grained copper and copper-iron sulfide minerals in 

strata-bound-to-stratiform disseminations in siliciclastic or 

dolomitic sedimentary rocks (Hitzman and others, 2005; 

Zientek and others, 2013). Ore minerals are characteristically 

zoned laterally and vertically across bedding from hematite to 

chalcocite to bornite to chalcopyrite to pyrite in the inferred 

direction of fluid flow (figs. P5B and P5C ). Deposits are hosted 

in reduced sedimentary strata within or above a thick section 

of oxidized red beds. These deposits are inferred to form by the 

flow of oxidized, metal-rich brines through permeable red-bed 

sedimentary rocks (mostly sandstones). Interaction of the 

oxidized ore fluids with a reducing agent, such as hydrocarbons 

(oil or natural gas), reduced organic carbon, and (or) sulfide 

minerals, results in deposition of the zoned sequence of iron, 

B

SOUTHWEST NORTHEAST

~10 meters

NOT TO SCALE

Rhenium Figure 5B. 

Chalcocite

Bornite

Chalcopyrite

Galena

Sphalerite

Rich chalcocite-bornite 

 mineralization

Red-colored 

 siltstone

Highest Re grades

EXPLANATION

Mineral zone

Rock type

Inferred direction of ore brine movement

Gray sandstone

KAZAKHSTAN

Dzhezkazgan

Location of section

Figure P5—Continued B, Sandstone-type strata-bound copper deposit at Dzhezkazgan, Qaraghandy, 

Kazakhstan, showing mineral zoning and rock type. Rhenium (Re) is enriched in all the ore zones, but the 

highest concentrations are thought to be in the copper zones, especially in chalcocite-bornite ore (Seyfullin 

and others, 1974). Figure modified from Box and others (2013) based on Daukeev and others (2004). 
Geology P13

Figure P5.—Continued C, Kupferschiefer (reduced-facies-type) strata-bound copper deposits showing lithology 

(top cross section) and copper ore distribution (stippled pattern) and metal and mineral zoning (bottom cross section). 

Rhenium is concentrated in copper ore in the Kupferschiefer black shale. Transitions between predominant sulfide 

minerals are gradational. Figure is based on Jowett (1986) and Oszczepalski (1999). 

C

Ore fluids

Ore fluids

Pyrite

Rote Faule

Transition zone

Rote Faule

Hematite

Bornite

Covellite

Chalcocite

Chalcopyrite

Sphalerite

Kupferschiefer

Re-enriched Cu ore in Kupferschiefer black shale

Galena

Break in section

Anhydrite

Limestone

Shale

Sandstone

Rhenium enrichment

Unaltered anhydrite 

 and limestone

Zinc

Lead

Rock type Metal zone

EXPLANATION

Copper

Copper ore zone

Red rock caused by

 hematite replacement

 of pyrite (Rote Faule)

Unaltered sandstone

VERTICAL SCALE GREATLY EXAGGERATED

1 KILOMETER

1 MILE

0

Lower 

anhydrite

Zechstein

limestone

Weissliegendes

Sandstone

Rotliegendes

Sandstone

Kupferschiefer

Rhenium figure 5C. 

WEST EAST

WEST EAST

Break in section

Basal limestone
P14 Critical Mineral Resources of the United States—Rhenium

copper, lead, and zinc sulfide and oxide minerals. Unlike 

porphyry copper deposits, in which rhenium is clearly associated with molybdenite and is derived from magmatic sources, 

the site and source of rhenium in sediment-hosted strata-bound 

copper deposits are poorly understood.

Occurrences in Sandstone-Type Strata-Bound 

Copper Deposits

In central Kazakhstan, rhenium is recovered from Late 

Paleozoic sandstone-type strata-bound copper deposits in 

the Chu-Sarysu basin (Dzhezkazgan and Zhaman-Aybat 

[also spelled Zhaman-Aibat] deposits) (fig. P3; Box and 

others, 2013). The deposits consist of copper sulfide minerals 

(bornite, chalcocite, and chalcopyrite) that form intergranular 

cement and replace grains in sandstone and conglomerate 

within a 600- to 1,000-m-thick Pennsylvanian fluvial red-bed 

sequence. Copper grades of individual deposits range from 

about 0.8 to 1.7 percent and average 1.1 percent, and the 

deposits contain significant quantities of silver and rhenium. 

The copper minerals are inferred to have precipitated from 

an oxidized, metal-rich brine that was progressively reduced 

by interaction with hydrogen sulfide (H2S)-bearing petroleum 

fluids as it flowed through permeable sandstone beds. The 

copper minerals are zoned from chalcocite to bornite to 

chalcopyrite in the inferred direction of the fluid (brine) flow 

(fig. P5B). The main deposits in Kazakhstan from which 

rhenium is recovered are the Zhaman-Aybat deposit (which 

has an average rhenium grade of 1.45 g/t and includes local 

zones that exceed 9 g/t), and the Dzhezkazgan deposit (which 

has an average grade of 1 to 2 g/t rhenium). Both deposits 

are located in Qaraghandy Province. In the Zhaman-Aybat 

deposit, the highest rhenium grades (>9 g/t) are associated 

with the boundary between the chalcopyrite and the bornitechalcocite zones in more central parts of the deposit.

The occurrence and mineralogy of rhenium in the 

Kazakhstan deposits is uncertain. Satpaeva and others (1959, 

summarized in Fleischer, 1960) and Seyfullin and others 

(1974) reported rhenium contents of different types of ores 

at Dzhezkazgan and other strata-bound copper deposits in 

Kazakhstan. Their data show that rhenium is present in appreciable abundances in lead-rich and zinc-rich ores, as well as in 

copper-rich ores, and that rhenium apparently is not associated 

with molybdenite. In particular, Seyfullin and others (1974) 

showed that the ratio of molybdenum to rhenium (Mo:Re) in 

copper ore averaged 1.2 in the Dzhezkazgan deposit and varied 

from 0.4 to 75 in other deposits in Kazakhstan, which indicates 

significant enrichment of rhenium relative to molybdenum in 

these deposits. For comparison, Mo:Re ratios in porphyry copper 

deposits range from about 100 to >10,000 and average about 

2,050 (table P2), which is close to estimates for the average 

Mo:Re ratios in continental crust, which range from 2,500 to 

2,750 (Taylor and McLennan, 1995; Rudnick and Gao, 2003).

Poplavko and others (1962) reported discovery of 

dzhezkazganite, a new rhenium mineral, at the Dzhezkazgan 

deposit and suggested that it had a chemical formula of CuReS4

. 

Subsequent electron microprobe analyses suggested a more 

complex formula of ReMoCu2PbS6 for this microscopic 

(typically <1 μm) phase (Genkin and others, 1994). Recent 

analyses of copper ore minerals in two samples from the 

Dzhezkazgan deposit show that the rhenium contents of 

chalcopyrite and bornite range from 3.3 to 10.1 ppm (Box 

and others, 2013), which are extremely high concentrations of 

rhenium in nonmolybdenite samples (Ruiz and Mathur, 1999; 

Selby and others, 2009). These data suggest that significant 

amounts of rhenium may be present in copper-iron sulfide 

minerals, although mass balance relations between copper and 

rhenium indicate that most of the rhenium is likely present 

in dzhezkazganite.

Occurrences in Reduced-Facies-Type (Kupferschiefer) 

Strata-Bound Copper Deposits

The thin black shale that forms the Upper Permian 

Kupferschiefer (copper slate) hosts the largest copper and 

silver deposits in Europe (Mansfeld-Sangerhausen and 

Lubin-Sieroszowice deposits) (fig. P3; Vaughan and others, 

1989; Jankowski, 1995; Kucha, 2003; Hitzman and others, 

2005). These reduced-facies-type strata-bound copper deposits 

also contain large rhenium resources. In the Kupferschiefer, 

rhenium is currently produced from the Lubin-Sieroszowice 

orebody in southwestern Poland (Bartlett and others, 2013), 

and high rhenium contents have been reported in the formerly 

mined Mansfeld-Sangerhausen deposits in southern Germany 

(Jankowski, 1995). Ore reserves in these deposits are large; 

2011 end-of-year reserves for the Legnica-Glogów copper belt 

in Poland, which includes the Lubin-Sieroszowice orebody, 

were 1,181 million metric tons of ore at average grades of 

1.58 percent copper and 48 g/t silver (Bartlett and others, 

2013); large undeveloped deposits have also been identified 

elsewhere in the Kupferschiefer in Poland. The average 

rhenium content of the Polish copper ores mined in 2009 was 

0.6 g/t, but the rhenium grade varied from about 1.1 g/t in 

shale ore to about 0.4 g/t in sandstone ore (Smakowski and 

others, 2010). In contrast, Jankowski (1995) reported a much 

larger average rhenium content of 21 ppm in Kupferschiefer 

ores in the Mansfeld-Sangerhausen deposits in Germany, 

suggesting that more than 2,500 metric tons of rhenium was 

contained in ores mined from 1200 through 1990. Only small 

amounts of rhenium were recovered from flue dusts near the 

end of mining of these deposits, however, and production data 

from 1962 suggests that rhenium recovery was only about 

0.1 g/t (Kruger, 2006).

The Kupferschiefer deposits occur along the southern 

margin of the Rotliegendes-Zechstein basin (part of an Early 

Permian intracontinental rift basin) where it overlies Paleozoic 

continental sedimentary and bimodal volcanic rocks (Vaughan 

and others, 1989; Kucha, 2003; Hitzman and others, 2005). 

The volcanic rocks in the Lower Rotliegendes are extensively 

albitized and may have supplied some of the copper in the 

Kupferschiefer deposits (Hitzman and others, 2005). The 

overlying Upper Rotliegendes is a thick red-bed sequence 
Geology P15

composed of fluvial, eolian, and playa (sabkha) deposits that is 

locally overlain by white eolian sandstones (Weissliegendes). 

The Upper Permian Kupferschiefer is a thin (typically 

30 to 70 centimeters thick) organic carbon-rich black shale 

that covers more than 800,000 square kilometers (km2) in 

central Europe. It forms the basal unit of the Zechstein layer 

and was deposited during the initial stages of a rapid marine 

transgression over the Rotliegendes-Weissliegendes sequence. 

The Kupferschiefer black shale is overlain by Zechstein 

carbonates (dolomite and limestone) and evaporites (anhydrite 

and halite), which may have supplied some of the sulfur in the 

ores. Copper, gold, PGEs, and silver are hosted by both the 

black shales and the underlying white sandstones (fig. P5C). 

Thin base-metal sulfide zones in the Kupferschiefer are tens 

to hundreds of square kilometers in area, and immediately 

adjacent hematitic alteration (in the Rote Fäule) covers more 

than 40,000 km2.

In the Kupferschiefer deposits, hydrothermal minerals 

and metals typically are zoned upward from hematite (Fe3+) 

to the sequence of covellite, chalcocite, bornite, chalcopyrite 

(copper zone), to galena (lead zone), to sphalerite (zinc zone), 

and finally to pyrite (Fe2+) (fig. P5C; Oszczepalski, 1999; 

Kucha, 2003; Hitzman and others, 2005). Gold and PGEs 

are concentrated in the transition zone between hematite and 

copper minerals, and silver is concentrated in the copper 

ores. The Kupferschiefer mineralization is complex, probably 

formed in multiple stages, and likely involved influx of 

oxidizing, metal-rich brines that moved through the oxidized 

red beds (Rote Fäule) from the Rotliegendes basin into the 

reduced Kupferschiefer (Zechstein) sediments (fig. P5C; see 

summary in Hitzman and others, 2005). The three main stages 

of ore formation involved (1) early diagenetic formation of iron 

monosulfides and possibly chalcocite near the sediment-water 

interface; (2) diagenetic replacement of iron monosulfides and 

(or) pyrite by copper-iron sulfides at least partly associated 

with Rote Fäule alteration, and (3) diagenetic remobilization 

of earlier copper sulfide mineralization during later Rote 

Fäule alteration and formation of gold-PGE mineralization 

(for example, Vaughan and others, 1989; Wodzicki and 

Piestrzynski, 1994; Kucha, 2003; Hitzman and others, 2005).

Rhenium in the Kupferschiefer is associated with 

molybdenum in copper-rich shale ore (Hammer and others, 

1990; Kucha, 2003). Kucha (2003) stated that copper, lead, 

and potassium-castaingite (cuprian molybdenite) have 

molybdenum-to-rhenium ratio of 70:1, although he did not 

report actual rhenium analyses. Jankowski (1995) reported 

molybdenum-to-rhenium ratio of 7:1 for ores in the MansfeldSangerhausen mines but did not report rhenium mineralogy. 

As in the Kazakhstan strata-bound copper deposits, these 

molybdenum-to-rhenium ratios suggest significant rhenium 

enrichment relative to molybdenum.

Geochemical studies of the Sangerhausen basin in 

Germany show large variations in the rhenium content of 

rocks; the content ranges from 0.24 to 27 ppm, and the 

highest rhenium concentrations are in the copper facies 

of the Kupferschiefer (Hammer and others, 1990). Pašava 

and others (2010) reported rhenium concentrations ranging 

from 63.6 to 1,380 ppb in six samples of unmineralized 

Kupferschiefer black shale in Poland (the copper content was 

less than or equal to 106 ppm and the molybdenum content 

was between 20 and 340 ppm). In these samples, rhenium 

strongly correlates with molybdenum (r2=0.93) and is 

concentrated in the lower part of the black shale. Pašava and 

others (2007b) reported rhenium concentrations ranging from 

249 to 22,000 ppb for six copper-rich black shale samples (the 

average copper grade was 4.9 percent and no molybdenum 

content was reported), whereas six samples from PGE- and 

gold-rich but copper-poor horizons had rhenium concentrations that ranged from 221 to 558 ppb (Pašava and others, 

2007a). Comparison of the minor element composition of 

unmineralized Kupferschiefer sediments in the Lower Rhine 

basin in northwestern Germany to black shales in the Late 

Devonian Exshaw Formation in Canada and to modern black 

shales indicate similar high levels of enrichment in arsenic, 

cadmium, molybdenum, nickel, rhenium, antimony, uranium, 

and vanadium, which are typical of syngenetic mineralization 

(Lüschen and others, 2000a, b). These data suggest that 

rhenium in Kupferschiefer ores could have been derived from 

the black shales and was locally remobilized into the copperrich zones.

Rhenium is produced from the Kupferschiefer deposits 

by roasting copper ore. During roasting, sulfur and rhenium 

are released as gases that precipitate in the flues. Wet scrubbing of the flue dusts produces sulfuric acid that contains 

dissolved rhenium. Resins are then used to adsorb the rhenium 

from the acid, and the rhenium is ultimately precipitated as 

ammonium perrhenate. All processing of Kupferschiefer ores 

and recovery of rhenium are done by KGHM Polska Miedź 

S.A. at their facilities in Lubin, Poland.

Rhenium in Sandstone-Type Uranium Deposits

Rhenium was formerly produced as a byproduct from 

roll-front-type sandstone uranium deposits in Uzbekistan, and 

rhenium is currently recovered from similar uranium deposits 

in Kazakhstan (figs. P3 and P5D; Dahlkamp, 2009a, b; 

Seltmann and others, 2012). Small amounts of rhenium were 

also recovered from sandstone uranium deposits in Texas from 

1969 to 1974 (Millensifer, 1997). Sandstone uranium deposits 

are generally hosted in permeable medium- to coarse-grained 

sandstones that were deposited in fluvial or nearshore marine 

environments. The large sandstone deposits in Kazakhstan and 

Uzbekistan are hosted in large Cretaceous to early Tertiary 

continental basins. The deposits form by the dissolving of 

uranium from nearby strata in oxidized groundwaters that 

flow through the sandstones. Uranium is precipitated at a 

redox boundary (the “roll-front”) when the oxidized ore fluids 

intersect and react with reducing agents in sandstone host 

rocks, such as carbonaceous (plant and algal) material, sulfide 

minerals, hydrocarbons, or interbedded volcanic rocks. Rollfront-type uranium deposits are crescent-shaped orebodies that 

crosscut the bedding in the sandstones (fig. P5D).
P16 Critical Mineral Resources of the United States—Rhenium

Figure P5.—Continued D, Roll-front-type sandstone uranium deposit in the Moynkum-Tortkuduk sector of the 

Chu-Sarysu basin, Kazakhstan, showing lithological and reduction-oxidation (redox) control on uranium, rhenium, 

and selenium mineralization. Rhenium and uranium are enriched at the front of the roll in reduced rocks, whereas 

selenium is concentrated in oxidized rocks. The top cross section shows the oxidized zone; the bottom cross 

section shows the zone of selenium enrichment. Figure modified from Dahlkamp (2009a). ppm, part per million

Medium-grained sandstone

Fine-grained sandstone

Clay and siltstone

Oxidized zone

Uranium orebody

Uranium halo

Rhenium halo—Concentration

 

 >0.1 ppm

Inferred direction of ore 

 fluid movement

Rhenium figure 5D.

EXPLANATION

Land

surface

10

20

30

40

0 40

VERTICAL EXAGGERATION ×4

80 120 METERS

0 100 200 300 FEET

METERS

Land

surface

10

20

30

40

METERS

Selenium halo—Concentration

 >50 ppm

KAZAKHSTAN

Moynkum

Location of section

D
Geology P17

At the Sugraly deposit in the Zarafshan mining district in 

Uzbekistan, rhenium was recovered along with selenium and 

molybdenum as byproducts of the mining of uranium deposits 

in Cretaceous to Eocene sandstone strata in the Kyzylkum 

basin (Dahlkamp, 2009b). Underground mining began in 1977 

and was later switched to in situ leaching until mining ended 

in 1994. The Sugraly deposit was a typical roll-front-type 

deposit with complex uranium-selenium-molybdenumrhenium ores that formed orebodies that were from several 

hundred meters to 20 km long and from 100 to 500 meters (m) 

wide. Rhenium content varied from 10 to 15 ppm (Seltmann 

and others, 2012). Rhenium was reportedly present in the form 

of ReS2 and ReO2 (Dahlkamp, 2009b).

Rhenium is reported in several roll-front-type sandstone 

uranium deposits in Kazakhstan (Dahlkamp, 2009a). These 

include deposits in the Moynkum (Moinkum), the Kanzhugan, 

and the Kenze-Budenovskaya mining districts in the 

Chu-Sarysu basin, and deposits in the Suluchekinskoye district. 

In the Moynkum and Kanzhugan districts, numerous uranium 

orebodies are associated with redox fronts in Paleogene 

sandstone aquifers. Rhenium grades generally range from 

0.08 to 0.38 g/t with rhenium content of as much as 4.8 g/t 

in some deposits. In the Tortkuduk sector at the north end of 

the Moynkum deposit, uranium mineralization is contained in 

a 40-m-thick arenite unit sandwiched between clay-silt beds 

(fig. P5D). Uranium mineralization extends for about 10 km 

along a winding redox front. The forward sections of the 

rolls are about 20 m thick; tails may be up to several meters 

thick and more than 400 m long. A rhenium halo ( > 0.1 ppm 

rhenium), which coincides with the uranium distribution, 

surrounds the front of the roll and extends for up to 60 m into 

the reduced rocks (fig. P5D). A selenium zone with greater 

than 50 ppm selenium is in the rear part of the roll (fig. P5D).

At the Zhalpak and the Akdala deposits in the KenzeBudenovskaya mining district, Cretaceous sandstones host 

roll-front uranium mineralization. Several orebodies in 

these deposits have rhenium content of as much as 62 ppm 

(Dahlkamp, 2009a; Seltmann and others, 2012). Samples 

containing greater than 0.2 ppm rhenium have higher uranium 

contents. In the Suluchekinskoye district in the Ily basin, rollfront-type orebodies are hosted by Cretaceous to Paleogene 

sandstones. Orebodies are zoned laterally from rhenium in 

reduced rocks, followed by a uranium-rhenium zone at the 

redox boundary, and then a selenium zone that forms the rear 

of the roll in oxidized rocks. Rhenium contents range from 

1 to 24 ppm and average 1 to 2 ppm.

Uranium and rhenium in the Kazakhstan deposits are 

recovered using in situ leaching (also known as solution 

mining). In mining that uses in situ leaching, ore minerals are 

dissolved by circulating groundwater fortified with sulfuric or 

other acids through undisturbed underground deposits (World 

Nuclear Association, 2013). The resulting pregnant solutions 

are pumped to the surface, and uranium, rhenium, and other 

metals are recovered using methods similar to those used in 

processing milled uranium ore.

Rhenium in the Merlin Deposit

The Merlin molybdenum-rhenium deposit in the Mount 

Isa Inlier in northwestern Queensland, Australia, is a highgrade rhenium resource in which rhenium and molybdenum are 

the primary commodities and only minor amounts of copper 

are present (fig. P3; Brown and others, 2010; Lycopodium 

Minerals QLD Pty Ltd., 2012). Mineralized rock consists 

dominantly of molybdenite in veins, infilling breccias and 

stylolites, and disseminated grains in a zone that is about 1 km 

long and up to 20 m thick. Rhenium is contained in molybdenite grading an average of about 1,000 ppm rhenium. In 

2012, indicated and inferred mineral resources were 6.9 million 

metric tons averaging 1.38 percent molybdenum and 22.7 ppm 

rhenium (Lycopodium Minerals QLD Pty Ltd., 2012).

The Merlin deposit was discovered in 2008 during 

drilling of the northern extension of the Mount Dore 

copper deposit, which mostly lies above and south of the 

Merlin deposit (Brown and others, 2010). The Merlin and 

the Mount Dore deposits are hosted by Early Proterozoic 

carbonaceous metashale and metasiltstone that are structurally 

overlain by the 1,500-mega-annum (Ma) anorogenic Mount 

Dore Granite along a moderately dipping reverse fault. The 

Merlin molybdenum-rhenium mineralization is associated 

with silica-albite alteration and interstitial clays that formed 

along reactivated fractures and shear zones and replaced the 

matrix of structurally controlled breccias. At Mount Dore, 

early regional-scale sodic-calcic alteration is cut by potassic 

alteration and quartz veins and by breccias that host an early 

phase of copper mineralization. A second phase of brecciation 

was followed by deposition of dolomite and a second stage of 

copper mineralization. The molybdenum-rhenium mineralization at the Merlin deposit cuts the copper mineralization in 

the Mount Dore deposit but is likely part of the same system. 

Zircon uranium-lead dating of the Mount Dore Granite and 

rhenium-osmium dating of molybdenite in the Mount Dore 

deposit show that both formed at about 1,500 Ma and suggest 

that copper-molybdenum mineralization at Mount Dore 

probably is related to a late-stage, evolved magmatic fluid 

that generated potassic alteration (Duncan and others, 2011). 

The average Mo:Re ratio of 600 is also consistent with an 

igneous origin.
P18 Critical Mineral Resources of the United States—Rhenium

Resources and Production

Identified Resources

Estimates of grades and tonnages of major rheniumbearing deposits are shown in figure P4 and presented in 

tables P1 and P2. There are no published grade and tonnage 

models for rhenium in porphyry deposits, although Sinclair 

and others (2009) and Millensifer and others (2014) present 

plots of rhenium grade versus deposit tonnages for some 

porphyry and other types of deposits. Publicly available data 

on rhenium contents in porphyry deposits are limited and often 

inconsistent when more than one data source is available. Very 

few grade and tonnage data are available for other types of 

deposits that produce rhenium. Descriptions of data compilation methods and limitations of these data are in appendix P1.

Production

Worldwide mine production of rhenium in 2012 was 

estimated to be 52,600 kg, of which about 27,000 kg, or 

51 percent, was produced from porphyry copper mines in 

Chile (fig. P6; Polyak, 2014). Because specialized processing 

facilities are required to recover rhenium from molybdenite 

concentrates, however, a significant amount of rhenium 

(approximately 15 to 20 metric tons) is contained in molybdenite concentrates roasted at facilities other than where it 

is mined (for example, molybdenite concentrates from the 

Bingham Canyon Mine), thereby making estimates of the 

sources of rhenium difficult. In recent years, rhenium also was 

recovered from porphyry copper deposits in the United States, 

Armenia, Kazakhstan, Mexico, Peru, Russia, and Uzbekistan. 

Porphyry copper deposits formed in continental arcs are the 

main sources of both rhenium production and resources (tables 

P1 and P2; fig. P3). All primary rhenium production in the 

United States is from processing facilities at the Sierrita Mine 

in Arizona, where rhenium is recovered from molybdenite 

concentrates produced at several porphyry copper mines. 

Rhenium resources in the United States are located mainly in 

Arizona and Utah, although smaller resources are located in 

Montana, New Mexico, and Nevada. The Pebble deposit in 

Alaska also contains a large inferred rhenium resource.

Sediment-hosted strata-bound copper deposits in Poland 

and Kazakhstan were the other major primary sources of 

rhenium in 2011, and both the Kupferschiefer deposits in 

Poland and the Dzhezkazgan deposit in Kazakhstan have 

large identified rhenium resources (fig. P4). Ore reserves 

in the underground deposits in Kupferschiefer deposits in 

the Legnica-Glogów copper belt, Poland, are sufficient to 

maintain the current production rate of about 30 million metric 

tons per year for 30 to 40 years (Bartlett and others, 2013).

Sandstone uranium deposits in Kazakhstan produce 

some byproduct rhenium, and published grade and tonnage 

United States

15%

World mine production—52,600 kilograms

Poland

12%

Kazakhstan

6%

Uzbekistan

10%

Armenia

1%

Russia

3%

Others

2%

Chile

51%

Figure P6. Pie chart showing world rhenium mine production in 

2012, by country and percent of world total. Compiled using data 

from Polyak (2014).

data suggest significant rhenium resources in some deposits 

(Dahlkamp, 2009a). The lack of grade and tonnage data for 

most deposits, however, precludes a thorough assessment 

of these resources (table P1; fig. P4). Similar deposits in the 

southern and middle Ural Mountains in Russia may produce 

rhenium, but no production or resource estimates are available 

(Khalezov, 2009).

The Merlin molybdenum-rhenium deposit in Queensland, 

Australia, is unique because rhenium is a primary commodity. 

Despite its high rhenium grade (22.7 ppm), the contained 

rhenium resources are relatively small compared with the 

resources of other deposit types (fig. P4).

Undiscovered Resources

Rhenium resources likely are present in undiscovered 

porphyry copper deposits formed in continental arcs 

throughout the world. Recent U.S. Geological Survey 

(USGS) assessments of porphyry copper deposits in the 

Southwestern United States (Ludington and others, 1996), 

the Andes Mountains in South America (Cunningham and 

others, 2008), Mexico (Hammarstrom and others, 2010), 

the Province of British Columbia and the Yukon Territory in 

Canada (Mihalasky and others, 2013), and the Tethys region, 

including Armenia, Azerbaijan, Georgia, Iran, and Turkey 
Environmental Considerations P19

(Lukas Zürcher, U.S. Geological Survey, written commun., 

2013) concluded that there is significant potential for undiscovered porphyry copper deposits in these regions. Because 

known porphyry copper deposits in these areas contain some 

of the largest identified rhenium resources in the world, it is 

highly probable that significant rhenium resources remain to 

be discovered in these areas.

In the United States, additional rhenium resources in 

porphyry copper deposits are most likely to occur in Arizona, 

where there are numerous identified deposits—several of 

which contain recoverable rhenium—and where there is 

a high probability of undiscovered deposits. Also, recent 

delineation of the giant Pebble porphyry copper-molybdenumgold deposit in Alaska, which has estimated rhenium resources 

that represent more than 40 years of production at the current 

level of worldwide mine production (Ghaffari and others, 

2011; Lang and others, 2013), suggests that there is the 

potential for significant rhenium resources in undiscovered 

porphyry copper deposits in Alaska.

Mongolia is another area where significant porphyry 

copper deposits have been discovered recently (for example, 

the giant Oyu Tolgoi porphyry copper-gold deposit in 

Omnogovi Aymag) (Khashgerel and others, 2006). The 

incompletely defined Zuun Mod molybdenum deposit has 

modest rhenium resources (table P1; Clark and Baudry, 2011). 

The presence of these deposits suggests the potential for 

significant undiscovered rhenium resources in Mongolia.

According to the recent USGS assessment of sedimenthosted strata-bound copper deposits, a significant number 

of undiscovered reduced-facies-type strata-bound copper 

deposits that have large resources are likely present in the 

Kupferschiefer in Poland (M.L. Zientek, U.S. Geological 

Survey, written commun., 2013) and sandstone-type stratabound copper deposits in Kazakhstan (Box and others, 

2012). Because known deposits in these areas are enriched in 

rhenium and contain significant rhenium resources (fig. P4), 

the undiscovered deposits also likely contain significant 

rhenium resources. Other large sediment-hosted strata-bound 

copper deposits, such as Udokan in the Transbaikalia Region 

in Russia and deposits in the Democratic Republic of the 

Congo, do not have reported anomalous rhenium contents 

(Hitzman and others, 2005; Zientek and others, 2013).

The recent discovery of the Merlin deposit in Australia 

opens up the possibility that there are other high-grade 

molybdenum-rhenium deposits in which these metals are the 

primary commodities. Until the genesis of the Merlin deposit 

is better understood, however, it is premature to speculate on 

the probability of the occurrence of similar deposits.

Exploration for New Deposits

With the exception of the Merlin molybdenum-rhenium 

deposit, all identified rhenium resources are contained in 

deposits mined primarily for other mineral commodities. In 

porphyry copper deposits, rhenium is recovered mostly from 

molybdenum, which is another byproduct. Until a better 

understanding of the genesis of the Merlin deposit emerges 

and it is determined that this type of deposit is economic, 

exploration for other primary rhenium deposits is unlikely 

to take place. Therefore, the discovery of significant new 

rhenium resources is likely to depend on successful exploration for porphyry copper deposits, especially in frontier areas, 

such as Alaska and Mongolia, where there have been recent 

discoveries of large porphyry copper deposits that may contain 

byproduct molybdenum and rhenium (for example, the Pebble 

and the Oyu Tolgoi deposits). Future exploration for porphyry 

copper deposits likely will involve increased remote sensing, 

geophysical (both potential field and electrical techniques), 

and lithogeochemical studies to discover covered deposits 

(for example, Graybeal and Vikre, 2010).

Environmental Considerations

Rhenium is present in Earth’s crust at very low concentrations, and its production is nearly always as a byproduct 

commodity from copper mining. Therefore, the environmental 

characteristics of rhenium extraction are closely linked to those 

associated with the mining of the two types of copper deposits 

that represent its main sources. The primary source of rhenium 

is from the mineral molybdenite (the main ore mineral of 

molybdenum) derived from porphyry copper deposits. Rhenium 

is also produced from sediment-hosted copper deposits. The 

economic geology of the two mineral deposit types that are the 

predominant sources of byproduct rhenium are significantly 

different from one another, which means that the environmental 

geology of these deposit types are also different. The environmental geology of porphyry copper deposits is dominated by 

their large size (which means that they are commonly mined by 

open pit methods), their low grade, and the variable potential 

of their solid mine wastes to generate minor amounts of acid 

drainage or neutralize it (John and others, 2010). The environmental geology of sediment-hosted copper deposits is characterized by their moderate size, the tabular geometry of their ores, 

their low grade, and the low potential of their solid mine

wastes to generate acid drainage (Hayes and others, 2015).
P20 Critical Mineral Resources of the United States—Rhenium

Sources and Fate in the Environment

Concentrations of rhenium in the natural environment, 

including in groundwater, surface water, sediment, soils, 

and biota, are low, which is a reflection of rhenium’s low 

crustal abundance (table P3). The distribution of rhenium in 

environmental samples is further influenced by its relatively 

higher solubility in oxygenated waters and its lower solubility 

in oxygen-free (anoxic) waters.

The overall concentration of rhenium in water is similarly 

lower than that of many other base and precious metals, as 

summarized by Colodner and others (1993) and Hodge and 

others (1996). In fact, the natural concentration of rhenium in 

surface water, groundwater, and seawater falls in the partsper-trillion range (table P3). Experimental studies suggest that 

the maximum solubility of ReO2 in water at room temperature 

(25 °C) is less than 160,000 parts per trillion (ppt) (0.16 ppm), 

but could be as low as 20 ppt (Kim and Boulègue, 2003; 

Xiong and others, 2006). Dissolved concentrations in the 

ocean average 8.3 ppt and do not vary with depth (Colodner 

and others, 1993). Variations in rivers span four orders of 

magnitude, from 0.004 to 76 ppt (Colodner and others, 1993; 

Rahaman and others, 2012). The geology of the watershed 

influences the rhenium concentration. Watersheds underlain by 

black shales typically have higher concentrations of rhenium 

than do those underlain by other rock types. Rhenium in 

seawater and surface water behaves conservatively, meaning 

that it stays dissolved and does not bind or adsorb to particulate matter, such as clays. Reducing sediments (sediments 

rich in organic matter and sulfide minerals) are known to be 

important for the removal of rhenium from the water column, 

however (Koide and others, 1986). The observation that 

black shales can serve as sources of elevated concentrations 

of rhenium in surface water is a reflection of this process 

operating in the geologic record (Rahaman and others, 2012).

In solid environmental media (soil, sediment, and 

biological material), the rhenium concentrations are also low, 

although they are higher than those typically found in water 

(table P3). River sediments, oxygenated marine sediments, 

and soils all have concentrations that are comparable to the 

average crustal abundance (0.4 ppb). Organic-rich, anoxic 

sediments have concentrations that can be significantly higher 

(1.8 to 110 ppb) (Koide and others, 1986). Plants are known to 

accumulate rhenium at concentrations greater than local soil 

concentrations (Bozhkov and Borisova, 2003).

Information on pre-mining environmental concentrations 

of rhenium is limited, in part because of its low concentrations. Leybourne and Cameron (2008) investigated the 

concentrations of arsenic, copper, molybdenum, rhenium, and 

selenium in groundwater associated with the undeveloped 

Spence porphyry copper deposit in the Atacama Desert in 

northern Chile. They found that groundwater concentrations 

within the deposit reached a maximum rhenium concentration 

of 30.7 ppb. Outside of the deposit, the minimum rhenium 

concentration was as low as 0.2 ppb. From an environmental 

perspective, baseline concentrations of other trace elements 

and related constituents associated with porphyry copper 

deposits or sediment-hosted copper deposits are of greater 

environmental interest than are those for rhenium because 

these elements can be present in significantly higher 

concentrations.

Environmental baseline characterization studies of 

porphyry copper deposits are limited in the literature, 

particularly with regard to the diversity of climatic settings in 

which the deposits are found. Some information is available 

in mine permit applications for recently proposed mines, 

however. Studies from such diverse climates as the cold 

climate of southwestern Alaska (Fey and others, 2008); 

the tropical climate of Puerto Rico (Learned and Boissen, 

1973; Plaza-Toledo, 2005); and the hot, arid climates of 

Table P3. Rhenium concentrations in rocks, soils, biota, waters, and sediments.

[ppb, part per billion; ppt, part per trillion]

Environment Rhenium

concentration Reference(s)

Upper continental crust 0.4 ppb Taylor and McLennan (1995)

Soils 0.208 to 1.72 ppb Uchida and others (2005)

Leaves, dried, birch and fir, unimpacted 4 to 6 ppb Bozhkov and Borisova (2003)

Leaves, dried, acacia, smelter and mine site vicinity 902 to 2,430 ppb Bozhkov and Borisova (2003)

Seawater (Atlantic Ocean and Pacific Ocean) 8.3±0.1 ppt Colodner and others (1993)

Rivers 0.004 to 76.7 ppt Colodner and others (1993); Rahaman and others (2012)

Groundwater 0.9 to 35.4 ppt Colodner and others (1993); Hodge and others (1996)

Groundwater, porphyry copper deposits 0.09 to 30.7 ppb Leybourne and Cameron (2006)

River sediment 0.233 to 0.285 ppb Uchida and others (2005)

Pelagic sediments <0.1 ppb Koide and others (1986)

Anoxic sediments 1.8 to 110 ppb Koide and others (1986)
Environmental Considerations P21

Arizona and Chile (Chaffee, 1976, 1977; Chaffee and others, 

1981; Leybourne and Cameron, 2006, 2008) show similar 

geochemical features. Soils show elevated concentrations 

of, in decreasing order of maximum concentration, iron 

(2.3 to 7.4 weight percent), sulfur (0.02 to 0.87 weight 

percent), copper (3.2 to 1,830 ppm), zinc (36 to 142 ppm), 

arsenic (7 to 78 ppm), molybdenum (0.6 to 27.1 ppm), 

lead (7.1 to 17.4 ppm), and cadmium (< 0.1 to 0.6 ppm). 

Stream sediments show elevated concentrations of, 

in decreasing order of maximum concentration, iron 

(6.3 to 10.6 weight percent), copper (64 to 804 ppm), 

zinc (59 to 291 ppm), lead (< 4 to 17 ppm), molybdenum 

(< 2 to 10 ppm), arsenic (< 10 ppm), and cadmium (< 2 ppm).

Only a limited amount of pre-mining baseline data 

are available in the literature for soil, stream sediment, 

groundwater, and surface water from a variety of deposits. 

Plaza-Toledo (2005) found that surface waters downstream 

from undeveloped porphyry copper deposits in the Cordillera 

Central of Puerto Rico contained maximum dissolved 

concentrations of 110 ppm sulfate, 15.9 ppm iron, 0.56 ppm 

aluminum, 0.13 ppm copper, and 0.04 ppm zinc. The pH 

was high (7.7 to 8.6). Fey and others (2008) documented 

surface-water pH values of between 4.1 and 7.3 in the vicinity 

of the Pebble deposit in southwestern Alaska. Alkalinity 

ranged between 0 and 100 ppm calcium carbonate (CaCO3 ) 

equivalent; sulfate, between 1 and 85 ppm; and hardness 

of water, between 2 and 130 ppm CaCO3 equivalent. 

Dissolved trace element concentrations showed a range of 

values for, in decreasing order of maximum concentration, 

iron (< 20 to 4,260 ppb), copper (< 0.5 to 688 ppb), zinc 

(< 0.5 to 68 ppb), arsenic (<1 to 36.2 ppb), molybdenum 

(< 2 to 21.9 ppb), lead (< 0.05 to 18.8 ppb), and cadmium 

(< 0.2 to 11.6 ppb). Leybourne and Cameron (2006, 2008) 

documented high salinity in groundwaters associated 

with the undeveloped Spence deposit where the salinity 

reached 10,000 to 55,000 ppm, with one outlier sample at 

145,000 ppm. The pH of the groundwaters varied widely, 

between 4.7 and 9.2. Dissolved sulfate concentrations 

in groundwaters were between 5,000 and 10,000 ppm. 

Dissolved trace element concentrations showed a range of 

values for, in decreasing order of maximum concentration, 

iron (4 to 54,454 ppb), copper (9 to 28,991 ppb), zinc 

(5 to 1,344 ppb), molybdenum (2 to 475 ppb), arsenic 

(<10 to 160.9 ppb), and lead (0.025 to 23.8 ppb).

Environmental baseline characterization studies of 

sediment-hosted copper deposits are more limited than those 

for porphyry copper deposits. For sedimentary-hosted copper 

deposits, soil and stream sediments may contain anomalous 

abundances of copper, lead, silver, and possibly arsenic, 

mercury, and zinc in the vicinity of the deposits (Lindsey 

and others, 1995). Stream-sediment samples collected in 

the Kafue River (Zambia) upstream from sediment-hosted 

copper deposits contained copper in concentrations of 

between 15 and 146 ppm, some of which were just below 

the stream-sediment criterion (that is, the probable effects 

concentration, or the concentration at which toxic effects 

would be expected for aquatic organisms) of 149 ppm 

(MacDonald and others, 2000; Pettersson and others, 2000; 

Pettersson and Ingri, 2001). Also, soil and sediment associated 

with some sandstone deposits in Montana contained anomalous copper (up to 2,000 ppm), lead (greater than 150 ppm), 

and silver (greater than 0.5 ppm), as reported by Cazes and 

others (1981) and Wells and others (1981).

In contrast to some of the sediment and soils that have 

anomalous metal concentrations, surface water in the Kafue 

River upstream from the mining area in Zambia contains 

low levels of copper (up to 3 ppm) and other trace elements 

(Norrgren and others, 2000; Pettersson and Ingri, 2001). The 

draft environmental impact statement for the Montanore 

Project, which is a proposed mine in the Rock CreekMontanore deposit in Montana, describes the surface-water 

quality in streams and lakes near the proposed mine as being 

“excellent,” which presumably means that all trace elements 

meet environmental criteria. The concentrations of total 

suspended solids, total dissolved solids, major ions, nutrient 

concentrations, and metal concentrations are generally low 

and frequently at or below detection limits. The surface waters 

generally contain iron, copper, and silver in low concentrations that are well below environmental guidelines. The 

surface waters tend to be slightly acidic (U.S. Department 

of Agriculture, Forest Service, and others, 2009).

Mine Waste Characteristics

The amount of mine waste associated with porphyry 

copper and sediment-hosted deposits varies with the size of 

the deposit. Porphyry copper deposits range from 30 million 

metric tons to more than 20 billion metric tons with a median 

size of 250 million metric tons. Because copper grades 

are typically less than 1 percent copper (median grade of 

0.44 percent copper), more than 95 percent of the material 

mined ends up as solid mine waste of one form or another. 

The solid mine waste typically falls into one or more of the 

following three types: tailings, waste rock, and leach-pad 

waste. Tailings and waste rock are usually found at all mines, 

whereas leach-pad waste is found only at those mines where 

the ores are amenable to a copper extraction method known as 

solvent extraction-electrowinning (SX–EW). SX–EW is viable 

only for deposits that have been subject to prolonged periods 

of weathering, which has produced a zone near the surface 

where sulfide minerals have been oxidized to copper oxide 

or copper carbonate minerals. The oxide zone will overlie an 

unweathered sulfide zone. The oxide ore is mined and crushed 

to a coarse grain size and placed on lined pads. The piles are 

leached with a sulfuric acid solution, which is recovered to 

collect the leached copper. Oxide zones are not universally 

present at porphyry copper deposits, so this type of mine waste 

is not always present at porphyry copper deposits.
P22 Critical Mineral Resources of the United States—Rhenium

Tailings are produced when the ore is crushed to a sand 

or silt size to facilitate the separation of the copper sulfide 

and other metal-sulfide ore minerals from gangue minerals 

(minerals of no value) using a technique known as froth 

flotation. After separation of gangue minerals, the tailings 

are discarded as a slurry or trucked dry to a tailings storage 

facility. Tailings storage facilities are typically impoundments 

surrounded by a retaining dam.

The other type of solid waste is waste rock. Because of 

their large size and low grade, porphyry copper deposits are 

commonly mined by open pit methods. Waste rock is uneconomic rock that must be removed (stripped) to access the 

ore. Waste rock is disposed of on site. It can also be used for 

construction on site if tests determine that it will not generate 

acid-rock drainage when it is exposed to the atmosphere and 

water. For open pit porphyry copper mines, waste-to-ore 

(stripping) ratios commonly can exceed 2:1, which means 

that for each ton of ore mined, two tons of waste rock must 

be removed (Porter and Bleiwas, 2003). Porphyry copper 

deposits at depth can also be mined by block caving. A vertical 

shaft or spiral decline is built to the base of the orebody, and 

the orebody is mined from below, leaving a large, unsupported 

cavity where the ore was removed. With this type of mining, 

the amount of waste rock is less than that in open pit mining 

and the waste material is dominated by tailings.

The size of sediment-hosted copper deposits depends 

upon their subtype (reduced type, red bed, or sandstone) and 

is smaller than most porphyry copper deposits. The volume 

of mine waste and tailings produced from a deposit depends 

on the deposit size, depth, geometry, ore grade, and mining 

method. Sediment-hosted copper deposits can be mined by 

underground or open pit methods. The reduced-facies-type 

deposits are generally high tonnage (the median tonnage is 

33 million metric tons) with a high copper grade (the median 

grade is 2.3 percent) (Lindsey and others, 1995; Cox and 

others, 2003). Red-bed deposits are generally low tonnage 

(the median tonnage is 1.2 million metric tons), have an 

intermediate copper grade (the median grade is 1.2 percent), 

and are mined either by open pit or underground methods. 

Sandstone deposits are intermediate tonnage (the median 

tonnage is 14 million metric tons) but they have lower copper 

grades (the median grade is 0.79 percent). Because of the 

low copper grades, most of the mined rock is waste. Copper 

can be separated by froth flotation of finely ground ore or by 

SX-EW recovery from oxidized ores.

The mineralogy of the ore and waste in porphyry 

copper and sediment-hosted copper mines determines the 

environmental characteristics of these deposits. In terms of 

both ecological risks and human health risks, the mineralogy 

dictates the acid-generating potential of the ores and wastes 

and the mobility of trace elements. In porphyry copper 

deposits, sulfide minerals, such as bornite, chalcopyrite, 

enargite, galena, molybdenite, pyrite, and sphalerite, are the 

main hosts of the trace elements that are of environmental 

concern; in total, they typically constitute less than 5 percent 

of the ore (John and others, 2010). In sediment-hosted copper 

deposits, the sulfide minerals are dominated by, in general 

order of decreasing abundance, chalcocite, digenite, bornite, 

chalcopyrite, pyrite, and subordinate galena and sphalerite 

(Hayes and others, 2015). The trace element geochemistry 

of tailings and waste dumps reflects the mineralogy of the 

waste. For porphyry copper deposits, tailings typically contain 

significant concentrations of copper (475 to 5,100 ppm), 

manganese (67 to 700 ppm), molybdenum (12 to 235 ppm), 

zinc (40 to 210 ppm), and arsenic (3.5 to 136 ppm) (John and 

others, 2010). Tailings from sediment-hosted copper deposits 

generally contain 1 to 4 weight percent iron, 0.1 to 0.3 weight 

percent copper, and 300 to 800 ppm cobalt (Hayes and 

others, 2015).

Approaches for managing solid waste depend upon its 

acid-generating potential. Acid generation can be considered 

a “master variable” for aqueous risks. Metals and many other 

trace elements tend to be more soluble at low pH (acidic) than 

at neutral or high pH (alkaline). Therefore, the acid-generating 

or acid-neutralizing potentials of the waste rock, tailings, 

and other solid waste material are of prime importance in 

identifying the potential environmental risks associated with 

mining and ore beneficiation. The acid-generating potential 

of mine waste is expressed in terms of the amount of calcium 

carbonate it would take to neutralize it; it is measured in 

kilograms of calcium carbonate per metric ton of mine waste 

(kg CaCO3/t) (Price, 2009; International Network for Acid 

Prevention, 2011). The acid-generating potential resides 

primarily in pyrite. Mine waste can also have acid-neutralizing 

potential, which resides in carbonate minerals, such as calcite, 

and in some silicate minerals, such as feldspars.

The rocks associated with porphyry copper deposits, in 

general, tend to straddle the boundary between having net 

acid-generating potential and not having net acid-generating 

potential. During mining and processing of porphyry copper 

deposits, a variety of materials with differing acid-base 

accounts may be encountered and produced, and each type 

must be managed according to its acid-generating potential. 

Net alkaline waste does not require any special handling 

and can be used for construction purposes, whereas net acid 

waste has to be managed to mitigate acid-mine drainage 

problems. As an example, the net neutralizing potentials for 

the hypogene and supergene ores from Morenci, Arizona, 

range from dominantly net acid to slightly net alkaline 

(–257.0 to 1.1 kg CaCO3/t) (Enders and others, 2006). The 

net acid-neutralization potentials reported in the literature for 

tailings from several porphyry copper deposits in Chile are 

net acid generating, ranging from –101.6 to –18.2 kg CaCO3/t 

(Dold and Fontboté, 2001). Tailings derived from metallurgical testing on an exploratory drill core from the Pebble 

deposit in southwestern Alaska have net neutralization 

potentials ranging from –110 to 27.2 kg CaCO3/t (Pebble 

Partnership, 2011).
Environmental Considerations P23

For sediment-hosted copper deposits, the abundance 

of carbonate minerals and slightly less reactive minerals, 

principally chlorite- and epidote-group minerals, determine the 

acid-neutralizing capacity, whereas the acid-generating potential is determined by the abundance of sulfides, such as pyrite 

and chalcopyrite. Many deposits contain a significant amount 

of carbonate and (or) silicate minerals in the ore or host rock, 

which neutralize the acidity generated by sulfide weathering. 

As for acid-generating sulfides, the ore for all three subtypes 

contain low pyrite contents. The pyrite content of the host rock 

for all types of sediment-hosted copper deposits is less than 

1 volume percent (Hayes and Einaudi, 1986).

Limited acid-base accounting data have been reported 

in the literature for sediment-hosted copper deposits. For 

tailings, no acid-generating potential has been detected in 

samples from the White Pine Mine in Michigan (Williams and 

others, 2002), whereas high acid-neutralizing capacities and 

neutral-to-alkaline conditions have been reported in samples 

from the Zambian copperbelt. For the Zambian tailings, the 

neutralization potential ratios are between 3.4 and 84; most 

samples at depth have values of greater than 20, and the 

highest values are for the leached surface materials (Sracek 

and others, 2010).

The mean net neutralization potentials of ore, waste 

rock, and tailings from sandstone deposits at Montanore, 

Rock Creek, and Troy, Montana, were summarized in 

the environmental impact statement for these projects 

(U.S. Department of Agriculture, Forest Service, and others, 

2009). Ore from the Montanore deposit has a mean net 

neutralization potential of –4 kg CaCO3/t, with values ranging 

from –24 to 11 kg CaCO3/t, meaning that its average value 

has a slight acid-generating potential, but ranges from having 

a slight excess of acid-generating potential to having a slight 

excess of acid-neutralizing potential. Values for the Rock 

Creek and the Troy ore samples average 5 and 8 kg CaCO3/t, 

respectively, meaning that they have acid-generating potential. 

The mean net neutralization potentials for the tailings from 

these sandstone deposits range from 2.8 to 10 kg CaCO3/t. 

These acid-base accounting results reflect the low amounts 

of sulfides in the tailings. The mean net neutralization 

potentials for the mine waste from these deposits range from 

3.6 to 15 kg CaCO3/t.

Human Health Concerns

No information is available for rhenium and its toxic 

effects on humans, partly (presumably) because of its low 

natural abundance. The U.S. Agency for Toxic Substances and 

Disease Registry (http://www.atsdr.cdc.gov/toxprofiles/index.asp)

does not have a toxicological profile for rhenium nor does the 

U.S. Environmental Protection Agency have drinking-water 

standards or soil guidelines for rhenium. Unnatural (not naturally occurring) radioisotopes of rhenium (188Re and 186Re), 

however, are being investigated for management of metastatic 

bone pain in cancer patients (Finlay and others, 2005).

Instead, the more significant human health risks associated with rhenium production are found with the broader 

aspects of the mining of porphyry and sediment-hosted copper 

deposits. Contaminated groundwater plumes associated with 

tailings impoundments may threaten drinking-water supplies, 

depending upon the geologic and hydrologic setting and 

engineering aspects of the mine and the waste piles for both 

deposit types. Host rocks with higher acid-neutralization 

potentials, such as carbonate rocks, tend to limit the mobility 

of metals and related compounds. Hydrologic and climatic 

settings that have net evaporative loss of water may cause 

evaporative concentration of solutes that may enter groundwater used as drinking-water supplies in the vicinity of mines 

and waste piles. Improperly constructed water-containment 

structures may allow contaminated mine waters to enter 

surrounding groundwater. For both porphyry copper deposits 

and sediment-hosted deposits, elements or compounds with 

the greatest likelihood of causing problems for drinking-water 

sources include aluminum, arsenic, copper, iron, manganese, 

sulfate, and zinc. The U.S. Environmental Protection Agency 

has primary drinking-water standards, meant to protect human 

health, for arsenic, cadmium, copper, and lead, and nonenforceable secondary standards meant to protect the cosmetic 

and aesthetic qualities of water (skin or tooth discoloration, 

taste, odor, and color) for aluminum, cadmium, iron, lead, 

manganese, sulfate, and zinc (U.S. Environmental Protection 

Agency, 2009a). The potential for groundwater contamination 

depends upon the waste and water management practices at 

any given mine.

An additional human health risk associated with 

porphyry copper and sediment-hosted copper deposits is 

from the smelting of ore concentrates. Historically, it was 

more common for each mine or mining district to operate 

its own smelter. In contrast, modern mines typically ship 

ore concentrates to smelters that may be far away from the 

mines. Therefore, smelter sites are not a necessary feature 

of a proposed mine. Smelters may emit carbon monoxide, 

hydrocarbons, metals, nitrogen oxide, particulates, and sulfur 

dioxide, which can contaminate surrounding soils, although 

in the United States, these emissions are currently regulated 

to protect the environment. For example, numerous studies on 

the soil surrounding the Glogów and the Legnica smelters in 

Poland, which serve sediment-hosted copper mines, reported 

high metal concentrations, especially of, in decreasing 

order of maximum concentration, lead (90 to 18,000 ppm), 

copper (250 to 10,000 ppm), zinc (55 to 4,000 ppm), and 

cadmium (0.3 to 10.9 ppm), that vary with the dominant wind 

direction and distance from the smelter (Roszyk and Szerszen, 

1988; Helios Rybicka and Jędrzejczyk, 1995; Karczewska, 

1996; Pilc and others, 1999; Grzebisz and others, 2001; 

Kabala and Singh, 2001). The concentrations of cadmium and 

lead in groundwater and surface waters in the vicinity of the 

two smelters (Pilc and others, 1999) exceed the World Health 

Organization (WHO) guidelines for drinking water (World 

Health Organization, 2006).
P24 Critical Mineral Resources of the United States—Rhenium

Ecological Health Concerns

No information is available for rhenium and its toxic 

effects on aquatic or terrestrial ecosystems, presumably 

because of its low crustal abundances. Therefore, as with 

human health risks, the more significant ecosystem risks 

associated with rhenium production are found with the broader 

aspects of the mining of porphyry and sediment-hosted copper 

deposits. Many of the ecological risks associated with both 

porphyry copper deposits and sediment-hosted deposits focus 

on the ability of mine wastes to generate acid, and the ability 

of the resulting acid-mine drainage to carry metals and other 

inorganic contaminants. As described above with respect 

to mine-waste characteristics, the acid-generating potential 

of mine waste is primarily found in its pyrite content, and 

it may be offset by the acid-neutralizing potential found in 

carbonate or less significantly silicate (chlorite and epidote 

group) minerals. Residual ore minerals in the waste material 

are the primary hosts of many metals and arsenic in mine 

drainage. Silicate minerals are common sources of aluminum, 

iron, and manganese in mine drainage. Sound waste management practices can mitigate these effects. Tailings are prone 

to being transported by waters, especially in the case of a 

tailings dam failure, and wind, because of the sand- to silt-size 

grains. Thus, they present additional potential risks to aquatic 

organisms through sediment contamination.

Mine-drainage data are available for porphyry copper 

deposits in British Columbia, Canada (Day and Rees, 2006); 

the Globe mining district in Arizona (Eychaner, 1991; 

Stollenwerk, 1994; Brown and others, 1998; Lind and others, 

1998; Conklin and others, 2001); the Morenci mining district, 

Arizona (Enders and others, 2006); and Iran (Khorasanipour 

and others, 2011). Many of the values exceed relevant waterquality guidelines for the protection of aquatic ecosystems 

(U.S. Environmental Protection Agency, 2009b). In British 

Columbia, Canada, Day and Rees (2006) documented mine 

waters with pH values ranging from 2.0 to 8.5, and sulfate was 

the dominant anionic species (1 to 30,000 ppm). The concentrations of trace elements varied widely: aluminum ranged 

from 0.001 to 1,000 ppm; copper, 0.0005 to 1,000 ppm; iron, 

0.005 to 1,000 ppm; manganese, 0.001 to 100 ppm; and 

zinc, 0.001 to 100 ppm. Khorasanipour and others (2011) 

found similar geochemical trends, but in a more arid environment, for drainage associated with waste-rock dumps at the 

Sar Cheshmeh Mine in Kermān Province, southeastern Iran. 

The pH values ranged from 3.1 to 6.3, and the concentration of 

sulfates was between 365 and 1,590 ppm. The concentrations 

of aluminum ranged from less than 0.05 to 60 milligrams 

per liter (mg/L); manganese, from 14.6 to 95.8 ppm; 

copper, 2.15 to 70 ppm; and zinc, 2.4 to 27.4 ppm. In the 

Globe mining district in Arizona, a stream was blocked by 

mill tailings, causing a lake to form. Water from this lake 

entered an alluvial aquifer by seepage, and the aquifer and 

a stream to the north were contaminated (Eychaner, 1991; 

Stollenwerk, 1994; Brown and others, 1998; Lind and others, 

1998; Conklin and others, 2001). The most contaminated 

groundwater in the aquifer had a pH of 3.3 and contained 

about 9,600 ppm sulfate, 2,800 ppm iron, 300 ppm aluminum, 

and 190 ppm copper. As the plume traveled north through 

the aquifer, the concentration of constituents decreased as 

the plume interacted with carbonate-bearing alluvium and 

was diluted by uncontaminated water (including groundwater 

flowing upward from lower basin fill, water in uncontaminated streams that join the contaminated wash, and surface 

rainwater). Enders and others (2006) reported analyses of 

seeps and springs in the Morenci district of Arizona, which 

had pH values of between 2.6 and 4.6. Sulfate concentrations were between 550 and 4,300 ppm; copper, between 

0.46 and 960 ppm; iron, between 15 and 420 ppm; aluminum, 

between 0.48 and 370 ppm; and zinc, between 0.8 and 

159 ppm.

Pit lakes, particularly in porphyry copper mining districts 

that have extensive historical underground mine workings, 

such as the Butte district in Montana, which has a 140-year 

mining history, can be problematic. The Berkeley Pit lake at 

Butte contains more than 100 billion liters of pH-2.5 mine 

water (Gammons and others, 2005; Gammons and Duaime, 

2005). The pit lake has high levels of dissolved solids and 

elevated concentrations of copper, iron, sulfate, and zinc 

(Gammons and Duaime, 2005). In contrast, pit lakes in the 

Yerington and Robinson districts in Nevada have pH values 

that typically range from 7.0 to 8.5 with a few outliers near 

4.7, which have been attributed to discharge from solventextraction operations rather than groundwater-rock interactions in the vicinity of the pits (Shevenell and others, 1999). 

In Nevada, the total dissolved solids are generally less than 

6,000 ppm, and concentrations of manganese are less than 

5 ppm; iron, less than 4.5 ppm; selenium, less than 0.14 ppm; 

and arsenic, less than 0.05 ppm.

Several studies provide insights into the ecosystem risks 

associated with sediment-hosted copper deposits. Detailed 

ecological studies have been carried out along the Kafue River 

in Zambia, which flows through the Zambian Copperbelt 

Province’s mining district. Syakalima and others (2001) 

analyzed water downstream of the mining-affected areas of 

two national parks. This study found concentrations of lead 

in water of between 0.29 and 0.36 ppm compared with the 

WHO drinking-water guideline of 0.01 ppm. Other studies 

focused on the toxic effects of metals in the Kafue River 

on tropical fish. River sediment collected downstream of 

several mines and near a city that has a major ore-processing 

facility was significantly more toxic to zebrafish (Pterois 

volitans) and tilapia (both redbreast tilapia [Tilapia rendalli] 

and spotted tilapia [Tilapia mariae]) than sediment collected 

further downstream (Mwase and others, 1998). The toxicity 

was directly related to the contamination of sediments by 

the mining activities (Mwase and others, 1998). The results 

of another study showed that tilapia exposed in situ to 

Kafue River water for 2 weeks bioaccumulated several trace 

elements (cadmium, chromium, cobalt, copper, and nickel) 

for experiments located downstream of mining activities and 

other industrial point sources (Norrgren and others, 2000). 
Acknowledgments P25

The Zambian copperbelt contains both reduced-facies and 

red-bed-associated deposits, and stream sediments from the 

Kafue River—an area affected by mining of these deposits—

contains up to 0.8 weight percent sulfur, 12,855 ppm copper, 

and 1,030 ppm cobalt (Pettersson and others, 2000; Pettersson 

and Ingri, 2001).

A study by the mining company to assess the local and 

long-term effects on Lake Superior from the discharge from 

the White Pine Mine in Michigan was conducted in 1991. 

The study indicated that discharge from the mine area was 

not causing measurable effects based on a lack of observable 

impact on the benthic community, an insignificant increase 

in chloride from mine discharges, and a lack of detectable 

effects on local water intakes (U.S. Environmental Protection 

Agency, 1994). According to the Surface Water Quality 

Division of the Michigan Department of Natural Resources, a 

diverse fish community exists in the nearby river that receives 

drainage from the mine. According to the U.S. Environmental 

Protection Agency (1994), the Michigan Department 

of Natural Resources report did show a reduced macroinvertebrate community downstream from the mine drainage 

compared with upstream in the river. They suggested that this 

reduction may be owing to major physical and natural stream 

quality differences in the White Pine area.

Several environmental impact statements for sandstone 

deposits in Montana discuss the mobility of trace elements to 

surface water. Tailing effluent and mine water chemistries for 

the Spar Lake (Troy) Mine had concentrations of cadmium, 

copper, and lead in the tailings outflow that exceed aquatic 

ecosystem guidelines (U.S. Department of Agriculture, 

Forest Service, and others, 1992).

Carbon Footprint

Rhenium extraction through mining does not have a 

unique carbon footprint beyond the general energy requirements of mining. In terms of its uses, rhenium is most directly 

linked to the global carbon cycle through its use in platinumrhenium catalysts to produce high-octane, lead-free gasoline 

(Polyak, 2013).

Mine Closure

The methods used to close porphyry copper mines 

and sediment-hosted copper mines depend primarily on the 

method of mining and the characteristics of the waste material. 

Open pit mining of either deposit type produces at least three 

different features after mining: the open pit, tailings storage 

facilities, and waste rock piles. Backfilling pits is typically not 

practical for a variety of reasons. If the water table is above 

the bottom of the pit, the pit will become a lake. The water 

quality of the lake will depend upon a number of factors, 

including the characteristics of the wall rock, the extent of 

underground mine workings that connect to the pit, the water 

level and volume in the pit, the local hydrology, and climate, 

among others (Castendyk and Eary, 2009).

The long-term fate of tailing storage facilities depends 

upon the nature of the tailings and the method of construction 

of the facility. Some tailings storage facilities can be regraded, 

capped, and revegetated. Others are designed to have a water 

cover in perpetuity to limit sulfide oxidation. Either type may 

have seepage that could require some form of water treatment. 

The long-term fate of waste rock piles typically includes 

regrading, capping, and revegetation. Depending upon the 

acid-generating potential of the waste, some piles may also 

require some form of water treatment.

For underground mines, tailings and waste rock may be 

handled similarly to how they are handled at open pit mines. 

Some of the tailings may be transferred back into the minedout workings, depending upon how the orebody is mined. The 

entire volume of tailings cannot be placed back in the mine 

workings, however, because of the volume expansion associated with crushing and milling the ore and the relatively small 

volume of the mineral commodity that is recovered.

Problems and Future Research

The lack of published data about the rhenium contents of 

most deposits, including many deposits from which rhenium is 

produced, hinders understanding of where additional resources 

may be located. The many uncertainties in the estimates of 

rhenium grades and rhenium contents of identified deposits 

preclude more-detailed analysis of rhenium resources.

The geology and geochemistry of rhenium-enriched 

deposits are not fully understood. Areas of future research 

will likely include investigating the following: (a) why 

rhenium is generally more enriched in molybdenite that occurs 

in porphyry copper deposits (where the rhenium content is in 

the hundreds to thousands of parts per million) than in molybdenite that occurs in porphyry molybdenum deposits (where 

the rhenium content ranges from less than one to tens of parts 

per million) and why molybdenite in gold-rich porphyry 

copper deposits tends to have higher rhenium concentrations 

than it does in other subtypes of porphyry copper deposits; 

(b) where rhenium occurs in sediment-hosted strata-bound 

copper deposits and in sandstone uranium deposits; (c) the 

source of rhenium in sediment-hosted strata-bound copper 

deposits; and (d) the origin of the Merlin molybdenumrhenium deposit.

Acknowledgments

Peter Vikre, Steve Ludington, and John DeYoung provided

helpful reviews. David Sinclair provided unpublished data 

and a preprint of the Millensifer and others (2014) paper. 

Keith Labay helped create the deposit location map and 

Suzanne Nicholson provided additional editorial comments 

after the technical review was complete.
P26 Critical Mineral Resources of the United States—Rhenium

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Table P2 and Appendix P1
P38 Critical Mineral Resources of the United States—Rhenium

Table P2. Rhenium data for selected porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS2, molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Deposit

name Country Deposit subtype1

Tectonic

setting

Grade

(wt. % Cu)

Grade

(wt. % Mo)

Grade

(g/t Au)

Minimum 

Re in 

MoS2

(ppm)

Maximum

Re

in MoS2

(ppm)

Estimated 

average

Re in

MoS2

(ppm)

Porphyry copper deposits 

Agarak Armenia Cu Continental arc 0.56 0.025 0.6 57 6,310 820 

Ajax West Canada Cu Island arc 0.31 0.005 0.2 None None 3,161 

Aksug Russia Cu Post-collisional 0.67 0.015 0.12 None None 460 

Aktogai (or 

Aktogay)

Kazakhstan Cu-Mo Continental arc 0.39 0.01 0.026 50 2,700 850 

Bagdad United States Cu-Mo Continental arc 0.4 0.01 0.0011 330 642 460 

Berg Canada Cu-Mo Continental arc 0.39 0.031 0.06 67 215 152 

Bethlehem–

Huestis

Canada Cu-Mo Island arc 0.4 0.005 0.012 None None 417 

Bethlehem–Iona Canada Cu-Mo Island arc 0.52 0.006 0.012 None None 1,015 

Bethlehem–JA Canada Cu-Mo Island arc 0.43 0.017 0.01 200 246 222 

Bingham United States Cu Post-collisional 0.882 0.053 0.38 130 2,000 250 

Borly Kazakhstan Cu Continental arc 0.34 0.011 0.3 250 5,500 3,160 

Boschekul Kazakhstan Cu-Mo Island arc 0.67 0.0023 0.049 230 1,500 825 

Brenda Canada Cu-Mo Island arc 0.152 0.037 0.013 80 145 115 

Bronson Slope Canada Cu-Au Island arc 0.17 0.006 0.44 None None 180 

Butte United States Cu-Mo Continental arc 0.673 0.028 0.042 None None 240 

Cananea Mexico Cu Continental arc 0.45 0.002 0.035 None None 700 

Casino Canada Cu Continental arc 0.25 0.025 — 65 289 197 

Castle Dome 

(Pinto Valley)

United States Cu-Au Continental arc 0.33 0.0055 0.34 1,200 1,750 1,750 

Cerro Verde Peru Cu Continental arc 0.495 0.01 — 3,060 3,497 3,280 

ChuquicamataRadomiro 

Tomic

Chile Cu-Mo Continental arc 0.86 0.04 0.013 93 262 265 

Collahuasi Chile Cu-Mo Continental arc 0.592 0.04 0.01 368 448 395 

Copper Creek United States Cu Continental arc 0.75 0.0046 — 534 2,107 1,165 

Cuajone Peru Cu Continental arc 0.69 0.0214 — None None 580 
Table P2 P39

Table P2. Rhenium data for selected rhenium porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Number 

of Re 

analyses2

Grade3

(g/t Re)

Deposit 

tonnage 

(Mt)

Contained 

Mo 

(t)

Contained 

Re 

(t)

Mo:Re 

ratio Data sources Deposit name

Porphyry copper deposits

106, 0, 0

1, 0, 0

0, 0, 1

30, 0, 0

0, 7, 2

4, 0, 0

1, 0, 0

1, 0, 0

4, 0, 0

36, 6, 1

19, 0, 0

23, 0, 0

11, 0, 2

1, 0, 1

1, 0, 0

0, 0, 1

4, 0, 0

3, 0, 2

0, 2, 0

1, 6, 2

0, 3, 0

0, 3, 0

1, 0, 1

0.342

0.263

0.115

0.14

0.08

0.079

0.035

0.102

0.063

0.221

0.579

0.032

0.071

0.018

0.112

0.023

0.082

0.160

0.12

0.18

0.26

0.089

0.207

125

365

337

2,636

1,600

238

1.4

30

260

3,230

94

1,000

182

79

5,220

5,141

559

1,438

2,528

21,277

3,100

75

1,630

31,300

18,300

50,600

260,000

160,000

73,800

70

1,780

44,200

1,710,000

10,400

23,000

67,200

4,740

1,460,000

103,000

140,000

79,090

250,000

8,500,000

1,200,000

3,500

348,800

43

96

39

370

130

19

0.05

3.0

16

714

55

32

13

1.4

585

118

46

230

300

3,800

820

6.7

337

730

190

1,300

700

1,200

4,000

1,400

590

2,700

2,400

190

730

5,200

3,300

2,500

871

3,100

344

830

2,300

1,500

520

1,030

Berzina and others, 2005; Singer and others, 2008

Sinclair and others, 2009

Berzina and others, 2005; Singer and others, 2008

Berzina and others, 2005; Singer and others, 2008

Sutulov, 1974; Nadler, 1997; Barra and others, 

2003; Singer and others, 2008

Sinclair and others, 2009

Sinclair and others, 2009

Sinclair and others, 2009

Sinclair and others, 2009

Giles and Schilling, 1972; McCandless and Ruiz,

1993; Chesley and Ruiz, 1998; Singer and 

others, 2008; Austen and Ballantyne, 2010; 

J.T. Chesley, Ph.D., written commun., 2013

Berzina and others, 2005; Singer and others, 2008

Singer and others, 2008; Sinclair and others, 2009

Sutulov, 1974; Sinclair and others, 2009; W.D. 

Sinclair, Ph.D., Geological Survey of Canada, 

written commun., 2013

Sinclair and others, 2009

Giles and Schilling, 1972; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; Singer 

and others, 2008

Sinclair and others, 2009

Giles and Schilling, 1972; Singer and others, 2008

Mathur and others, 2001; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Singer and others, 2008; 

Barra and others, 2013

Mathur and others, 2001; Masterman and others, 

2004; Singer and others, 2008

McCandless and Ruiz, 1993; Barra and others, 

2005; Singer and others, 2008

Nadler, 1997; Mathur and others, 2001; Singer 

and others, 2008

Agarak

Ajax West

Aksug

Aktogai (or 

Aktogay)

Bagdad

Berg

Bethlehem–

Huestis

Bethlehem–Iona

Bethlehem–JA

Bingham

Borly

Boschekul

Brenda

Bronson Slope

Butte

Cananea

Casino

Castle Dome 

(Pinto Valley)

Cerro Verde

ChuquicamataRadomiro 

Tomic

Collahuasi

Copper Creek

Cuajone
P40 Critical Mineral Resources of the United States—Rhenium

Table P2. Rhenium data for selected porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Deposit 

name Country Deposit subtype1

Tectonic 

setting

Grade 

(wt. % Cu)

Grade 

(wt. % Mo)

Grade 

(g/t Au)

Minimum 

Re in 

MoS 2

(ppm)

Maximum 

Re 

in MoS 2

(ppm)

Estimated 

average 

Re in 

MoS 2

(ppm)

Porphyry copper deposits—Continued

Cuatro Hermanos Mexico Cu Continental arc 0.431 0.035 — None None 469

Cumobabi Mexico Cu-Mo Continental arc 0.266 0.099 — 189 368 279

Dastakert Armenia Cu Continental arc 0.62 0.048 — 130 300 220

Duobaoshan China Cu Island arc 0.46 0.016 0.128 122 885 560

Elatsite (or Bulgaria

Elatzite)

El Salvador Chile

Cu

Cu

Continental arc

Continental arc

0.39

0.86

0.01

0.022

0.26

0.1

273

None

2,740

None

1,250

585

El Teniente Chile Cu-Mo Continental arc 0.62 0.019 0.005 25 1,154 420

Ely United States Cu Continental arc 0.613 0.01 0.27 1,250 2,840 1,600

Erdenet (Erdene- Mongolia

tuin-Obo)

Gibraltar Canada

Cu

Cu

Post-collisional

Island arc

0.62

0.29

0.025

0.006

—

0.07

104

238

534

750

199

443

Granisle (or Bell) Canada Cu Continental arc 0.43 0.005 0.13 522 528 526

Highmont Canada

West Pit

Cu-Mo Island arc 0.15 0.05 0.04 137 176 157

Huckleberry Canada

Hushamu Canada

Cu

Cu-Au

Continental arc

Island arc

0.49

0.198

0.014

0.0092

0.04

0.278

247

None

258

None

253

3,140

Ingerbelle (or Canada

Similco)

Island Copper Canada

Cu-Au

Cu

Island arc

Island arc

0.329

0.338

0.002

0.0088

0.17

0.19

None

1,654

None

1,863

1,620

1,730

Kadjaran Armenia

(Kadzharan)

Kalmakyr Uzbekistan

(Almalyk)

Kemess South Canada

Cu

Cu-Au

Cu-Au

Continental arc

Continental arc

Island arc

0.27

0.38

0.22

0.055

0.006

0.008

0.65

0.6

0.65

33

700

3,106

2,620

2,000

4,609

280

1,500

3,858

Kounrad Kazakhstan Cu Continental arc 0.589 0.011 0.19 620 4,050 1,540
Table P2 P41

Table P2. Rhenium data for selected rhenium porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Number 

of Re 

analyses2

Grade3

(g/t Re)

Deposit 

tonnage 

(Mt)

Contained 

Mo 

(t)

Contained 

Re 

(t)

Mo:Re 

ratio Data sources Deposit name

Porphyry copper deposits—Continued

0, 1, 0

0, 2, 0

8, 0, 1

0, 8, 0

19, 0, 0

1, 0, 2

1, 14, 2

4, 0, 1

2, 1, 1

3, 0, 1

0, 0, 5

2, 0, 0

2, 0, 0

1, 0, 0

1, 0, 0

0, 0, 11

237, 0, 1

20, 0, 1

2, 0, 0

20, 0, 1

0.274

0.460

0.176

0.149

0.21

0.215

0.133

0.27

0.043

0.044

0.044

0.13

0.059

0.481

0.054

0.262

0.257

0.150

0.514

0.282

233

67

36

951

350

3,836

20,731

754

1,780

935

171

0.8

73

510

78

600

1,700

2,000

213

637

81,600

66,000

17,000

152,000

35,000

844,000

3,940,000

75,000

445,000

56,100

8,560

400

10,300

46,900

1,600

52,800

935,000

120,000

17,000

70,100

64

31

6.2

142

73

825

2,760

200

77

41

7.5

0.10

4.3

245

4.2

157

437

300

109

180

1,300

2,100

2,700

1,070

480

1,020

1,430

380

5,810

1,360

1,100

3,800

2,378

191

380

336

2,140

400

155

390

Barra and others, 2005; Singer and others, 2008

Barra and others, 2005; Singer and others, 2008

Sutulov, 1974; Berzina and others, 2005; Global 

Metals (ARM) Ltd., 2015 

Zhao and others, 1997; Singer and others, 2008; 

Deng and others, 2013

Singer and others, 2008; Sinclair and others, 

2009

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Maksaev and others, 2004; 

Cannell, 2004; Klemm and others, 2007; 

Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Singer and others, 2008

Watanabe and Stein, 2000; Berzina and others, 

2005; Singer and others, 2008

Sinclair and others, 2009; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Sinclair and others, 2009; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Sinclair and others, 2009

Sinclair and others, 2009

Giroux and Casselman, 2012

Sinclair and others, 2009

Sinclair and others, 2009; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Nadler, 1997; Berzina and others, 2005; 

Singer and others, 2008

Sutulov, 1974; Singer and others, 2008; 

Pašava and others, 2010

Sinclair and others, 2009

Sutulov, 1974; Berzina and others, 2005; 

Singer and others, 2008

Cuatro Hermanos

Cumobabi

Dastakert

Duobaoshan

Elatsite (or 

Elatzite)

El Salvador

El Teniente

Ely

Erdenet (Erdenetuin-Obo)

Gibraltar

Granisle (or Bell)

Highmont 

West Pit

Huckleberry

Hushamu

Ingerbelle (or 

Similco)

Island Copper

Kadjaran 

(Kadzharan)

Kalmakyr 

(Almalyk)

Kemess South

Kounrad
P42 Critical Mineral Resources of the United States—Rhenium

Table P2. Rhenium data for selected porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Deposit 

name Country Deposit subtype1

Tectonic 

setting

Grade 

(wt. % Cu)

Grade 

(wt. % Mo)

Grade 

(g/t Au)

Minimum 

Re in 

MoS 2

(ppm)

Maximum 

Re 

in MoS 2

(ppm)

Estimated 

average 

Re in 

MoS 2

(ppm)

Porphyry copper deposits—Continued

La Caridad Mexico Cu Continental arc 0.452 0.0247 — 72 570 570

La Escondida Chile Cu-Au Continental arc 0.769 0.0062 0.25 95 1,805 886

Lomex Canada Cu-Mo Island arc 0.404 0.014 0.006 286 427 351

Los BroncesRio Blanco 

(Andina)

Los Pelambres

Chile

Chile

Cu-Mo

Cu-Mo

Continental arc

Continental arc

0.601

0.617

0.02

0.015

—

0.028

104

450

898

820

265

600

Machangqing China Cu-Au Collision belt(?) 0.64 0.08 0.35 31 125 80

Maggie Canada Cu-Mo Continental arc 0.28 0.029 — None None 643

Majdanpek Serbia Cu-Au Continental arc 0.6 0.005 0.35 2,320 3,550 2,770

Medet Bulgaria Cu Continental arc 0.37 0.01 0.1 None None 905

Miami United States Cu-Mo Continental arc 0.63 0.01 0.009 None None 600

Mineral Park 

(Ithaca Peak)

Mission-Pima

United States

United States

Cu

Cu

Continental arc

Continental arc

0.489

0.52

0.011

0.015

—

—

250

None

290

None

270

600

Mitchell 

(Sulphurets)

Morenci

Canada

United States

Cu-Au

Cu-Mo

Island arc

Continental arc

0.18

0.524

0.005

0.0095

0.69

0.028

7,012

270

8,170

640

7,590

455

Mt. Tolman United States Cu-Mo Continental arc 0.09 0.054 — None None 182

OK Canada Cu Continental arc 0.34 0.016 — None None 746

Pebble United States Cu Post-collisional 0.592 0.0243 0.342 329 2,070 1,100

Qulong China Cu-Mo Post-collisional 0.52 0.032 — 16 303 125

Ray United States Cu Continental arc 0.68 0.001 — 440 1,500 820

San ManuelKalamazooUnited States Cu-Mo Continental arc 0.6 0.011 0.017 700 1,200 900

Santa Rita United States Cu Continental arc 0.468 0.008 0.056 700 1,200 800
Table P2 P43

Table P2. Rhenium data for selected rhenium porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Number 

of Re 

analyses2

Grade3

(g/t Re)

Deposit 

tonnage 

(Mt)

Contained 

Mo 

(t)

Contained 

Re 

(t)

Mo:Re 

ratio Data sources Deposit name

Porphyry copper deposits—Continued

0, 2, 1

0, 7, 0

2, 0, 20

0, 13, 2

0, 3, 0

0, 0, 5

1, 0, 0

3, 0, 0

22, 0, 1

1, 0, 0

2, 0, 1

0, 1, 1

2, 0, 0

3, 1, 0

0, 0, 1

1, 0, 0

0, 6, 2

0, 4, 0

9, 0, 0

2, 0, 2

8, 0, 1

0.235

0.092

0.081

0.09

0.150

0.11

0.311

0.231

0.15

0.10

0.050

0.150

0.633

0.072

0.163

0.199

0.446

0.067

0.014

0.165

0.107

1,800

11,158

460

16,816

7,458

39

181

1,000

244

1,591

876

900

734

6,470

2,177

64

5,940

1,517

1,583

1,390

3,030

444,600

691,800

64,400

3,400,000

1,120,000

31,000

52,600

50,000

24,000

160,000

96,400

135,000

36,700

614,700

1,180,000

10,000

1,443,000

485,000

15,800

153,000

242,000

423

1,030

37

1,500

1,120

4

56

231

37

160

44

135

465

466

355

13

2,650

102

22

229

324

1,050

672

1,728

2,300

1,000

7,429

932

216

651

1,000

2,200

1,000

79

1,320

3,330

790

545

4,770

710

667

746

Nadler, 1997; Valencia and others, 2005; 

Singer and others, 2008

Mathur and others, 2001; Singer and others, 

2008; Romero and others, 2010

Sinclair and others, 2009; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Mathur and others, 2001; Singer and others, 

2008; Deckart and others, 2013

Mathur and others, 2001; Singer and others, 2008

Hou and others, 2006

Sinclair and others, 2009

Todorov and Staikov, 1985; Singer and others, 

2008

Sutulov, 1974; Berzina and others, 2005; 

Singer and others, 2008

Berzina and others, 2005; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Singer and others, 2008

Sutulov, 1974; McCandless and Ruiz, 1993; 

Singer and others, 2008

Sinclair and others, 2009

Giles and Schilling, 1972; McCandless and 

Ruiz, 1993; Singer and others, 2008

Carten and others, 1993; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Sinclair and others, 2009

Ghaffari and others, 2011; Lang and others, 2013

Singer and others, 2008; Hou and others, 2009

Giles and Schilling, 1972; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Singer and others, 2008.

Giles and Schilling, 1972; Sutulov, 1974; 

Singer and others, 2008

La Caridad

La Escondida

Lomex

Los BroncesRio Blanco 

(Andina)

Los Pelambres

Machangqing

Maggie

Majdanpek

Medet

Miami

Mineral Park 

(Ithaca Peak)

Mission-Pima

Mitchell 

(Sulphurets)

Morenci

Mt. Tolman

OK

Pebble

Qulong

Ray

San ManuelKalamazoo

Santa Rita
P44 Critical Mineral Resources of the United States—Rhenium

Table P2. Rhenium data for selected porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Deposit 

name Country Deposit subtype1

Tectonic 

setting

Grade 

(wt. % Cu)

Grade 

(wt. % Mo)

Grade 

(g/t Au)

Minimum 

Re in 

MoS 2

(ppm)

Maximum 

Re 

in MoS 2

(ppm)

Estimated 

average 

Re in 

MoS 2

(ppm)

Porphyry copper deposits—Continued

Sar Cheshmeh Iran Cu Continental arc 1.2 0.03 0.27 11 517 597

Schaft Creek Canada Cu Island arc 0.25 0.019 0.18 None None 590

Sierrita- United States

Esperanza

Cu-Mo Continental arc 0.294 0.0292 0.003 90 1,800 238

Silver Bell United States Cu-Mo Continental arc 0.66 0.013 0.026 340 620 531

Skouriés Greece Cu-Au Post-collisional 0.35 0.002 0.47 800 1,000 900

Snowfields Canada Cu-Au Island arc 0.08 0.008 0.50 None None 3,600

Sora (Sorsk) Russia Cu Post-collisional? 0.17 0.058 — 6 18 14

Tominskoe Russia Cu Island arc 0.58 0.004 0.12 None None 1,080

Tongchankou China Cu uncertain 0.94 0.04 — 176 235 208

Toquepala Peru Cu-Mo Continental arc 0.55 0.04 — 387 1,496 600

Tsagaan-Suvarga Mongolia Cu Continental arc 0.53 0.018 0.084 80 156 118

Twin Buttes United States Cu-Mo Continental arc 0.502 0.023 0.019 None None 600

Valley Copper Canada Cu-Au Island arc 0.44 0.0067 0.006 None None 294

Veliki Krivelj Serbia Cu Continental arc 0.44 0.004 0.068 None None 302

Wunugetushan China Cu-Mo Post-collisional? 0.46 0.053 — 142 369 199

Yulong China Cu-Au Post-collisional? 0.99 0.028 0.35 291 665 444

Zuun Mod Mongolia

Molybdenum

Cu-Mo Continental arc? 0.069 0.059 — 250 300 275

Porphyry molybdenum deposits

Boss Mountain Canada Arc-related Continental arc — 0.074 — 49 157 80

Carmi Canada Arc-related Continental arc — 0.064 — 10 139 58

Endako Canada Arc-related Continental arc 0.002 0.07 — 15 67 35

Glacier Gulch 

(Davidson)

Canada Arc-related Continental arc 0.04 0.177 — 34 41 38
Table P2 P45

Table P2. Rhenium data for selected rhenium porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Number 

of Re 

analyses2

Grade3

(g/t Re)

Deposit 

tonnage 

(Mt)

Contained 

Mo 

(t)

Contained 

Re 

(t)

Mo:Re 

ratio Data sources Deposit name

Porphyry copper deposits—Continued

15, 0, 5

1, 0, 0

6, 1, 2

18, 1, 0

4, 0, 0

1, 0, 0

9, 0, 0

1, 0, 0

0, 6, 0

1, 2, 2

0, 2, 0

0, 0, 1

0, 0, 1

1, 0, 0

0, 7, 0

0, 0, 2

2, 0, 0

0.30

0.187

0.116

0.115

0.030

0.480

0.014

0.072

0.14

0.40

0.035

0.230

0.033

0.020

0.176

0.207

0.270

1,200

1,393

2,262

268

568

2,203

300

241

45

2,320

240

940

791

750

850

628

218

360,000

265,000

660,500

34,900

11,400

17,600

174,000

9,640

18,000

930,000

43,200

216,000

53,000

30,000

450,000

176,000

129,000

360

260

262

31

17

1,060

4.2

17

6.2

930

8.4

216

26

15

150

130

59

1,000

1,020

2,520

1,130

670

17

41,000

560

2,900

1,000

5,100

1,000

2,030

2,000

3,000

1,350

2,190

Singer and others, 2008; Aminzadeh and 

others, 2011

Sinclair and others, 2009

Giles and Schilling, 1972; Sutulov, 1974; 

McCandless and Ruiz, 1993; Nadler, 1997; 

Singer and others, 2008

Giles and Schilling, 1972; Barra and others, 

2005; Singer and others, 2008

Singer and others, 2008; Sinclair and others, 2009

Armstrong and others, 2011

Sotnikov and others, 2001; Berzina and others, 

2005; Berzina and Korobeinikov, 2007

Singer and others, 2008; Sinclair and others, 2009

Xie and others, 2007; Singer and others, 2008

Giles and Schilling, 1972; Sutulov, 1974; 

Nadler, 1997; Mathur and others, 2001; 

Singer and others, 2008

Watanabe and Stein, 2000; Singer and others, 2008

Sutulov, 1974; Singer and others, 2008

Sinclair and others, 2009; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Singer and others, 2008; Sinclair and others, 2009

Chen and others, 2011

Hou and others, 2006

Clark and Baudry, 2011

Sar Cheshmeh

Schaft Creek

SierritaEsperanza

Silver Bell

Skouriés

Snowfields

Sora (Sorsk)

Tominskoe

Tongchankou

Toquepala

Tsagaan-Suvarga

Twin Buttes

Valley Copper

Veliki Krivelj

Wunugetushan

Yulong

Zuun Mod 

Molybdenum

Porphyry molybdenum deposits

7, 0, 0

3, 0, 0

14, 12, 1

2, 0, 0

0.099

0.062

0.04

0.112

63

21

600

75

46,600

13,300

420,000

134,000

6.2

1.3

25

8.4

7,500

10,000

17,000

16,000

Sinclair and others, 2009

Sinclair and others, 2009

Giles and Schilling, 1972; Selby and Creaser, 

2001; Sinclair and others, 2009; W.D. Sinclair, Ph.D., Geological Survey of Canada, 

written commun., 2013

Sinclair and others, 2009

Boss Mountain

Carmi

Endako

Glacier Gulch 

(Davidson)
P46 Critical Mineral Resources of the United States—Rhenium

Table P2. Rhenium data for selected porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Deposit 

name Country Deposit subtype1

Tectonic 

setting

Grade 

(wt. % Cu)

Grade 

(wt. % Mo)

Grade 

(g/t Au)

Minimum 

Re in 

MoS 2

(ppm)

Maximum 

Re 

in MoS 2

(ppm)

Estimated 

average 

Re in 

MoS 2

(ppm)

Porphyry molybdenum deposits—Continued

Kitsault (Lime Canada

Creek)

Lucky Ship Canada

Mount Haskin Canada

Nithi Mountain Canada

Quartz Hill United States

Red Bird Canada

Red Mountain Canada

Storie Moly Canada

Thompson Creek United States

Trout Lake Canada

(Max)

Adanac (Ruby Canada

Creek)

Climax United States

Donggou China

Jinduicheng China

Questa United States

Shapinggou China

Urad-Henderson United States

Xiaodonggou China

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Arc-related Continental arc

Alk-granite/ Extensional conhybrid? tinental arc

Alk-granite Continental rift

Alk-granite Collision belt

Alk-granite Collision belt

Alk-granite Continental rift

Alk-granite Collision belt

Alk-granite Continental rift

Alk-granite Collision belt(?)

0.004

—

—

—

0.003

0.07

—

—

—

—

0.001

—

—

—

—

—

—

—

0.115

0.067

0.09

0.02

0.0762

0.065

0.10

0.078

0.071

0.12

0.059

0.20

0.116

0.099

0.15

0.126

0.228

0.109

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

36

None

None

None

None

6

None

15

None

56

8

10

4.1

15.5

6

0.4

7

4.5

129

None

None

None

None

43

None

22

None

73

22

80

4.3

16.2

145

14.7

20

8.4

71

41

108

76.9

149

25

32

20

120

56

12

13

4.2

15.9

36

4.7

20

7.1

1

Cox and Singer (1992) porphyry Cu models; Taylor and others (2012) and Ludington and Plumlee (2009) porphyry Mo models.

2

Number of analyses of MoS separates, MoS analyzed for Re-Os dating, MoS mill concentrates. Number in bold indicates samples used to 2 2 2

calculate Re grade. In cases where no number is bold, average grade of total resources calculated from drilling was used as Re grade.

3

Re grade calculated from mean Re content of MoS and Mo grade of deposit. 2
Table P2 P47

Table P2. Rhenium data for selected rhenium porphyry copper and porphyry molybdenum deposits of the world.—Continued

[The names, locations, and types of most of the deposits are taken from the U.S. Geological Survey Mineral Resources Data System (MRDS) (U.S. Geological 

Survey, 2014). The minimum and maximum values are concentrations of rhenium in molybdenite separates; the estimated average is based on mill concentrates. 

Elements and compound: Au, gold; Cu, copper; Mo, molybdenum; MoS , molybdenite; Os, osmium; Re, rhenium. Units of measure: g/t, gram per metric ton; 2

Mt, million metric tons; ppm, part per million; t, metric ton; wt. %, weight percent. —, no data; None, no minimum and maximum values are available because 

only a single analysis was used to calculate the Re grade]

Number 

of Re 

analyses2

Grade3

(g/t Re)

Deposit 

tonnage 

(Mt)

Contained 

Mo 

(t)

Contained 

Re 

(t)

Mo:Re 

ratio Data sources Deposit name

Porphyry molybdenum deposits—Continued

9, 0, 0

1, 0, 0

1, 0, 0

0, 1, 0

0, 0, 1

2, 0, 0

1, 0, 0

3, 0, 0

0, 0, 1

1, 0, 1

4, 0, 0

13, 0, 4

0, 2, 0

0, 7, 0

14, 8, 1

0, 9, 0

2, 0, 2

0, 6, 0

0.136

0.046

0.16

0.03

0.189

0.027

0.05

0.026

0.142

0.11

0.012

0.04

0.008

0.026

0.090

0.010

0.076

0.013

104

62

12

240

1,600

75

187

101

212

43

144

800

594

1,089

424

1,270

437

42

120,000

41,000

11,000

47,900

1,220,000

49,000

190,000

78,400

151,000

51,000

84,800

1,600,000

689,000

1,080,000

640,000

1,600,000

996,000

45,000

14

2.8

2.0

6.2

302

2.0

10

2.6

30

4.8

1.7

35

4.8

28

38

13

33

0.5

8,500

15,000

5,600

7,700

4,030

24,000

19,000

30,000

5,000

11,000

50,000

45,000

145,000

38,000

17,000

126,000

30,000

83,000

Sinclair and others, 2009

Sinclair and others, 2009

Sinclair and others, 2009

Selby and Creaser, 2001; Mosher, 2001

Hudson and others, 1979; Wolfe, 1995; W.D. 

Sinclair, Ph.D., Geological Survey of Canada, 

written commun., 2013

Sinclair and others, 2009

Sinclair and others, 2009

Sinclair and others, 2009

Carten and others, 1993; W.D. Sinclair, Ph.D., 

Geological Survey of Canada, written commun., 2013

Sinclair and others, 2009

Sinclair and others, 2009

Giles and Schilling, 1972; Nadler, 1997; Singer 

and others, 1993; Sinclair and others, 2009; 

W.D. Sinclair, Ph.D., Geological Survey of 

Canada, written commun., 2013

Mao and others, 2011; Deng and others, 2013

Mao and others, 2011; Deng and others, 2013

Giles and Schilling, 1972; Singer and others, 

1993; Rosera and others, 2013; W.D. Sinclair, 

Ph.D., Geological Survey of Canada, written 

commun., 2013

Mao and others, 2011; Deng and others, 2013

Giles and Schilling, 1972; Nadler, 1997; Seedorff and Einaudi, 2004; Markey and others, 

2007; W.D. Sinclair, Ph.D., Geological Survey of Canada, written commun., 2013

Zheng and others, 2010

Kitsault (Lime 

Creek)

Lucky Ship

Mount Haskin

Nithi Mountain

Quartz Hill

Red Bird

Red Mountain

Storie Moly

Thompson Creek

Trout Lake (Max)

Adanac (Ruby 

Creek)

Climax

Donggou

Jinduicheng

Questa

Shapinggou

Urad-Henderson

Xiaodonggou
P48 Critical Mineral Resources of the United States—Rhenium

Appendix P1. Rhenium Data Sources and Limitations of Data Used in Rhenium

Resource Estimates

Kupferschiefer resources in Poland.—Ore reserves in 

deposits in the Legnica- Glogów copper belt, Poland, which 

includes the Lubin-Sieroszowice orebody, are estimated to 

be 1,180 million metric tons (Bartlett and others, 2013) that 

averages 0.6 ppm rhenium (Smakowski and others, 2010). 

These tonnage and grade data suggest a rhenium resource of 

709 metric tons.

Kupferschiefer resources in the Mansfeld-Sangerhausen 

area, Germany.—Jankowski (1995) reported production of 

120 million metric tons of ore from 1200 to 1990 that had 

an estimated average rhenium grade of 21 grams per metric 

ton (g/t). This suggests that more than 2,500 metric tons of 

rhenium was mined (but not recovered). Jankowski (1995) 

lists a pre-mining resource of 155 million metric tons with 

an average grade of 21 g/t, which suggests a total pre-mining 

resource of about 3,300 metric tons of rhenium and that about 

800 metric tons of rhenium remains. Jankowski (1995) also 

reports remaining resources of 35.4 million metric tons of ore 

with an average rhenium grade of 21 g/t, which suggests that 

about 740 metric tons of rhenium remains in these deposits.

Average rhenium grades and deposit tonnages for the 

Dzhezhazgan and the Zhaman-Aybat sandstone-type stratabound copper deposits in Kazakhstan are available from Box 

and others (2013). Rhenium grades for other similar deposits 

in Kazakhstan are not available.

Rhenium contents of some roll-front-type sandstone 

uranium deposits in Kazakhstan and Uzbekistan are reported 

by Dahlkamp (2009a, b). For most deposits, however, total 

contained uranium or total uranium production and ranges 

of uranium and (or) rhenium concentration data are reported, 

which precludes calculation of rhenium grades, tonnages, 

and resources.

The estimated tonnage and rhenium grade for the 

Merlin deposit in Australia is a Canada National Instrument 

43–101-compliant estimate of indicated and inferred mineral 

resources made in April 2012 (Lycopodium Minerals QLD Pty 

Ltd., 2012).

Because of these limitations in the data for known 

deposits, estimated rhenium resources in table P1 and shown 

in figure P4 are likely imprecise, but are accurate enough 

to allow discussion of some aspects of U.S. and global 

rhenium resources.

Publicly available data for rhenium resources are limited. 

Of more than 225 porphyry copper deposits with published 

molybdenum grades and tonnages (Singer and others, 2008), 

rhenium concentrations are available for only about 80 

deposits, several of which are represented by a single rhenium 

analysis (table P1; John and Taylor, 2016). There are similarly 

few rhenium data available for porphyry molybdenum 

deposits. The available rhenium analytical data are a mixture 

of analyses of (a) small molybdenite separates (for example, 

Giles and Schilling, 1972; Sinclair and others, 2009), (b) 

small molybdenite separates used in rhenium-osmium dating 

studies (for example, McCandless and Ruiz, 1993; Barra 

and others, 2013), and less commonly, (c) bulk molybdenite 

mill concentrates (mostly from Sutulov, 1974, and Nadler, 

1997). The molybdenite separates and mill concentrates are 

subject to impurities, and some of the variation in rhenium 

content within deposits may be the result of variable purity of 

these molybdenite separates. Electron microprobe analyses 

of rhenium contents of molybdenites are available for some 

deposits (Newberry, 1979; McCandless and others, 1993), but 

these analyses have relatively high detection limits and low 

precision and were not included in the data compilation.

Calculated rhenium resources in porphyry copper and 

porphyry molybdenum deposits are based on the average 

concentration of rhenium in molybdenite, the average molybdenum grade for the entire deposit, and the total tonnage of 

the deposit. The type of analysis used to determine rhenium 

concentrations in molybdenite is indicated in table P1.

Molybdenum grades and tonnages for porphyry copper 

deposits are mostly from Singer and others (2008) and are 

subject to the rules specified in their data compilation. For 

example, average molybdenum grades and the associated 

tonnages are based on the total production, reserves, and 

resources at the lowest possible cutoff grade, and all mineralized 

and altered rock within 2 km are combined into one deposit. 

Tonnages, therefore, are pre-mining resources. Because many 

porphyry copper deposits have been mined for decades or 

longer (for example, the Bingham Canyon deposit has been 

mined since 1906), remaining tonnages and rhenium resources 

for these deposits are smaller than indicated in table P1.

Rhenium data for strata-bound copper deposits are 

limited to a few deposits in the Kupferschiefer in Poland and 

Germany and two deposits in Kazakhstan.
Appendix P1 P49

Appendix P1 References Cited

Note: All Web links listed were active as of the access date but may no 

longer be available.

Barra, Fernando, Alcota, Hugo, Rivera, Sergio, Valencia, 

Victor, Munizaga, Francisco, and Maksaev, Victor, 2013, 

Timing and formation of porphyry Cu–Mo mineralization in 

the Chuquicamata district, northern Chile—New constraints 

from the Toki cluster: Mineralium Deposita, v. 48, no. 5, 

p. 629–651. [Also available at http://dx.doi.org/10.1007/

s00126-012-0452-1.]

Bartlett, S.C., Burgess, Harry, Damjanović, Bogdan, 

Gowans, R.M., and Lattanzi, C.R., 2013, Technical report on 

the copper-silver production operations of KGHM 

Polska Miedź S.A. in the Legnica-Glogów copper belt 

area of southwestern Poland, NI 43–101 technical report, 

prepared for KGHM Polska Miedź S.A.: Toronto, Ontario, 

Canada, Micon International Ltd., 159 p., accessed 

September 2, 2015, at http://kghm.com/en/technical-reportcopper-silver-production-operations-kghm-polska-miedzsa-legnica-glogow-copper-belt.

Box, S.E., Syusyura, Boris, Seltmann, Reimar, Creaser, R.A., 

Dolgopolova, Alla, and Zientek, M.L., 2013, Dzhezkazgan 

and associated sandstone copper deposits of the Chu-Sarysu 

basin, central Kazakhstan, in Hedenquist, J.W., Harris, Michael, 

and Camus, Francisco, eds., Geology and genesis of major 

copper deposits and districts of the world—A tribute to 

Richard H. Sillitoe (1st ed.): Littleton, Colo., Society of 

Economic Geologists Special Publication no. 16, p. 303–328.

Dahlkamp, F.J., 2009a, Kazakhstan, in Dahlkamp, F.J., ed.,

Uranium deposits of the world—Asia: Berlin, Germany,

Springer-Verlag, p. 181–189. [Also available at 

http://dx.doi.org/10.1007/978-3-540-78558-3.]

Dahlkamp, F.J., 2009b, Uzbekistan, in Dahlkamp, F.J., ed.,

Uranium deposits of the world—Asia: Berlin, Germany,

Springer-Verlag, p. 397–400. [Also available at 

http://dx.doi.org/10.1007/978-3-540-78558-3.]

Giles, D.L., and Schilling, J.H., 1972, Variation in the rhenium

content of molybdenite, in Program—24th International 

Geological Congress, Montreal, Quebec, Canada, 1972 

Proceedings of sec. 10: Ottawa, Ontario, Canada, International Geological Congress, p. 145–153.

Jankowski, Gunter, ed., 1995, Zur geschichte des Mansfelder 

Kupferschiefer-Bergbaus [The history of the Mansfeld 

copper shale mining]: Clausthal-Zellerfeld, Germany, 

Gesellschaft Deutscher Metallhütten- und Bergleute, 

370 p., folded page with 2 maps.

John, D.A., and Taylor, R.D., 2016, Byproducts of porphyry 

copper and molybdenum deposits, in Verplanck, P.L., and 

Hitzman, M.W., eds., Rare earth and critical elements 

in ore deposits: Reviews in Economic Geology, v. 18, 

p 137–164.

Lycopodium Minerals QLD Pty Ltd., 2012, NI 43-101 

technical report, Merlin molybdenum rhenium project, 

feasibility study, northwest Queensland, Australia: Spring 

Hill, Queensland, Australia, Lycopodium Minerals 

QLD Pty Ltd., 302 p., accessed March 14, 2013, at

http://www.ivanhoeaustralia.com/i/pdf/NI43-101_

Merlin_Technical_Report.pdf.

McCandless, T.E., Ruiz, Joaquin, and Campbell, A.R., 1993, 

Rhenium behavior in molybdenite in hypogene and nearsurface environments—Implications for Re-Os geochronometry: Geochimica et Cosmochimica Acta, v. 57, no. 4, 

p. 889–905. [Also available at http://dx.doi.org/10.1016/

0016-7037(93)90176-w.]

Nadler, H.-G., 1997, Rhenium, in Habashi, Fathi, ed., 

Handbook of extractive metallurgy—Precious metals, 

refractory metals, scattered metals, radioactive metals, 

rare earth metals: New York, N.Y., Wiley-VCH, v. 3, 

p. 1491–1501.

Newberry, R.J., 1979, Polytypism in molybdenite (II)—

Relationships between polytypism, ore deposition/alteration 

stages and rhenium contents: American Mineralogist, v. 64, 

nos. 7–8, p. 768–775.

Sinclair, W.D., Jonasson, I.R., Kirkham, R.V., and Soregaroli, A.E.,

2009, Rhenium and other platinum-group metals in porphyry

deposits: Geological Survey of Canada Open File 6181,

1 sheet. [Also available at http://dx.doi.org/10.4095/247485.]

Singer, D.A., Berger, V.I., and Moring, B.C., 2008, Porphyry 

copper deposits of the world—Database and grade and tonnage models, 2008 (version 1.0): U.S. Geological Survey 

Open-File Report 2008–1155, 45 p., accessed April 23, 2015,

at http://pubs.usgs.gov/of/2008/1155/.

Smakowski, Tadeusz, Ney, Roman, and Galos, Krzysztof, 

eds., 2010, Minerals yearbook of Poland 2009: Kraków, 

Poland, Department of Mineral Policy, Mineral and Energy 

Economy Research Institute, Polish Academy of Sciences, 

537 p. [Also available at http://www.min-pan.krakow.pl/

pliki/Minerals_Yearbook_of_Poland_2009.pdf.]

Selenium

By Lisa L. Stillings

Chapter Q of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–Q

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter Q

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................Q1

Introduction..................................................................................................................................................Q1

Uses and Applications ......................................................................................................................Q2

Demand and Availability of Supply .................................................................................................Q4

Supply.. .......................................................................................................................................Q6

Demand.......................................................................................................................................Q6

Recycling ....................................................................................................................................Q7

Geology.........................................................................................................................................................Q7

Crustal Abundance ............................................................................................................................Q7

Mineralogy..........................................................................................................................................Q7

Deposit Types .....................................................................................................................................Q9

Magmatic and Associated Hydrothermal Deposits ............................................................Q9

Selenide-Bearing Deposits ...................................................................................................Q13

Sedimentary Rocks .................................................................................................................Q14

Resources and Reserves .........................................................................................................................Q14

Identified Resources .......................................................................................................................Q19

Undiscovered Resources ...............................................................................................................Q21

Environmental Considerations ................................................................................................................Q21

Sources and Fate in the Environment ..........................................................................................Q21

Natural and Anthropogenic Sources in Air, Water, and Soil ...........................................Q21

Average Ambient Concentrations in Air, Streams, Groundwater, and Soil ...................Q23

Above-Background Concentrations of Selenium in the Environment Near 

Major Deposits and Enriched Bedrock .................................................................Q23

Mobility in Water and Soils ...................................................................................................Q26

Mine Waste Characteristics ..........................................................................................................Q27

Human Health Concerns .................................................................................................................Q29

Pathways for Selenium Ingestion in Humans ....................................................................Q30

Related Diseases, Cancers, and Human Health Effects From Industrial 

Exposures ...................................................................................................................Q30

Human Health Guidelines ......................................................................................................Q30

Ecological Health Concerns ...........................................................................................................Q30

Pathways for Selenium Ingestion in Aquatic and Terrestrial Organisms......................Q31

Aquatic Toxicity and Aquatic Ecosystem Health Guidelines ...........................................Q32

Problems and Future Research ..............................................................................................................Q32

Acknowledgments....................................................................................................................................Q32

References Cited.......................................................................................................................................Q33

Suggested citation:

Stillings, L.L., 2017, Selenium, chap. Q of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, and Bradley, D.C., eds., 

Critical mineral resources of the United States—Economic and environmental geology and prospects for future 

supply: U.S. Geological Survey Professional Paper 1802, p. Q1– Q55, http://dx.doi.org/10.3133/pp1802Q.
Selenium

By Lisa L. Stillings

Abstract

Selenium (Se) was discovered in 1817 in pyrite from 

copper mines in Sweden. It is a trace element in Earth’s crust, 

with an abundance of three to seven orders of magnitude less 

than the major rock-forming elements. Commercial use of 

selenium began in the United States in 1910, when it was used 

as a pigment for paints, ceramic glazes, and red glass. Since 

that time, it has had many other economic uses—notably, in 

the 1930s and 1940s, when it was used in rectifiers (which 

change alternating current to direct current), and in the 1960s, 

when it began to be used in the liner of photocopier drums. 

In the 21st century, other compounds have replaced selenium 

in these older products; modern uses for selenium include 

energy-efficient windows that limit heat transfer and thin-film 

photovoltaic cells that convert solar energy into electricity.

In Earth’s crust, selenium is found as selenide minerals, 

selenate and selenite salts, and as substitution for sulfur in 

sulfide minerals. It is the sulfide minerals, most commonly 

those in porphyry copper deposits, that provide the bulk of the 

selenium produced for the international commodity market. 

Selenium is obtained as a byproduct of copper refining and 

recovered from the anode slimes generated in electrolytic 

production of copper. Because of this, the countries that have 

the largest resources and (or) reserves of copper also have the 

largest resources and (or) reserves of selenium.

Because selenium occurs naturally in Earth’s crust, its 

presence in air, water, and soil results from both geologic 

reactions and human activity. Selenium is found concentrated 

naturally in soils that overlie bedrock with high selenium 

concentrations. Selenium mining, processing, use in industrial 

and agricultural applications, and disposal may all contribute 

selenium to the environment. A well-known case of selenium 

contamination from agricultural practices was discovered in 

1983 in the Kesterson National Wildlife Refuge in California. 

There, waters draining from agricultural fields created 

wetlands with high concentrations of dissolved selenium in 

the water. The selenium was taken up by aquatic wildlife and 

caused massive numbers of embryonic deformities and deaths.

Regulatory agencies have since worked to safeguard 

ecological and human health by creating environmental 

exposure guidelines based upon selenium concentrations in 

water and in fish tissue. Any attempt to regulate selenium 

concentrations requires a delicate balance because selenium 

occurs naturally and is also a vital nutrient for the health of 

wildlife, domestic stock, and humans. Selenium is commonly 

added as a vitamin to animal feed, and in some regions of the 

United States and the world, it is added as an amendment to 

soils for uptake by agricultural crops.

The important role of selenium in economic products, 

energy supply, agriculture, and health will continue for well 

into the future. The challenge to society is to balance the 

benefits of selenium use with the environmental consequences 

of its extraction. Increased understanding of the elemental 

cycle of selenium in the earth may lead to new (or unconventional) sources of selenium, the discovery of new methods of 

extraction, and new technologies for minimizing the transfer 

of selenium from rock to biota, so to protect environmental 

and human health.

Introduction

Selenium (Se) was discovered in 1817 (and named in 

1818) by the Swedish chemist Jöns Jacob Berzelius, who 

became interested in the sludge left over after the processing 

of pyrite from the copper mines at Falun, Sweden. The sludge 

was red in color and developed the unpleasant smell of rotten 

radishes when it was heated. Berzelius had first thought 

that the sludge contained tellurium but determined that the 

material, though having close affinities with tellurium, was 

a previously unknown element. Because of its affinity with 

tellurium (from Latin tellus, meaning Earth), the element was 

named selenium after the Greek word selene, meaning Moon, 

or satellite of Earth (Sindeeva, 1964; Lakin and Davidson, 

1973; Boyd, 2011).

In the past, selenium has been described as a somewhat 

mysterious element, one that can be “a metal or nonmetal, 

a conductor or nonconductor, amorphous or crystalline, 

colorant or deodorant, [or] a hydrogenator or dehydrogenator” 

(Sargent, 1954, p. 1). It can be either red or black (Sargent, 

1954) and, when ingested, either a necessary nutrient or a 

poison (Trelease and Beath, 1949). It is fair to say that today, 

almost 200 years after the discovery of selenium, scientific 
Q2 Critical Mineral Resources of the United States—Selenium

understanding has reconciled these seemingly contradictory 

descriptions. The complexity of selenium chemistry in 

economic, environmental, and geologic contexts is better 

understood yet remains a topic of research.

Selenium is a trace element in Earth’s crust with an 

average crustal abundance of 0.05 part per million (ppm). Its 

abundance is three to seven orders of magnitude less than that 

of the major rock-forming elements (fig. Q1), and it is one-half 

to three orders of magnitude less abundant than the rare-earth 

elements. It is, however, more abundant (by one to four 

orders of magnitude) than the rarest metals, such as platinum, 

palladium, and gold.

Selenium is photovoltaic, meaning that it converts light 

into electricity. It is also photoconductive, meaning that its 

electrical conductivity increases with increasing light. These 

are the properties that make selenium useful in photovoltaic 

cells and solar panels, and are among the reasons why 

selenium has been classified as an energy-critical element 

(American Physical Society Panel on Public Affairs and the 

Materials Research Society, 2011). This chapter provides an 

overview of the geochemical cycle of selenium at Earth’s 

surface and includes sections on selenium’s occurrence in 

mineral deposits, its production and consumption, its presence 

in the environment, and its role in both human and environmental health.

Uses and Applications

Before World War I, according to Sargent (1954, p. 8), 

“selenium was regarded as the abomination of the smelterman 

and was discarded.” Today, in comparison, selenium is 

considered an “energy critical element” that is important for 

renewable energy technologies (American Physical Society 

Panel on Public Affairs and the Materials Research Society, 

2011), and it is also known to be an essential nutrient for 

humans and livestock.

Commercial use of selenium in the United States began in 

1910; at that time, selenium was used as a pigment for paints, 

ceramic glazes, and red glass. The first large-scale commercial 

use of selenium began in 1915, when selenium was substituted 

for manganese in glassmaking. Manganese, which was used 

to neutralize the green tint in glass caused by iron impurities, 

had previously been obtained from Russia, but the supply of 

manganese was disrupted during World War I. Selenium, in 

concentrations of 0.04 to 0.14 kilogram per metric ton (kg/t) 

in glass, was found to be a good substitute for manganese. In 

addition, greater concentrations of selenium were found to 

provide a pink tint to glass, and at concentrations of 0.45 to 

2.3 kg/t, could produce a desirable ruby red glass for use 

in tableware, automobile taillights, and traffic signals. The 

modern sheet-glass industry uses selenium in energy-efficient 

Li

Be

B

Sc

Cu Zn

Y

Mo

Mo

Cd

Sn

Sb

Pb

Bi

Tl

I

N

V Cr

Co

Ga

Ge As

Sr

Rb Zr

Nb

Hf

W

Th

U

Hg

Ba

In

Ni

Se Ag

Ta

Br Cs

Nd

SmGd Er

Dy Yb

Ce

La

Eu

Ho

Tm Tb

Pr

Lu

Cl

O

Na

C F P

S

K Ca

Al Si

Rock-forming elements

Rare-earth

elements

Mg

Fe

Mn

Ti

H

Te

Rh

Re

Os

Pt

Au

Ir

Ru

Au

Pd

Rarest “metals”

100 20 30 40 50 60 70 80 90

Atomic number, Z

Abundance (atoms of element per 106 atoms of Si)

13

103

10–3

10–6

106

109

Rock-forming elements

Rare-earth elements—

 Lanthanides (La to Lu) 

 and Y; labeled in blue

Major

Minor

“Rarest” metals—Six 

 platinum-group elements

 plus Au, Re, and Te 

 (a metalloid)

EXPLANATION

Selenium figure Q1

Major industrial metal—Global 

 production greater than about

 3×107

 kg/yr; labeled in bold

Precious metal—Labeled in italic

Lu

Figure Q1. Graph showing relative abundance of the chemical elements in Earth’s upper crust (from Haxel and 

others, 2002). Selenium (Se) is a trace element with an estimated concentration of approximately 0.05 part per 

million in Earth’s crust (Greenwood and Earnshaw, 1997). See the periodic table at the front of this publication for 

definitions of element abbreviations. kg/yr, kilogram per year
Introduction Q3

windows to help limit the heat transfer from sunlight (Sargent, 

1954; Butterman and Brown, 2004).

After glassmaking, the next largest commercial use 

for selenium was the selenium rectifier. Rectifiers, which 

change alternating current (AC) to direct current (DC), had 

been known about since 1884, but they were not extensively 

used until inventors improved the devices in 1930 (Sargent, 

1954). After this time, rectifiers began to be used in arc lamps, 

battery chargers, DC motors, electroplating, magnet coils, and 

welding. The military used selenium rectifiers in World War II 

for aircraft controls, ground starters for aircraft, guided 

missiles, radar equipment, and numerous other applications. 

In 1947, miniature rectifiers were developed that could be 

used in radio and television circuits. This miniaturization 

was a tremendous improvement because the new technology 

could deliver 200 to 500 volts of DC from a 117-volt AC 

source, thus eliminating the need for the older heavy and 

bulky transformers and rectifier tubes, and saving greater than 

0.23 kilogram (kg) of copper and 3.4 kg of steel per television 

set (Sargent, 1954).

Soon after World War II, many other uses for selenium 

were developed. Selenium became important in the rubber 

industry where it acted as an accelerator and vulcanizing 

agent, and it was also used to improve the resilience of rubber 

by increasing its resistance to heat, oxidation, and abrasion. 

Selenium was used as a pigment to produce orange, red, and 

maroon colors in the post-World War II plastics industry and 

in ceramics, dyes, inks, leather, paints, paper, rubber, soap, 

and other products. It was used to either hydrogenate or 

dehydrogenate organic compounds, crude oil, and coal, and it 

assisted in the production of edible fats, plastics, soaps, and 

waxes. It provided antioxidant properties to inks and oils, as 

well as nondrying properties to linseed, oiticica, and tung oils. 

Selenium began to be used in personal care products, such as 

dandruff and dermatitis shampoos and soaps, and in deodorants. It was also included in fungicides, herbicides, and insecticides. Finally, selenium came to be used in the steel industry 

to improve the machinability of stainless steel. Forging steels 

contained 0.18 to 0.22 percent selenium, and machining-grade 

steel contained 0.30 to 0.35 percent. A small amount of 

selenium (0.01 to 0.05 percent) assured a fine-grained and 

bubble-free texture in cast steel. Selenium was also added to 

copper and copper alloys to improve machinability and tensile 

strength (Sargent, 1954; Butterman and Brown, 2004).

The 1960s saw yet another expansion in the industrial use 

of selenium—as a coating on photocopier drums. Xerography, 

which was the forerunner to modern-day photocopying 

machines, was invented in 1938 and patented in 1942. It 

gained commercial popularity in the late 1950s when an automated process was developed using a cylindrical drum coated 

with selenium. Even today, in the 21st century, the majority 

of paper documents are produced xerographically, either on 

photocopiers or laser printers, which employ a similar process 

but use materials other than selenium as the photoreceptor 

(Owen, 2004).

By 1973, 40 percent of the selenium consumed in the 

United States was a high-purity grade used for electrostatic 

printing (xerography), rectifiers, exposure meters, photoelectric cells, and optical lenses. Thirty percent went toward the 

manufacturing of glass. Fourteen percent was used for inorganic pigments, mainly in the form cadmium sulfoselenide, 

which have good light stability and heat resistance and were 

used in enamels, inks, paints, plastics, and rubber. In the 

1990s, amorphous silicon and organic compounds started 

to replace selenium in xerographic drums, and the use of 

selenium in the photocopier industry began to decline (fig. Q2; 

Lakin and Davidson, 1973).

Year

0

100

200

300

400

500

600

700

800

Selenium end use, in metric tons

Other, including metallurgy and agriculture

Electronics and photocopier components

Chemicals and pigments

Glass manufacturing

EXPLANATION

Selenium figure Q2

1975 1980 1985 1990 1995 2000 2005 2010

Figure Q2. Graph showing end uses for selenium in the 

United States from 1975 to 2012. Data are unavailable 

for the years 2004, 2006, and 2007; hence, these data are 

interpolated. End use amounts for the years 2008 through 

2012 are calculated as domestic apparent consumption 

of selenium multiplied by the fraction of world consumption attributed to each type of use. Data are from 

U.S. Geological Survey (2006–13, 2010–14, 2014b).

In the 21st century, potentially large-scale demand for 

selenium could result from the manufacture of thin-film photovoltaic cells that employ a copper-indium-gallium-selenide 

(CIGS) alloy for converting solar energy into electricity. As 

of 2010, silicon-based photovoltaic cells still dominated the 

solar energy market and accounted for greater than 90 percent 

of all photovoltaic cells in use. These photovoltaic cells are 

comparatively inefficient, however, and convert only about 

15 percent of adsorbed light to electricity in actual outdoor use 

(Bleiwas, 2010). Recent research on the second-generation 
Q4 Critical Mineral Resources of the United States—Selenium

photovoltaic thin-film cells that use CIGS technology is 

showing increased efficiencies of 20 to 40 percent and leading 

to lower production costs. Selenium is a critical element in the 

CIGS technology and contributes to the higher levels of light 

absorption and conversion to electricity demonstrated by these 

second-generation thin-film cells.

Bleiwas (2010) described the quantities of selenium 

and other metals required to produce thin-film CIGS photovoltaic cells with enough capacity (1 gigawatt) to produce 

8,760 gigawatt-hours (GWh) of electricity (enough to supply 

800,000 average U.S. households) annually. He calculated 

that 180 metric tons of selenium would be needed to produce 

this quantity of CIGS photovoltaic cells, which is 7.6 percent 

of the estimated 2013 world refinery production of selenium. 

The required quantities of gallium and indium necessary for 

this technology are 30 and 90 metric tons, or 7 percent and 

11 percent of the 2014 worldwide production of these metals, 

respectively (Bleiwas, 2010).

Demand and Availability of Supply

Sargent (1954) provided a picture of selenium as a critical 

element in the post-World War II years. As noted above, 1947 

was a critical year because of the development of a miniature 

selenium rectifier. For much of the 1940s and 1950s, selenium 

consumption outpaced primary refinery production, and 

the difference was made up from imports and releases from 

industry stockpiles (fig. Q3).

To safeguard the supply of selenium, the Federal Government had begun holding selenium in a strategic stockpile in 

the 1940s, although data on the size of the stockpile were 

confidential (Clark, 1950). Increased concerns regarding 

possible shortages became apparent when, in February 1952, 

the National Production Authority placed both imported and 

domestically produced selenium under a “complete allocation” 

in order to meet essential defense requirements with respect 

to steelmaking and rectifiers. This allocation was discontinued 

in early 1953, although consumption continued to be higher 

than domestic production through 1960 (with the exception 

of 1957, fig. Q3; Sargent, 1954; Massachusetts Institute of 

Technology, 2010; American Physical Society Panel on Public 

Affairs and the Materials Research Society, 2011). Ingerson 

(1964) noted that, throughout the 1950s and early 1960s, 

production and consumption of selenium was “nip and tuck,” 

and that there was not a sufficient stockpile of selenium to last 

more than a few months. The situation was so “precarious” 

that selenium was constantly listed in Group I (that is, one of 

the most important) in the national stockpile list of strategic 

and critical materials. In 1961, the U.S. Bureau of Mines 

reported that the Federal stockpile held 254,000 pounds (lb) 

(115 metric tons) of selenium (U.S. Bureau of Mines, 1963), 

and, in 1964, the Office of Emergency Planning increased the 

stockpile objective for selenium to 475,000 lb (215 metric 

tons) (U.S. Bureau of Mines, 1965).

By 1970, the balance between supply and demand had 

shifted—selenium consumption became consistently less than 

the total available supply (primary refinery production plus 

imports plus release from stockpiles) and the excess selenium 

was available for export (fig. Q3; Lakin and Davidson, 

1973). In August 1971, Congress authorized the disposal 

of the national stockpile, and the last lot was sold on May 

26, 1976 (U.S. Bureau of Mines, 1973, 1978). Despite the 

sufficient domestic supply of selenium, however, there was 

some concern in 1973 of an impending shortage of selenium 

owing to (a) the increased use of in situ leaching for copper 

production from its ores, which did not allow for recovery of 

associated selenium, and (b) the discovery that selenium was 

necessary for animal nutrition. It was learned that selenium 

deficiency in commercial feed could result in death and 

disease for poultry and livestock; this happened in a number 

of countries and led to heavy economic losses. The deficiency 

could be reversed, however, by adding selenium to fertilizer 

for agricultural lands. Lakin and Davidson (1973) estimated 

that 5.7 million kilograms per year of selenium would be 

required to treat the most severely deficient agricultural lands 

in the United States.

Jensen (1985) tracked the supply of and demand 

for selenium from 1973 to 1983 by comparing selenium 

production from U.S. mines to total U.S. supply, including 

available stockpiles. During this period, U.S. primary refinery 

production peaked at 360 metric tons in 1973, dropped to a 

low of 140 metric tons in 1980, and recovered to 350 metric 

tons (revised) in 1983 (U.S. Bureau of Mines, 1989). The total 

U.S. supply was always higher than mine production during 

this period, with a high of 831 metric tons in 1974 and a low 

of 581 metric tons in 1977. (Jensen’s calculations include 

the entire quantity of selenium held in stockpiles, and not the 

change in stockpile mass, as indicated in fig Q3.) The difference between mine production and total supply was owing to 

the refineries’ use of U.S. Government selenium stockpiles, 

selenium imports, and industry stockpiles. The largest releases 

from the U.S. Government stockpile took place in 1973 and 

1974 (150 and 55 metric tons, respectively) (U.S. Bureau of 

Mines, 1980). The last shipment took place in 1976, and since 

then, refineries have relied on an increased supply of selenium 

from imports (fig. Q3).

George (2013) tracked the price of selenium from 1970 

to 2010 (fig. Q4). As with any commodity, the price reflects 

supply and demand. By using yearly prices in 1992 constant 

dollars, George showed that the price of selenium in the 

1970s was higher than it was in 2010. Prices spiked far above 

2010 prices during 1974–76 because of the elimination of 

the U.S. Government selenium stockpile, low commercial 

inventories, reduced domestic inventories, and increased 

demand for plain-paper copiers. Prices remained relatively 

steady from 1981 to 2003, at approximately $4 to $12 per 

pound (in 1992 dollars), because there was sufficient worldwide supply, because selenium substitutes were available for 

many industrial applications, and because there were no new 

large-scale uses for selenium. The price rise and volatility for 
Introduction Q5

Selenium figure Q3

Selenium, in metric tons

–1,000

0

1,000

2,000

EXPLANATION

Primary refinery production

Imports for consumption

Exports

Change in Government stockpile

Change in industry stockpile

Apparent consumption

Imports as proxy for apparent consumption

Year

1943

1945

1947

1949

1951

1953

1955

1957

1959

1961

1963

1965

1967

1969

1971

1973

1975

1977

1979

1981

1983

1985

1987

1989

1991

1993

1995

1997

1999

2001

2003

2005

2007

2009

2011

2013

2015

Figure Q3. Graph illustrating a mass balance for selenium in the United States 

from 1943 to 2015. The supply of selenium to the United States is estimated from 

primary refinery production plus imports for consumption (meaning that material is 

added to the supply for that year), minus exports from the United States (removal 

of selenium from the supply for that year—shown as a negative quantity), plus 

or minus changes in Government and industry stockpiles (releases are shown as 

positive quantities and additions are shown as negative quantities), to calculate 

domestic apparent consumption. 

Data from 1943 to 2014 are from U.S. Geological Survey (2014c), and data for 2015 are from U.S. Geological Survey (2017). 

Primary refinery production numbers are not available for 1985 to 1987, and from 1997 to 2015, they are withheld because 

they are proprietary; hence, no primary refinery production data are plotted for these years. For the years 1985 to 1987, 

apparent consumption data have been estimated by linear interpolation. For 1997 to 2015, domestic apparent consumption 

data are withheld to avoid disclosing company proprietary data, and imports for consumption are used as a proxy for 

apparent consumption (U.S. Geological Survey, 2014c, 2017). The Federal Government held selenium in a strategic stockpile 

in the 1940s, but data on the size of the stockpile were confidential (Clark, 1950, p. 1342). The U.S. Bureau of Mines reported 

the quantity of selenium held in the national stockpile in 1961 (U.S. Bureau of Mines, 1966, p. 1071) and continued to report 

the size of the stockpile until the last of the selenium was sold on May 26, 1976 (U.S. Bureau of Mines, 1978, p. 1467), and 

there has been no stockpile since 1976. The USGS reported numbers for the industry-held stockpile from 1944 through 

1984, but these values have been withheld since 1985 to avoid disclosing company proprietary data (U.S. Geological 

Survey, 2014c). Domestic secondary production of selenium (selenium recovered from recycled rectifiers, catalysts, and 

photocopier drums) was never a large percentage of domestic supply and was inconsistently reported during the period 

1943–2016 and thus is not included in this compilation.
Q6 Critical Mineral Resources of the United States—Selenium

selenium since 2003 is attributed mainly to increased demand 

for selenium from manganese producers in China. Manganese, 

which is an alloying element in steel production, is produced 

by an electrolytic process. Substitution of selenium dioxide 

(SeO2(g)) for sulfur dioxide (SO2(g)) lowers the energy requirements of the electrolytic cells, allowing manganese to be 

produced more cheaply.

Supply

The period from 1945 to 1960 was a period when selenium 

demand (apparent consumption) outpaced supply from primary 

refinery production (1957 was an exception). During this time, 

U.S. primary refinery production averaged 290 metric tons 

per year; U.S. producers were American Smelting & Refining 

Co., U.S. Metals Refining Co., and International Smelting & 

Refining Co. (fig. Q3). During 1961–64, the total supply of 

selenium was greater than consumption, and the extra supply 

was stored in industry and Government stockpiles (fig. Q3). 

Primary refinery production decreased in 1965–68, and 

selenium was released from stockpiles to meet demand. Beginning in 1969, the total supply of selenium has been sufficient 

to meet demand, and excess supply has been exported.

Domestic production from 1947 to 1953, which was the 

beginning of selenium consumption for modern technologies, 

contrasts markedly with production nearly 50 years later. 

The latest available report of domestic selenium production 

was for 1996, when 379,000 kg (379 metric tons) of selenium 

was produced as a primary product. Production figures were 

not published after 1996 to avoid disclosing proprietary data 

from the one or two known domestic refineries, and because 

selenium-containing residues and anode slimes from other 

domestic copper refiners were exported for processing.

Just as U.S. domestic production of selenium increased 

in the years following 1947, so did global production. Sargent 

(1954) reported that during 1926–31, the Soviet Union 

produced between 2.3 and 2.7 metric tons per year of selenium. In 1946, Canada produced 237 metric tons, although 

this production level could not be sustained owing to a 

reduction in output of selenium-rich high-grade ores from the 

Horne Mine at Noranda, Quebec, Canada; by 1950, Canada’s 

production fell to less than 160 metric tons. In 1951, Sweden 

produced 41 metric tons; Belgium, 10 metric tons; and West 

Germany, 7.3 metric tons. In 1952, Belgium increased its 

production to 17 metric tons, and Africa produced 14 metric 

tons. Belgium increased production again in 1953 to about 

27 metric tons, and production in the Soviet Union in that year 

was estimated to have increased to 45 metric tons.

Total global production in 1996 (the last year for which 

domestic production numbers were reported) was 2,250 metric 

tons, and the United States accounted for almost 17 percent of 

global production, or about 380 metric tons. U.S. production 

ranked third behind Canada (694 metric tons) and Japan 

(588 metric tons). During the years 1997 to 2015, Japan 

was the leading international producer of selenium, although 

it briefly slipped into second place in 2005. Germany was 

the second-ranked international producer in 2015. China 

and the United States produced selenium during this time 

period as well, but their production numbers are not available 

(U.S. Geological Survey, 1997–2014).

Demand

From 1943 to 2015, U.S. consumption of selenium 

was between 184 (1946) and 896 (1969) metric tons per 

year (fig. Q3) and was affected by increased substitution 

0

10

20

30

40

50

60

1970 1975 1980 1985 1990 1995 2000 2005 2010

Dollars per pound

Year

Price

1992 dollars

EXPLANATION

Significant events affecting selenium prices since 1970

1974–76 Government stocks liquidated by 1974; low commercial 

inventories; reduced domestic production from recession and 

copper industry strike; increased import dependence; 

continued growth in plain-paper copiers

1976–80 Stock buildup and reduced demand following 1973–75 

recession; production level established at about one-half 

of 1969 peak

1981–83 Demand surged; stocks remained high; plain-paper copiers 

and glass manufacturing dominate demand

1984–89 World stocks declined as demand outstrips production; 

speculation encouraged price fluctuations; domestic demand 

averaged 1.3 million pounds

1990–91 World production rose; demand slackened owing to recession; 

stock decline was reversed

1990–99 Organic photoreceptor compounds replaced selenium 

compounds for plain-paper copiers

2003–10 Increased but fluctuating demand from manganese producers 

in China; global financial crisis in 2008 reduced demand but 

supplies remained tight, and prices remained relatively high

Figure Q4. Graph showing average annual prices of commercial-grade selenium from 1970 to 2010. Both the nominal 

(unadjusted) prices and real prices (in 1992 constant dollars) are shown. Graph is from George (2013).
Geology Q7

of other materials for selenium and changes in technology 

that made many selenium products obsolete. As examples, 

high-purity silicon replaced, or has become a major substitute 

for, selenium in rectifiers; silicon can substitute for selenium 

in photovoltaic cells; and amorphous silicon and organic 

photoreceptors have replaced selenium in photocopier drums. 

Organic pigments can substitute for cadmium sulfoselenide 

pigments; cerium oxide (CeO) can substitute for selenium in 

glass; tellurium can substitute for selenium in pigments and 

rubber compounding; bismuth, lead, and tellurium can substitute for selenium in metal alloys; and tellurium can substitute 

for selenium in lead-free brass. Even with a potential increased 

use of selenium in CIGS photovoltaic solar cells, the demand 

for selenium will likely remain steady for the near future.

Recycling

Traditionally, selenium has been used in U.S. manufacturing for glass decolorizers, pigments, fertilizers, animal 

feeds, and metal alloys. These have been termed “dissipative” 

uses because of the small amounts of selenium involved, the 

sheer number of these products, and the difficulty in recovering selenium from them because of the disposal or recycling 

practices for the products. Even when selenium-containing 

glass and metal alloys are recycled, they are not segregated 

from other forms of scrap, and the selenium content is 

probably volatilized during melting. Major nondissipative 

uses for selenium have been as photoreceptors in photocopier 

drums and as rectifiers; however, manufacturers began to 

replace selenium with organic photoreceptor compounds in the 

1980s; consequently, there is no longer a substantial supply 

of photocopier drums containing selenium for recycling, and 

large selenium rectifiers have been replaced with ones that 

use silicon. Even the use of selenium in AC current rectifiers 

has ended, as silicon-based rectifiers have been developed 

that are cheaper and more reliable. As a result of this lack of 

supply, no selenium recycling facilities remain in operation in 

the United States. This may change in the future as selenium 

is increasingly used in the manufacture of thin-film CIGS 

photovoltaic cells for solar panels (George and Wagner, 2009).

Geology

The distribution of selenium in Earth’s crust is closely 

associated with that of sulfur (S), and because of their similar 

ionic radii—198 picometers (pm) for Se(-II) versus 184 pm 

for S(-II)—selenium readily substitutes for sulfur in structures 

of sulfide minerals. Selenium is less commonly found in 

sulfate deposits. Selenium is brought to Earth’s surface by way 

of magmatic intrusions, where it is found in sulfide minerals, 

and by way of volcanic emissions, where it is vented as the 

gases selenium dioxide (SeO2

) and hydrogen selenide (H2Se).

Crustal Abundance

The crustal abundance of selenium has been estimated in 

different ways. Sindeeva (1964) estimated 0.14 ppm selenium 

for rocks in the Soviet Union; Brunfelt and Steinnes (1967) 

estimated 0.04 ppm based on standard rock samples provided 

by the U.S. Geological Survey (USGS), and Wells (1967) 

reported an average of 0.6 ppm selenium in volcanic ash 

samples from New Zealand and 0.4 ppm in New Zealand lava 

flows (Lakin and Davidson, 1973, and references therein).

An indirect method for estimating selenium abundance 

was suggested by Goldschmidt and Strock (1935) based upon 

an estimated sulfur-to-selenium ratio of 6,000:1 in sulfide 

minerals. Starting with the crustal abundance for sulfur of 

520 ppm, they divided this number by the sulfur-to-selenium 

ratio and arrived at an estimate of 0.09 ppm selenium in 

Earth’s crust. Turekian and Wedepohl (1961) revised the 

crustal abundance of sulfur to 300 ppm, and by using the same 

sulfur-to-selenium ratio of 6,000:1, they estimated 0.05 ppm 

selenium in Earth’s crust. This figure of 0.05 ppm remains the 

accepted estimate of the crustal abundance of selenium today 

(fig. Q1; Greenwood and Earnshaw, 1997).

Mineralogy

Selenium forms selenide minerals and selenite and 

selenate salts with chalcophile elements (for example, 

bismuth, copper, lead, mercury, and silver). Some of the 

natural selenides, such as berzelianite (Cu2Se), clausthalite 

(PbSe), guanajuatite (Bi2Se3), krutaite (CuSe2), tiemannite 

(HgSe), and umangite (Cu3Se2) form from selenium-rich 

hydrothermal fluids (Pirri, 2002) in deposits that “span the 

magmatic-hydrothermal spectrum” (Ciobanu and others, 

2006, p. 163)—from 65 to 300 degrees Celsius (°C), as 

determined from fluid inclusion and isotopic studies, up to 

800 °C in laboratory synthesis studies. Other selenium-bearing 

minerals, such as aguilarite (Ag4SeS), contain selenium and 

sulfur in isomorphic substitution, and some systems show a 

continuous solid solution between selenium and sulfur end 

members. Selenium forms only one oxide mineral: downeyite 

(SeO2

). The selenium minerals recognized by the International 

Mineralogical Association (IMA) are listed in table Q1 at 

the back of this chapter. In a literature survey of selenium 

concentrations reported in worldwide sulfide and selenide 

deposits, selenium concentrations in selenides can be as high 

as 70.23 weight percent, such as in trogtalite (CoSe2), which is 

found in the Musonoi Mine in the Democratic Republic of the 

Congo (Congo [Kinshasa]), and up to 4,617 ppm in sulfides, 

such as in bornite (Cu5FeS4) in the Glava Mine (in Sweden) 

and the Grusen, Moberg, Tinnsjå, and Tjøstølflaten Mines (all 

in Norway) (table Q2 at back of chapter).

In the Dana classification system for minerals (Gaines and

others, 1997), selenium minerals are included in Classes 02 and 03

with sulfides and sulfosalts, respectively. This hierarchical 
Q8 Critical Mineral Resources of the United States—Selenium

system emphasizes the similarity between sulfides and 

selenides with respect to their chemistry, atomic characteristics, and crystal structure. In selenium-bearing hydrothermal 

systems, selenium may commonly substitute into sulfides, 

such as bornite, chalcocite (Cu2S), chalcopyrite (CuFeS2), 

digenite (Cu9S5), galena (PbS), molybdenite (MoS2), pyrite 

(FeS2), and vaesite (NiS2) (Simon and others, 1997).

Significant selenium concentrations can be found in 

sulfide minerals, and numerous studies have been conducted 

to determine whether the selenium-sulfur content represents 

a continuous solid solution between end members, whether a 

compositional gap exists in an apparent solid-solution series, 

or whether the presence of selenium in a sulfide mineral is 

merely an opportunistic isomorphic substitution.

Phases in the silver-gold-sulfur-selenium system show 

ready substitution of selenium for sulfur and, in some cases, 

continuous solid solution exists; for example, from αAg2Se to 

αAg3AuSe2 (α indicates the high temperature structural modification of these phases—naumannite [Ag2

Se] to fischesserite 

[Ag3AuSe2]) (Simon and Essene, 1996). Many mineral phases 

in this system are found in the Broken Hills deposit of the 

Hauraki goldfield in New Zealand. Using electron microprobe 

analysis (EMPA), Cocker and others (2013) observed phases 

with morphologies and compositions similar to the known 

minerals aguilarite and fischesserite, and uytenbogaardtite 

(Ag3AuS2 ) and petrovskaite (AuAg(S,Se)), but with atomic 

sulfur-to-(sulfur+selenium) ratios falling between the known 

members. For example, they observed substitution of selenium 

into uytenbogaardtite, with a sulfur-to-(sulfur+selenium) ratio 

that ranges from 0.55 to almost 1. They noted, however, that 

uytenbogaardtite and fischesserite cannot be end members 

of a solid solution series because the two minerals possess 

different crystal structures (tetragonal and isometric [cubic], 

respectively).

Likewise, solid solution would not be expected between 

the phases acanthite (Ag2S) and aguilarite (Ag4SeS) because 

they belong to different crystal systems. Whereas chemical 

analyses have shown significant substitution of sulfur and 

selenium between these two phases, this may be an analytical 

error owing to their occurrence as fine intergrowths with one 

another. It is unclear whether the chemical analyses come 

from chemically homogenous grains or whether the size of the 

analytical spot size produces a composite analysis. Many of 

the phases in the Broken Hills deposit showed compositions 

that do not match known minerals and may form a continuum 

between approximately (Ag,Au)8

S3 and (Ag,Au)8Se3 (Cocker 

and others, 2013).

Sometimes a solid solution may be present in only one 

polymorph of a mineral composition. For example, solid 

solution may exist between the low-temperature orthorhombic 

mineral structures of naumannite, and aguilarite at temperatures from below 120 to 135 °C. At higher temperatures, they 

invert to their cubic polymorphs (Petruk and others, 1974). 

EMPA analyses of Cocker and others (2013) show a compositional gap between these low-temperature orthorhombic 

phases, but data from Izawa and others (1990) suggest 

that these phases may show solid solution between atomic 

selenium-to-(selenium+sulfur) ratios of 0.57 to 0.88.

Other systems with solid solution compositions 

include lead-sulfur-selenium systems, where solid solution 

has been observed between clausthalite (PbSe) and galena 

(PbS) (Simon and Essene, 1996). Examples are found in 

various uranium-vanadium deposits of the Colorado Plateau 

(Coleman, 1959) and in the Niederschlema-Alberoda uraniumselenium-polymetallic deposit in Erzgebirge, Germany, where 

Förster (2005) documented the entire clausthalite to galena 

solid solution covering the range of PbS1.00 to Pb(S0.04Se0.96).

Complete solid solution can also be found between 

cinnabar (HgS) and tiemannite (HgSe); sphalerite (ZnS) and 

stilleite (ZnSe); and native selenium (Se0) and native tellurium 

(Te0). Quaternary systems show extensive or complete solid 

solution at higher temperatures (Simon and Essene, 1996). 

At 800 °C, Drabek (1995) observed a complete solid solution 

between molybdenite (MoS2 ) and molybdenum diselenide 

(MoSe2) under dry laboratory conditions with a high fugacity 

of selenium to sulfur (f

Se2/fS2 ) ratio.

Selenium-bearing sulfides may form as a replacement 

texture in ore deposits. Paniagua (1995) suggested that 

selenium-rich vaesite (Ni(S,Se)2) and selenium-rich villamaninite ((Cu,Ni,Co,Fe)(S,Se)2) of the Providencia Mine in Leon, 

Spain, are breakdown products of copper-nickel disulfides 

and that these phases show incremental substitution of sulfur 

by selenium with increasing copper+nickel contents because 

of increasing bonding distances within the structure. Many 

studies point to the formation of selenium-bearing sulfides 

during late stages of ore mineralization. For example, in 

epithermal gold-silver deposits in western Nevada, selenium 

enrichment is observed in late-stage stibnite (Sb2S3) and 

sphalerite, and in arsenian rims on earlier formed pyrite and 

marcasite (FeS2; John and others, 2003). John and others 

(2003) conclude that selenides are common in late-stage ores 

found in veins and breccias that fill open-spaces, and form 

from the reaction of late-stage selenium-rich ore fluid (formed 

after substantial boiling and loss of hydrogen sulfide [H2S] to 

vapor) with the initial alteration sulfide assembly.

Because of the close association between selenium and 

sulfur, Simon and Essene (1996) made a systematic thermodynamic study of the stability of selenide minerals relative 

to sulfides, oxides, and native elements. Their conclusions 

were that (a) the stability fields of selenides, in general, are 

more restricted than sulfide minerals, given similar fugacity 

values of S2(g) and Se2(g) (represented by fSe2(g) and fS2(g), 

respectively); (b) in hydrothermal fluids, most oxide minerals 

are unstable relative to their corresponding selenide mineral; 

and (c) for a given selenide, the larger the stability field, the 

more commonly it is found in ore deposits; for example, 

clausthalite.

In a companion paper to the above, Simon and others 

(1997) discussed the thermodynamic conditions that favor 

deposition of selenide minerals instead of selenium-substituted 

sulfide minerals. They conclude that an oxidizing environment is needed to form most selenide minerals. High oxygen 
Geology Q9

fugacity values (fO2(g) ) enhance separation of selenium from 

sulfur, leading to high fSe2(g)/fS2(g), which enables deposition 

of selenide minerals and limits the formation of sulfide 

minerals. In contrast, mineralizing systems characterized by 

low oxygen fugacity and the presence of sulfide minerals can 

deposit only silver selenides and selenium-substituted sulfide 

minerals.

Although some selenium is produced from gold and 

silver selenide deposits that have anomalously high concentrations of selenium (table Q2), most of the world’s production 

of selenium metal is from copper, copper-nickel, and some 

lead and zinc sulfide deposits where selenium is found 

in isomorphous substitution with sulfur (Bleiwas, 2010). 

Table Q2 illustrates the range of selenium concentrations 

found in sulfur and selenium minerals worldwide.

Deposit Types

The compilation of selenium deposits provided by 

Sindeeva (1964) remains one of the most exhaustive in 

today’s literature. Because the information is well organized 

and informative in its original form, the reader is referred 

to table 7 (p. 174) in “Mineralogy and Types of Deposits of 

Selenium and Tellurium” (modified as table Q3 in this report). 

Here, the author noted that selenium can be found in several 

genetic deposit types—magmatic, volcanic, hydrothermal, 

and exogenic (that is, produced by chemical and mechanical 

weathering at Earth’s surface; Jackson, 1997). In Sindeeva’s 

table, examples from hydrothermal deposits outnumber the 

rest combined. In the majority of the examples, selenium is 

found substituted into sulfide minerals, such as chalcopyrite, 

cinnabar, cobaltite (CoAsS), galena, molybdenite, pentlandite 

((Fe,Ni,Co)9S8), pyrite, pyrrhotite (Fe1–xS), sphalerite, and 

stibnite. Selenide minerals are less common than seleniumbearing sulfides, but such phases as clausthalite (PbSe), 

guanajuatite (Bi2Se3), naumannite (Ag2Se), and penroseite 

(blockite) ((Ni,Co,Cu)Se2), are found in veins and stockworks 

of hydrothermal deposits. The compilation includes some 

deposits where selenium can be found as selenites, selenates, 

and native selenium (for example, volcanic tuffs in Shoshone 

and Riverton, Wyoming).

As a result of a thorough investigation into selenium 

content of ore deposits and its mineral associations, Sindeeva 

(1964) concluded the following:

• Deposits containing independent selenium minerals 

are rare; instead, selenium usually occurs as an 

isomorphous substitution for sulfur in sulfide 

minerals.

• The largest resources of selenium are connected with 

ultrabasic and basic rocks and found in magmatic 

and hydrothermal chalcopyrite-pentlandite-pyrrhotite 

deposits, especially in sulfide veins that occur during 

late stages of formation.

• Pyrite deposits contain large resources of selenium, 

which seem to accompany late stages of mineralization.

• Copper and molybdenum deposits contain selenium, 

where the enrichment occurred in late-stage molybdenites.

• Polymetallic lead-zinc deposits can contain selenium.

• Gold deposits can contain selenium. Selenium will 

form independent minerals in these deposits if the 

ore contains very little sulfur; otherwise, it will be 

dispersed in sulfide minerals.

• Quartz-wolframite ((Fe,Mn)WO4)-bismuth and cassiterite (SnO2

)-quartz-sulfide deposits are sometimes 

accompanied by selenium.

• Cobaltite-selenide-telluride resources that contain 

selenium are not frequently observed, but selenium can 

be associated with chloritization in this assemblage.

• Uranium deposits carry selenium but the co-occurrence 

of these elements could use more study.

• Cinnabar-stibnite deposits might contain appreciable 

selenium and could use more study.

Magmatic and Associated Hydrothermal Deposits

Aside from selenium minerals, selenium-bearing 

sulfide deposits represent the greatest known concentration 

of selenium in Earth’s crust. The reported average selenium 

concentration in magmatic sulfides is 0.02 percent (Sargent, 

1954). In intrusive rocks, selenium will substitute for sulfur 

in sulfide minerals because of their similar ionic radii. Hightemperature sulfides, such as pentlandite and pyrrhotite, have 

a sulfur-to-selenium ratio of 7,000:1, and Goldschmidt (1954) 

estimated hydrothermal pyrite to have a sulfur-to-selenium ratio 

of between 10,000 and 20,000:1. In the Falun mining area of 

Sweden, selenium is present in the following sulfide minerals, in 

decreasing order of concentration: galena, chalcopyrite, arsenopyrite, sphalerite, pyrite, and pyrrhotite (Sindeeva, 1964).

Attempts have been made to standardize the estimate of 

selenium in sulfide ore minerals and the most commonly used 

number for the average selenium content in magmatic and 

associated hydrothermal sulfides is 0.02 percent. Other ratios 

exist; for example, in porphyry copper ore with 0.5 percent 

copper, the selenium content is 2.5 ppm, or a copper-toselenium ratio of 2,000:1 (Sargent, 1954; Lakin and Davidson, 

1973). The U.S. Geological Survey and the former U.S. Bureau 

of Mines have long based their selenium production estimates 

on copper production figures using a factor of 0.0215 percent. 

Because copper ore from Canada has a higher selenium 

content, however, a factor of 0.064 percent is used to estimate 

selenium produced from its ores (for example, selenium in 

ore from mines located in the Sudbury basin of Ontario, 

Canada (table Q4; Jensen, 1985; George and Wagner, 2009). 
Q10 Critical Mineral Resources of the United States—Selenium

Table Q3. Selenium concentrations in sulfide minerals and other phases, in various deposit types.—Continued

[Data are from Sindeeva, 1964. Do, do., ditto; wt. %, weight percent; n.r., not reported, although the deposit was described as containing selenium and (or) 

selenide minerals; b.d.l., below detection limit; Hg, mercury; S, sulfur; Se, selenium]

Deposit type Deposit name, Location Mineral phase

Selenium 

concentration

(weight percent 

unless otherwise 

indicated)

Hydrothermal

Chalcopyrite-molybdenite Agarak deposit, Armenia Chalcopyrite 0.004

 Do. do. Molybdenite 0.0131 to 0.0134

 Do. do. Pyrite 0.006

 Do. Dastakert deposit, Armenia Bornite 0.0066

 Do. do. Chalcopyrite-pyrite 0.003

 Do. do. Galena 0.0057

 Do. do. Molybdenite 0.0016

 Do. do. Pyrite 0.0036

 Do. do. Sphalerite n.r.

 Do. Dzhindara deposit, Armenia Molybdenite 0.04

 Do. Kadzharan deposit, Armenia Chalcopyrite 0.0052

 Do. do. Molybdenite 0.028 to 0.0513

 Do. do. Pyrite 0.0015 to 0.0063

 Do. Miskhana deposit, Armenia Molybdenite 0.0056

 Do. do. Pyrite 0.008

Cinnabar-stibnite Almaden, Idria, Spain Mercury ores may contain

Hg,Se-tiemannite, sometimes

in an isomorphous series

with metacinnabar

n.r.

 Do. San Onofrio, Mexico Onofrite, Hg(S,Se) 1:3.9 S:Se ratio

Cobaltite-selenide-telluride Upper Seymchanskoye deposit, 

Verina River basin, Russia

Selenide ore 0.0001 to 0.002

Gold-selenium Eisenberg, near Korbach, Germany Clausthalite (common) n.r.

Gold-silver-quartz-adularia Comstock and Goldfield, Nevada, 

United States

n.r. n.r.

 Do. El Oro, Mexico n.r. n.r.

 Do. Redzhang Lebong deposit, Sumatra, 

Indonesia

n.r. 0.02

Pyritic Uchaley deposit, Russia Pyrite with chalcopyrite 

and sphalerite

0.0004 to >0.01

Quartz-gold ore Kirovskoye deposit, Russia Arsenopyrite 0.0006 to 0.019

 Do. do. Bismuthinite 0.2800 to 0.3275

 Do. do. Chalcopyrite 0.0008 to 0.003

 Do. do. Jamesonite ≤0.003

 Do. do. Pyrite ≤0.003
Geology Q11

Table Q3. Selenium concentrations in sulfide minerals and other phases, in various deposit types.—Continued

[Data are from Sindeeva, 1964. Do, do., ditto; wt. %, weight percent; n.r., not reported, although the deposit was described as containing selenium and (or) 

selenide minerals; b.d.l., below detection limit; Hg, mercury; S, sulfur; Se, selenium]

Deposit type Deposit name, Location Mineral phase

Selenium 

concentration 

(weight percent 

unless otherwise 

indicated)

Hydrothermal—Continued

Quartz-wolframite-bismuth

 Do.

 Do.

Bukuka deposit, Russia

 do.

 do.

Bismuthinite

Chalcopyrite

Galena

0.016

0.0012

0.0116

 Do.

 Do.

 Do.

 Do.

 Do.

 do.

 do.

 do.

 do.

 do.

Molybdenite

Pyrite

Pyrrhotite

Sphalerite

Wolframite

0.006

0.0175

0.0013

0.0017

b.d.l.

Selenide

Uranite-selenide

 Do.

Pacajake deposit, Bolivia

Katanga deposit, Democratic Republic 

of the Congo (Congo [Kinshasa])

Shinkolobse deposit, Democratic 

Republic of the Congo (Congo 

[Kinshasa])

Blockite (penroseite) is main 

ore mineral, also naumannite; 

mined for gold in World War II

Reported to contain selenides

Selenide minerals in 114th horizon

n.r.

n.r.

19

Magmatic

Magmatic

 Do.

 Do.

 Do.

Noril’sk, Siberia, Russia

 do.

 do.

 do.

Bornite

Chalcopyrite

Magnetite 

Millerite

0.01 to 0.015

0.005 to 0.020

Trace

0.001

 Do. do. Pentlandite 0.0067

 Do.

 Do.

 do.

 do.

Pyrite

Pyrrhotite

Trace

0.0036 to 0.014

Volcanic

Andesite tuffs

Native sulfur

Shoshone, Jasper, and Riverton deposits, Wyoming, United States

Hawaiian Islands

n.r.

n.r.

0.003 to 0.012

5.18

 Do. Kurile Islands, additional 

Kamchatka locations, Russia

n.r. 0.0005 to 0.19

 Do.

 Do.

Lipari Islands , Italy

Mendeleev Volcano, Kunashir Island, 

Russia

n.r.

n.r.

1.13

0.12

 Do. Mutnovskii Volcano, Kamchatka, 

Russia

n.r. 0.08

 Do. New Zealand n.r. 0.3 and 0.19

 Do. Paramushir Island, Russia n.r. 0.19
Q12 Critical Mineral Resources of the United States—Selenium

Exceptions can always be found to these general estimates; 

for example, selenium can be present in concentrations of 

hundreds of parts per million in tuffaceous rocks, and concentrations of several percent can be found in sulfides that host 

mercury and antimony deposits (Lakin and Davidson, 1973).

In magmatic sulfide deposits associated with mafic 

rocks, selenium may be found substituted for sulfur in sulfide 

veins. Sindeeva (1964) estimated the selenium content in such 

deposits to vary between 0.002 and 0.01 percent, although 

concentrations of as high as 0.017 percent and as low as 

0.0003 percent have been found. The Permian-to-Triassicage Noril’sk group of sulfide deposits, which are located in 

Siberia, Russia, inside the Arctic Circle, are representative 

of this type of mineralization (table Q3), as are the sulfide 

deposits from the Sudbury basin in Ontario, Canada 

(table Q4).

At Noril’sk, the selenium of economic interest is 

contained in sulfide minerals, where it occurs from isomorphous substitution for sulfur in concentrations of 2 to 74 ppm. 

Massive vein deposits at Noril’sk display greater concentrations of selenium than its disseminated ores, with a range 

of 0.0036 to 0.015 percent selenium in massive vein deposits 

owing to the greater proportion of copper sulfides in the vein 

deposits; however, no correlation has been found between 

selenium and any of the elements copper, lead, nickel, sulfur, 

or tellurium. Distribution of selenium depends upon the 

composition of the ore as well as the location of the sample 

within the orebody and with respect to the structural setting 

(table Q3). The rocks surrounding the ore deposits (andesites, 

diabases, and gabbros) contain merely 0.00001 percent 

(0.1 ppm) selenium (Sindeeva, 1964).

In addition to magmatic sulfide ores, Sindeeva (1964) 

considered the volcanic deposits of native sulfur and tuffs. She 

observed selenium concentrations of 0.0005 to 5.18 percent 

in sulfur deposits and 0.003 to 0.012 percent selenium in 

andesite tuffs that had partially transformed to bentonite in 

the Oligocene-age Jasper, Riverton, and Shoshone deposits in 

Wyoming (table Q3).

The porphyry copper deposits that host much of the 

selenium produced worldwide are large breccia and veinhosted deposits found genetically related to granitoid porphyry 

intrusions. They form at convergent plate margins, and most 

are of Cenozoic or Mesozoic age, although they can range 

from Archean to Quaternary (John and others, 2010). John 

and others (2010) argued that porphyry copper mineralization 

forms from magmatic fluids released during emplacement of 

porphyritic stocks. Selenium in sulfides of porphyry copper 

deposits is originally transported in hypogene fluids but may 

also be remobilized during hydrothermal alteration of the 

porphyry copper system.

Many workers have attempted to employ the seleniumto-sulfur ratio in the ore minerals to ascertain whether 

external sulfur was involved in the genesis of mineralization. 

Selenium-to-sulfur ratios ranging from 0.00023:1 to 0.00035:1 

have been assigned to mantle-derived rocks (Eckstrand and 

Hulbert, 1987), and ratios lower than this are found when 

sulfur is contributed from external fluids and (or) host rocks 

(for example, Ripley, 1990). Fitzpatrick (2008) measured 

selenium-to-sulfur ratios in minerals from three deposits: a 

volcanogenic massive sulfide (VMS) deposit in Flin Flon, 

Canada; a high-sulfidation epithermal deposit in Pierina, Peru; 

and an iron oxide-copper-gold (IOCG) deposit in Mantoverde, 

Chile. By combining the selenium-to-sulfur ratios with δ34S 

values, Fitzpatrick demonstrated that the VMS and epithermal 

deposits displayed high selenium-to-sulfur ratios and low δ34S 

values, which is typical of magmatic fluids; the IOCG deposit, 

however, showed the opposite signal, suggesting dilution of 

a magmatic fluid with a seawater-derived basinal fluid. He 

concluded that selenium-to-sulfur ratios can help identify the 

fluid source(s) for mineralization, and that the addition of δ34S 

values extended the usefulness of this tracer.

Other studies have employed selenium as an exploration 

‘vector’ to discriminate among mineralization signatures found 

proximal or distal to a fluid source. Bornite and chalcocite 

show some potential for this work (Cook and others, 2011). 

Zonation of selenium concentration has also been observed in 

marine systems, and high selenium concentrations have been 

observed in black smoker deposits and iron-copper massive 

sulfides associated with the East Pacific Rise. Here, high selenium values occur in high-temperature mineral assemblages 

of the inner part of the deposit, which are in equilibrium with 

pure, unmixed hydrothermal fluids. The lowest selenium 

values are found in the outer part of the deposit, where 

mineralization is influenced by mixing of hydrothermal fluids 

and seawater (Auclair and others, 1987).

Table Q4. Selenium concentrations in copper-nickel ores in the 

Sudbury basin, Ontario, Canada.

Mine 

name Ore type1

Selenium

concentration2

(parts per million)

Murray Massive sulfide 32 to 64

Gabbro peridotite inclusion sulfide 82

Creighton Interstitial sulfide 26

Ragged disseminated sulfide 20

Gabbro peridotite inclusion sulfide 42

Massive sulfide 65

Garson Contorted schist inclusion sulfide 22

Levack Interstitial sulfide in sublayer norite 25

Frood Disseminated sulfide in quartz diorite 28

Interstitial sulfide 36

Ragged disseminated sulfide 20

1

Ore type classification is from Schwarcz (1973).

2

Selenium analysis is from Nriagu and Wong (1983).
Geology Q13

Selenide-Bearing Deposits

In contrast to Sindeeva (1964), who discussed all 

selenium-bearing deposits whether they contained seleniumbearing sulfides, selenides, selenates, selenites, or native 

selenium, Simon and others (1997) discussed deposits that 

contained selenide minerals only. Their purpose was to 

“determine the conditions that favor deposition of selenide 

minerals over substitution of selenium in sulfide minerals” 

(Simon and others, 1997, p. 468). Deposits were grouped into 

the following four types: (a) telethermal selenide vein-type, 

(b) unconformity-related uranium deposits, (c) sandstonehosted uranium deposits (roll-front type), and (d) gold-silver 

epithermal volcanic-hosted deposits. The following brief 

descriptions of these four deposit types are taken mostly 

from Simon and others (1997):

a. Telethermal selenide vein-type deposits include 

selenide-bearing deposits like those of the TilkerodeZorge-Lerbach deposit in the Harz Mountains of 

Germany, the Pacajake and El Dragon deposits in 

Bolivia, the Hope’s Nose deposit in the United 

Kingdom, and the Sierra de Umango and Sierra de 

Cacheuta deposits in Argentina. In these deposits, 

selenide minerals occur in veins or veinlets with 

carbonate minerals, gold, hematite (Fe2

O3), platinumgroup elements, and rare sulfides, and are commonly 

hosted in sedimentary rocks. Some of these deposits 

were classified as the “selenide-type” by Sindeeva 

(1964). Conditions for formation of these deposits 

include fO2(g) above the hematite-magnetite buffer (that 

is, in the hematite stability field) and a ratio of 

selenium-to-sulfur fugacities that is greater than 

one (fSe2(g) /fS2(g) > 1). Simon and others (1997) argue 

that the high fO2 value helps to separate selenium from 

sulfur, thus preventing the incorporation of selenium into 

sulfides and allowing the system to 

form selenide minerals.

b. Selenide minerals in unconformity-related uranium 

deposits are found in vein deposits in the Bohemian 

Massif of the Czech Republic, the Massif Central of 

France, and the Athabasca area of northern Saskatchewan, Canada. Cobalt, copper, and nickel selenides in 

these deposits are associated with carbonates; cobalt, 

copper, and nickel sulfides and arsenides; hematite; 

and uraninite minerals. As with the vein-type deposits 

above, the conditions for selenide formation include 

fO2(g) values within the hematite stability field and a ratio 

of selenium-to-sulfur fugacities that is greater than one 

(fSe2(g) /fS2(g) > 1).

c. Selenium-bearing sandstone-hosted uranium deposits 

(roll-front-type) include the uranium and vanadium 

sandstone deposits of the Colorado Plateau, and similar 

deposits in Oklahoma, Wyoming, and Texas (Davidson, 

1963; Howard, 1977). Davidson (1963) surveyed 

uranium deposits in oxidized sandstones ranging from 

Paleozoic to Cenozoic age in several States across the 

United States and found selenium concentrations that 

ranged from 3 to 4,500 ppm. In a survey by Coleman 

and Delevaux (1957), selenium concentrations up to 

3 weight percent were measured in uranium deposits 

of the Colorado Plateau.

In the roll-front uranium deposits, selenium is 

commonly found in isomorphic substitution for sulfur in 

chalcocite, galena, marcasite, and pyrite. It is also found 

as the selenide minerals ferroselite (FeSe2), clausthalite, 

and native selenium. Howard (1977) provided a good 

description of the location of selenium mineralization 

with respect to the redox interface of a roll-front-type 

uranium orebody. He described the uranium ore deposits 

as forming in gently dipping permeable sandstone 

units, sandwiched between units of lower permeability. 

Mineralization occurred as oxidizing meteoric solutions 

moved through the host sandstone and dissolved pyrite 

and other redox-active minerals. At the front of the 

plume of meteoric water, the solution lost its oxidizing 

potential and deposited iron, selenium, and uranium at 

the interface between oxidizing and reducing environments. In this model, selenium is commonly found as 

native selenium in the sandstone on the oxidized side 

of the redox boundary, as ferroselite in both goethite 

(FeO(OH))-hematite and pyrite-marcasite associations 

on either side of the redox interface, and as ferroselite 

and native selenium within the uraninite-coffinite 

(U(SiO4

)1–x(OH)4x) assemblages of the orebody. In 

the unaltered reduced rock, selenium will be found as 

ferroselite or in isomorphous substitution with sulfur 

in pyrite. It is the continued cycle of oxidative dissolution (of reduced minerals in the unaltered rock) and 

redeposition (at the redox interface at the front of the 

groundwater plume) that serves to concentrate selenium, 

iron, and the uranium-ore minerals of these deposits 

(Howard, 1977).

d. Selenide-bearing gold-silver epithermal volcanic-hosted 

deposits are found at Hishikari, Japan; the western Great 

Basin, Nevada; the Great Barrier Island, New Zealand; 

Salida and Redzhang Lobong, Indonesia; Guanajuato, 

Mexico; and Prasolovskoye, Kunashir Island, Russia. 

Selenium in these deposits is primarily found in silver 

selenides and, less commonly, in bismuth, copper, and 

lead selenides, all associated with sulfides, sulfosalts, 

and native elements. John (2001) observed that the silver 

selenides aguilarite and naumannite are found in almost 

all low-sulfidation epithermal gold-silver deposits associated with Miocene bimodal magmatism in the Great 

Basin, Nevada. Selenium-bearing stibnite and sphalerite 
Q14 Critical Mineral Resources of the United States—Selenium

are found among the ore-stage minerals, and selenium 

can be enriched in the gold-bearing arsenian rims found 

on pyrite and marcasite. It is thought that these deposits 

formed from hydrothermal fluids with fO2(g) below the 

hematite-magnetite buffer and have a ratio of seleniumto-sulfur fugacities that is less than one (fSe2(g) /fS2(g) 

< 1) 

(Simon and others, 1997). These conditions lead to a 

greater abundance of sulfides than selenides.

Sedimentary Rocks

The substitution of selenium for sulfur in biological 

processes leads to the presence of selenium in pyrite contained 

in black shales and coal, yet compared with magmatic rocks, 

the sulfur-to-selenium ratio in diagenetic pyrite is estimated 

to be 200,000:1 or higher (Goldschmidt, 1954). Selenium in 

detrital organic matter may be enriched many times above 

its crustal abundance. Highly organic shales may contain up 

to 1,500 ppm selenium. Marine phosphorites may contain as 

much as 300 ppm selenium. Coals of the United States have 

a geometric mean concentration of 1.7 ppm (Coleman and 

others, 1993).

Selenium concentrations in coal have been a topic 

of interest for many decades owing to the release of trace 

elements to the atmosphere during coal combustion. Coal 

contains 5 to 300 times the amount of selenium as average 

sedimentary and igneous rocks, and it is enriched 82 times 

above the average selenium concentration in Earth’s crust 

(table Q5; Coleman and others, 1993). The study by Coleman 

and others (1993) analyzed almost 9,000 coal samples from 

across the United States (fig. Q5) and found that the highest 

selenium values (with a geometric mean of 4.9 ppm) are 

found in coals from Texas. Coal found in the Atlantic Coast 

region (Virginia and North Carolina) and coal in Alaska have 

the lowest geometric means, at 0.2 and 0.42 ppm selenium, 

respectively. The highest concentration in an individual 

sample (75 ppm) was observed in coals from northern 

Appalachia. Because of the close association of selenium 

with sulfur, it might be expected that selenium concentrations 

would correlate with concentrations of pyritic sulfur or organic 

sulfur, but this was not generally observed throughout the 

dataset. Coleman and others (1993) speculated that the lack 

of correlation may be owing to multiple sources of selenium, 

including detrital particles, plant matter, volcanic ash, surface 

water and groundwater, and epigenetic mineralization. They 

note, in addition, that selenium concentrations can be modified by diagenetic processes. Whereas selenium is shown 

to be closely associated with the organic fraction in coal, it 

is also found in selenium-bearing pyrite and galena, and as 

clausthalite. Selenium is also found in water-soluble and ion 

exchangeable forms.

Much of the Western United States is underlain by 

seleniferous sedimentary rocks whose ages range from 

Permian (for example, the Phosphoria Formation, which has 

phosphate rock members containing 1 to 1,200 ppm selenium), 

to Cretaceous (for example, the Niobrara Formation and the 

Pierre Formation, which have selenium concentrations of 

0 to 113 ppm, and 0 to 103 ppm, respectively) to Upper Cretaceous-Paleogene (for example, the Moreno Formation and 

the Kreyenhagen Formation of the California Coast Ranges, 

which have selenium concentrations that together range from 

0.6 to 45 ppm) (table Q5; figs. Q6A–C; Presser and others, 

1990). Trelease and Beath (1949) extensively mapped and 

documented the widespread extent of seleniferous bedrock in 

the Western United States as part of their study of the source 

of selenium in livestock food and forage (figs. Q6A–B). 

Seiler and others (2003) extended this survey in their work 

to identify the source of selenium contamination to irrigation 

waters in the Western United States (fig. Q6C). Presser and 

others (2004) attributed the high selenium concentration in 

these sedimentary bedrock units to its depositional origin 

in marine basins with high primary biological productivity. 

Selenium is cycled from ocean waters into bacteria, algae, 

fungi, and plants with the production of selenium-containing 

amino acids. Deposition of these species as organic detritus 

on the sea floor, and subsequent loss of organic matter from 

diagenesis resulted in accumulation and concentration of 

selenium and other trace elements in the sediments. This is the 

general model for trace-element enrichment in the Permian 

Phosphoria Formation (Piper, 1994, 2001) as well as in the 

Cretaceous and Paleogene marine formations of the Western 

United States (Presser and others, 2004). Preservation of 

selenium and other trace elements in these rocks is attributed 

to their high initial concentrations, retention by the sediments 

(often in an organic fraction), and extremely low accumulation 

of other diluting phases.

Resources and Reserves

More than 80 percent of global production of selenium 

is obtained as a byproduct of copper refining; it is recovered 

from anode slimes generated in electrolytic production 

of copper (Brown, 2002). The blister copper (that is, the 

partly refined copper formed during smelting) that serves 

as the anode in an electrolytic cell contains an average 

of 0.05 percent selenium (0.5 kg of selenium per metric 

ton of copper), but recovery ranges from 0.02 percent to 

0.038 percent (0.2 to 0.38 kg of selenium per metric ton 

of copper; Lakin and Davidson, 1973). Other methods of 

refining copper, such as in situ leaching or the ISASMELTTM 

technology, do not recover selenium; hence, the global 

production of selenium is dependent on the method of copper 

refining. Because selenium production is mainly a byproduct 

of copper refining, countries with selenium resources and (or) 

reserves are those with resources and (or) reserves of copper, 

and they include Canada, Chile, China, Congo (Kinshasa), 

Mexico, Peru, the Philippines, Russia, the United States, and 

Zambia (Jensen, 1985; John and others, 2010). Minor amounts 

of selenium have also been recovered from lead, nickel, and 

zinc ores.
Resources and Reserves Q15

1

2

3

5

4

6 7

8

9

10

11

12

13

14

15

19

18

23

24 16

17

22

21 20

 1 Rhode Island Meta-anthracite —

 2 Pennsylvania Anthracite 2.6

 3 Atlantic Coast —

 4 Appalachian 2.7

 5 Northern Interior 2.3

 6 Eastern Interior 2.3

 7 Western Interior 2.4

 8 Southwestern Interior —

 9 Mississippi 4.4

 10 Texas 4.9

 11 Fort Union 0.7

 12 Powder River 0.8

Region number and name Types of coal in fields

EXPLANATION

Anthracite

Low-volatile bituminous

Medium- and high-

 volatile bituminous

Subbituminous

Lignite

Selenium

concentration,

in ppm

 13 Black Hills —

 14 North Central 1.4

 15 Tertiary Lake Beds —

 16 Bighorn Basin —

 17 Wind River 1.6

 18 Hams Fork 1.1

 19 Uinta 1.3

 20 Southwestern Utah 1.1

 21 San Juan River 1.9

 22 Raton Mesa 2.3

 23 Denver 1.5

 24 Green River 1.1

Region number and name

Selenium

concentration,

in ppm

0 250 500 KILOMETERS

0 125 250 MILES

Selenium figure Q6

125°

45°

35°

115° 105° 95° 85° 75° 65°

25°

Figure Q5. Map showing selenium concentrations (in parts per million) in coal samples, by region of the United States. 

Selenium concentrations were analyzed in almost 9,000 U.S. coal samples. Data are plotted as the geometric means of 

samples from 24 regions. The map shows the highest selenium values (with a geometric mean of 4.9 ppm) are found in 

coals from Texas. Coal found in the Atlantic Coast region (Virginia and North Carolina) and coal in Alaska (not shown 

on the map) have the lowest geometric means, at 0.2 and 0.42 ppm selenium, respectively. The highest concentration in 

an individual sample (75 ppm) was observed in coals from the northern portion of the Appalachian region. Map is from 

Coleman and others (1993).
Q16 Critical Mineral Resources of the United States—Selenium

Table Q5. Selenium concentrations in selected Earth and lunar materials.—Continued

[Biologic samples are dry weight unless otherwise noted. Abbreviations: ppm, part per million; ppb, part per billion; ng/L, nanogram per liter; nmol/kg, 

nanomole per kilogram]

Material Source of material

Selenium 

concentration

 (and additional

explanation)

Unit Reference(s)

Rocks

Earth’s crust Continental crust 0.05 ppm Taylor (1964, p. 1280)

Earth’s crust Granite 0.01 to 0.05 ppm Kabata-Pendias and Pendias (1984, p. 186)

Earth’s crust Limestone 0.03 to 0.10 ppm Kabata-Pendias and Pendias (1984, p. 186)

Earth’s crust Phosphate rock 1 to 300 ppm McNeal and Balistrieri (1989, and references therein)

Earth’s crust Sandstone 0.05 to 0.08 ppm Kabata-Pendias and Pendias (1984, p. 186)

Earth’s crust Shale 0.6

(average)

ppm Adriano (1986, p. 392)

Coal World 3 ppm Minkin and others (1984)

Coal China 3.65 ppm Zhang and others (1999)

Coal United States 0.46 to 10.65 ppm Pillay and others (1969, p. 480)

Coal United States 1.7

(geometric mean)

ppm Coleman and others (1993)

Marine shale Kreyenhagen Formation 8.7

(median)

ppm Presser (1994, p. 145)

Marine shale Mancos Formation 0.1 to 5.46 ppm Tuttle and others (2007, appendix tables)

Marine shale Mead Peak Member, 

Phosphoria Formation

1 to 1,200; 182

(average)

ppm Presser and others (2004)

Marine shale Moreno Formation 6.5 (median) ppm Presser (1994, p. 145)

Marine shale Niobrara Formation 0 to 30 ppm Moxon and others (1938, p. 797–799)

Marine shale Pierre Formation 0 to 103 ppm Moxon and others (1938, p. 799–802)

Moon rocks Apollo 11 0.12 to 1.6 ppm Leutwein (1972, p. 34-C-3)

Moon rocks Apollo 16 0.40 to 314 ppb Krähenbühl and others (1973, p. 1328)

Soils and dusts

Soil Worldwide, seleniferous, 

upper limit

1,200 ppm Winkel and others (2012, p. 572)

Soil Worldwide, selenium-deficient, 

lower limit

<0.01 ppm Winkel and others (2012, p. 572)

Soil England and Wales <0.01 to 4.7 ppm Thornton and others (1983; as cited in McNeal and Balistrieri, 1989)

Soil Finland, near a Cu-S orebody 0.1 to 5.5 ppm Koljonen (1976, p. 263)

Soil Ireland, seleniferous ≤1,200 ppm Fleming (1962, p. 30)

Soil Italy, volcanic 0.12 to 2.25 ppm Floor and others (2011, p. 238)
Resources and Reserves Q17

Table Q5. Selenium concentrations in selected Earth and lunar materials.—Continued

[Biologic samples are dry weight unless otherwise noted. Abbreviations: ppm, part per million; ppb, part per billion; ng/L, nanogram per liter; nmol/kg, 

nanomole per kilogram]

Material Source of material

Selenium 

concentration

 (and additional 

explanation)

Unit Reference(s)

Soils and dusts—Continued

Soil

Soil

Soil

Soil

Soil

Soil

Atmospheric dust

Sweden, agricultural

Sweden, background, not affected by agriculture

United States

United States

United States, seleniferous

United States, Hawaii, volcanic

Major U.S. cities, 1941

0.065 to 1.7

<0.45

<0.1 to 4.3

<0.2 to 8.3

2.7 to 90

1 to 20 

0.05 to 10

ppm

ppm

ppm

ppm

ppm

ppm

ppm

Johnsson (1992, section 4.1)

Johnsson (1992, section 4.1)

Shacklette and Boerngen (1984, p. 4); 

Tidball (1984, p. H6)

Smith and others (2014, p. 197–202)

Trelease and Beath (1949, p. 111–112)

Floor and Román-Ross (2012, p. 521)

Lakin and Byers (1941, p. 20)

Biologic samples

Algae

Animal tissue

Whole fish

Whole fish

Freshwater

Domestic stock

Freshwater

Marine

<2.0

0.44 to 4.0

0.05 to 2.87 

(wet weight)

0.1 to 20

ppm

ppm

ppm

ppm

Jenkins (1980; as cited in McNeal and Balistrieri, 1989)

Frost (1972, p. 471)

May and McKinney (1981, p. 35)

Hall and others (1978, p. 94–98)

Water

Seawater

River water

River water

River water

River water

Lake water

Lake water

Lake water

Lake water

Lake water

Pacific Ocean, 18° N., 108° W.

Amazon River

Colorado River, United States

Finland, stream sediments

Mississippi River, United States

Finland, lakes affected by 

agriculture 

Finland, lakes, background, not 

affected by agriculture runoff

Sweden

United States, Lake Michigan

Finland, lake sediment

0.95 to 2.44

0.00021

0.01 to 0.4

0.26

0.000114

76.8

56 to 62

43 to 209

0.083

0.27 to 3.64

nmol/kg

ppm

ppm

ppm

ppm

ng/L

ng/L

ng/L 

ppb

ppm

Cutter and Bruland (1984, p. 1183)

Kharkar and others (1968, p. 290)

Scott and Voegeli (1961; as cited in 

McNeal and Balistrieri, 1989)

Wang (1994; as cited in Parkman and Hultberg, 2002)

Kharkar and others (1968, p. 290)

Wang (1994; as cited in Parkman and Hultberg, 2002)

Wang (1994; as cited in Parkman and Hultberg, 2002)

Parkman and Hultberg (2002)

Robberecht and Van Grieken (1982, p. 839)

Wang (1994; as cited in Parkman and Hultberg, 2002)
Q18 Critical Mineral Resources of the United States—Selenium

In 1973, the USGS estimated the world’s identified 

resources of selenium to be almost 13 million metric tons 

(28.6 billion pounds) based upon the selenium content of 

identified resources of coal, copper, lead, and pyrite. Estimates 

from copper resources were made using a ratio of 0.325 kg 

of selenium recovered per metric ton of copper produced. 

Selenium from lead ores was estimated based upon an average 

of 10 ppm selenium in galena and the estimate of identified 

lead resources. Estimates for selenium in pyrite were made 

assuming 10 ppm selenium in pyrite and estimates of the 

sulfur resources in pyrite. Selenium resources in coal were 

estimated using estimates for coal resources and an average 

of 1.5 ppm Se in coal. Lakin and Davidson (1973) noted that 

although the estimated resource of selenium in coal was two 

orders of magnitude greater than that in copper resources, it 

could not be easily recovered. As of 2016, the recovery of 

selenium from coal is technically feasible but not economical.

In contrast, modern estimates of selenium reserves 

in table Q6 are based only on copper reserves and do not 

include selenium found in other copper resources or in lead, 

Selenium figure Q7A

EXPLANATION

Cretaceous, seleniferous, marine

 sedimentary outcrops

Plant sample with ≥50 ppm selenium

A

0 100 200 300 MILES

0 100 200 300 KILOMETERS

125°

45°

35°

115° 105° 95°

25°

WA

ID MT

WY

OR

CA NV

UT

AZ NM

TX

KS

SD

ND

MN

IA

MO

AR

LA

NE

CO

OK

Figure Q6. Maps showing locations of seleniferous sedimentary outcrops and deposits and plant samples with significant 

selenium content. This map, A, shows the spatial relation among locations of Cretaceous, seleniferous, marine sedimentary 

outcrops and plant samples with greater than or equal to 50 parts per million (ppm) selenium. Map A is from Trelease and 

Beath (1949) and, with map B, shows the widespread extent of seleniferous bedrock across the Western United States and 

its correlation with selenium in plant samples.
Resources and Reserves Q19

pyrite, and coal resources. The fraction of copper reserves 

expected to be recovered by leaching and electrowinning 

is subtracted from the total because these processes do not 

recover selenium. The estimates are derived, as mentioned 

previously, by using a recovery factor of 0.215 kg of selenium 

per metric ton of primary recoverable electrolytic copper 

(0.0215 percent). In Canada, where the selenium content of 

copper ore is higher, a factor of 0.64 kg of selenium per metric 

ton of copper (0.064 percent) is employed (Butterman and 

Brown, 2004; George and Wagner, 2009). Global supplies 

of selenium have steadily increased over the past few years 

owing to an increase in copper production.

Selenium figure Q7B

EXPLANATION

Paleocene and Eocene seleniferous, 

 marine sedimentary outcrops

Plant sample with ≥50 ppm selenium

B

0 100 200 300 MILES

0 100 200 300 KILOMETERS

125°

45°

35°

115° 105° 95°

25°

WA

ID MT

WY

OR

CA NV

UT

AZ NM

TX

KS

SD

ND

MN

IA

MO

AR

LA

NE

CO

OK

Identified Resources

No deposits in the world are mined for selenium alone. 

Instead, as discussed above, selenium is produced mostly as 

a byproduct of copper refining. Whereas Japan and Germany 

(in order of output) were the leading producers of selenium 

in the world in 2015 (U.S. Geological Survey, 1997–2014), 

the selenium was sourced from copper concentrates and 

residues supplied from mining operations in Africa, Asia, 

Australia, and South America (Bleiwas, 2010). Of the global 

reserves of 120,000 metric tons of selenium, almost 60 percent 

is located in Chile, China, and Russia combined, whereas 

Figure Q6.—Continued Maps showing locations of seleniferous sedimentary outcrops and deposits and plant samples 

with significant selenium content. This map, B, shows the spatial relation among locations of Paleocene and Eocene, 

seleniferous, marine sedimentary outcrops and plant samples with greater than or equal to 50 ppm selenium. Map B is 

from Trelease and Beath (1949) and, with map A, shows the widespread extent of seleniferous bedrock across the Western 

United States and its correlation with selenium in plant samples.
Q20 Critical Mineral Resources of the United States—Selenium

EXPLANATION

Selenium figure Q7C

Tertiary marine

Tertiary continental

Upper Cretaceous marine

0 250 500 MILES

0 250 500 KILOMETERS

Sedimentary deposits

C

125° 115° 105° 95°

45°

40°

35°

25°

State boundaries from TIGER/Line digital data, 1:100,000, 1992 

Coastlines and hydrology from U.S. Geological Survey

 1:2,000,000-scale digital data, 1970

Albers Equal-Area Conic projection

Standard parallels 29°30' N and 45°30' N, central meridian 96° W

WA

ID MT

WY

OR

CA NV

UT

AZ NM

TX

KS

SD

ND

NE

CO

OK

Figure Q6.—Continued Maps showing locations of seleniferous sedimentary outcrops and deposits and 

plant samples with significant selenium content. This map, C, shows locations of potentially seleniferous 

marine and continental sedimentary deposits in the Western United States. Map C is from Seiler and others 

(2003, p. 51, fig. 20) and shows additional areas with potentially seleniferous bedrock not identified by 

Trelease and Beath (1949).
Environmental Considerations Q21

the United States has an estimated 10,000 metric tons of world 

selenium reserves, or about 8 percent. As of 2012, the leading 

copper-producing mines in the United States were located in 

Arizona, Nevada, New Mexico, and Utah and accounted for 

all U.S. domestic selenium reserves (Brininstool, 2015).

Globally, the countries with the largest mine production 

of copper in 2012 were Australia, Canada, Chile, China, 

Congo (Kinshasa), Indonesia, Kazakhstan, Mexico, Peru, 

Poland, Russia, the United States, and Zambia (Brininstool, 

2015). Copper mines from these countries account for the 

global reserves of selenium and are the sources of concentrates 

and residues for the leading selenium producers.

Undiscovered Resources

Although selenium is produced currently from copper 

ores—mostly from porphyry copper deposits—it also occurs 

in many other types of mineral deposits and earth materials 

(tables Q2 through Q5). The highest concentrations of 

selenium are likely to occur in deposits that host selenide 

minerals, such as the selenide vein-type deposits described by 

Simon and others (1997). Although the selenium grade is high 

in selenide deposits, the overall tonnage is low. Deposits of 

native sulfur may also contain high concentrations of selenium 

(table Q3), yet the tonnage of these deposits compared with 

that of porphyry copper deposits is also low.

The close association of selenium with sulfur is the 

reason that copper ores are a resource for selenium. One 

approach to identifying new selenium resources would be to 

investigate this selenium-to-sulfur relationship in other rock 

types. Rocks enriched in organic carbon are likely to have high 

selenium concentrations, where selenium has substituted for 

sulfur in diagenetic sulfide minerals. These resources might be 

coal (where selenium recovery is technically feasible but not 

currently economic), marine phosphate deposits, or oil shales. 

Presser and others (2004) have mapped a global distribution 

of selenium-rich rocks found in sedimentary basins enriched 

in organic carbon (fig. Q7). Although the purpose of their 

mapping was to identify selenium source rocks that might 

create environmental concerns, their analysis might also be 

useful in identifying new selenium resources.

Environmental Considerations

Selenium is both a necessary nutrient and a potential 

hazard to Earth ecosystems, and it is present in the environment both naturally and as a result of human activity. Selenium mining, processing, use, and disposal may all potentially 

release selenium to the environment. Environmental 

considerations are integral to decisions regarding extraction 

and use of selenium.

Table Q6. Estimated world selenium reserves in 2014, in 

metric tons.

[Reserves are defined as the amount of selenium that could be economically extracted at the time the estimate was made (see U.S. Geological 

Survey, 2014a, appendix C, for more information). Selenium reserves, 

including those in China, are based upon the selenium content of 

identified copper deposits and do not include lead ores and coal. Other 

countries with reserves may include Australia, India, Kazakhstan, the 

United Kingdom, and Zimbabwe. Data have been rounded to no more 

than two significant digits. Data are from George (2014).]

Country Reserves

(metric tons)

United States 10,000

Canada 6,000

Chile 25,000

China 26,000

Peru 13,000

Poland 3,000

Russia 20,000

Other countries 21,000

World total 120,000

Sources and Fate in the Environment

The geochemical processes that control the presence 

and distribution of selenium in mineral deposits also operate, 

more broadly, in Earth’s surface environment. The processes 

of oxidation and reduction, adsorption at mineral surfaces, 

and isomorphic substitution of selenium for sulfur also 

govern selenium distribution and fate, including its uptake 

in biological organisms.

Natural and Anthropogenic Sources in Air, 

Water, and Soil

Volcanoes are the primary source of selenium in Earth’s 

upper crust and atmosphere. The selenium species H2Se 

and SeO2 are volatile and found along with sulfur gases in 

volcanic emissions. Lakin and Davidson (1973) estimated that 

1 kilogram per square meter (reported as 2 pounds per square 

yard) of selenium has been deposited on Earth’s surface from 

volcanic emissions over geologic time.

Selenium is produced in volcanic emissions when 

magmatic selenium is volatilized while in the volcanic 

plumbing system, and tens of kilograms of selenium can 

be released daily from a single volcano (Greenland and 

Aruscavage, 1986; Floor and others, 2011). Regardless of how 

selenium is emitted into the atmosphere, once there, it can be 
Q22 Critical Mineral Resources of the United States—Selenium

deposited onto surfaces by dry deposition or incorporated into 

water droplets and removed by wet deposition.

After selenium has been introduced to the upper crust, 

it cycles among the atmosphere, biosphere, and hydrosphere. 

Natural biologic emissions that cycle selenium between the 

biosphere and atmosphere account for approximately one-half 

to two-thirds of atmospheric selenium (Mosher and others, 

1987); these emissions are composed mostly of dimethyl 

selenide gas from marine biogenic sources and account for 

70 percent of natural biologic emissions of selenium. Other 

natural sources of atmospheric selenium are the continental 

biosphere (20 percent), volcanoes (8 percent), sea salt 

(2 percent), and crustal weathering (less than 1 percent) 

(Cahill and Eldred, 1998).

Anthropogenic selenium emissions to the atmosphere are 

derived mostly from coal and oil combustion (50 percent and 

10 percent, respectively), and base-metal production of copper, 

zinc, and lead (20 percent, 4 percent, and 4 percent, respectively) using roasters and autoclaves (Cahill and Eldred, 1998; 

Winkel and others, 2012). It has been known for some time that 

coal combustion is a major source of selenium to the environment and that coal can contain 5 to 300 times the amount of 

selenium as other rocks (table Q5). Emission controls on coalfired powerplants are limited in their ability to remove volatile 

selenium species (Ranville and others, 2010); hence, selenium 

escapes into the atmosphere. Coal combustion also generates 

fly ash that can concentrate selenium 10 to 25 times beyond its 

original concentration in the coal (Lemly, 1985). Other than 

coal, the remaining anthropogenic selenium emissions come 

from other forms of combustion and from ceramic and glass 

manufacturing (Cahill and Eldred, 1998).

One important natural source of selenium to surface and 

groundwaters is weathering of marine shales. A significant 

portion of the bedrock in the Western United States is composed 

of marine shales (most notably of Cretaceous age but with a 

range of Permian to Paleogene ages). These shales are noted for 

their high selenium content (table Q5) derived from the volcanic 

eruptions that were especially extensive during the Cretaceous 

Period. Selenium became incorporated into marine sediments, 

often substituting for sulfur in pyrite, and then mobilized 

as sediments were uplifted and weathered. Selenium-rich 

soils in the northern Great Plains of the United States owe 

their selenium content to the underlying shales of the Pierre 

Formation and the Niobrara Formation (Trelease and Beath, 

1949; Mayland, 1985). Moxon and others (1938) hypothesized 

that the selenium content of the Pierre Formation and the 

Niobrara Formation was inversely related to the rate of 

sediment deposition because they observed that thin stratigraphic sections (which were produced from low deposition 

rates) had high concentrations of selenium, whereas thicker 

Petroleum basin

Marine phosphate

 deposit

EXPLANATION

Selenium figure Q8. 

180° 90° W 90° E 180° 0°

15° N

15° S

30° N

30° S

45° N

45° S

60° N

60° S

75° N

75° S

0°

Figure Q7. A predictive map of selenium source rocks associated with organic-rich depositional marine basins. 

This map is useful both for identifying selenium source rocks that could be of potential environmental concern and 

for identifying potential new selenium resources. Modified from Presser and others (2004, p. 313).
Environmental Considerations Q23

sections (representing fast deposition rates) had low selenium 

concentrations. Other marine shales, such as the Meade Peak 

member of the Permian Phosphoria Formation in southeastern 

Idaho and western Wyoming (Stillings and Amacher, 2004; 

Presser and others, 2004), the Cretaceous-Paleocene Moreno 

Formation, and the Eocene-Oligocene Kreyenhagen Formation 

in the California Coast Range (Presser, 1994) are sources of 

selenium to groundwater, surface water, and soil. Additionally, coal and other organic-rich sediments release selenium 

through weathering (McNeal and Balistrieri, 1989).

In other regions of the United States where the soils are 

selenium-deficient, and in other countries, selenium is added 

to soils by way of selenium-containing fertilizers, lime, and 

manure. It is widely recognized that selenium is a necessary 

dietary nutrient for humans, as well as for other animals, 

and in the United Kingdom, selenium fertilizers are used in 

pastures and forage areas to increase the human intake of selenium through meat and dairy products. In Finland, a nationalscale use of selenium fertilizers began in 1984 and has resulted 

in a fourfold increase in the intake of dietary selenium in 

the human study population (Winkel and others, 2012, and 

references therein). Because crops take up only 5 to 20 percent 

of the selenium applied in fertilizers, the rest of the selenium 

is either immobilized in the soil through chemical reduction 

or leached from the soil and added to surface runoff waters. 

The Finnish study calculated that application of selenium 

fertilizers was expected to have increased selenium in soils by 

20 percent, but because no increase in the soil was observed, 

selenium must have been leached from the soils. In Tennessee, 

the use of selenium in fertilizers has caused selenium runoff in 

surface waters, resulting in an accumulation in aquatic biota 

(Winkel and others, 2012, and references therein).

Average Ambient Concentrations in Air, Streams, 

Groundwater, and Soil

Ambient concentrations of selenium in the environment 

are dependent upon selenium concentrations in the underlying 

bedrock. Soils in the United States that overlie pre-Cretaceousage sedimentary rocks, volcanic ash, or granitic rock tend to 

be low in selenium because most of these bedrock materials 

are also low in selenium (McNeal and Balistrieri, 1989).

Recent data on soil chemistry in the conterminous 

United States show selenium concentrations to range from less 

than 0.2 to 8.3 ppm (figs. Q8A–C; Smith and others, 2014). 

The data were collected from three soil sections—0 to 5 centimeters (cm) depth, the A horizon, and the C horizon. Although 

the spatial distribution of selenium varied among the three soil 

sections (figs. 8A–C), the concentration range and histogram 

of selenium concentrations were fairly consistent among 

them—from less than 0.2 to 6.9 ppm for the 0-to-5-cm-depth 

section; from less than 0.2 to 8.3 ppm for the A horizon 

section; and from less than 0.2 to 7.5 ppm for the C horizon 

section. Many of the general trends in selenium concentration 

in the Western United States, as mapped by Trelease and 

Beath (1949), were confirmed by current maps, as illustrated 

in figs. Q8A and Q8B. A striking feature in the recent maps, 

however, is the trend of high concentrations of selenium in 

the Dakotas, and eastward through the mid-continent and into 

the Appalachian and New England regions. Future scientific 

studies may debate the authigenic versus anthropogenic origin 

of this selenium.

Scandinavia is another region with low selenium content 

in the bedrock. In Sweden, 89 percent of the mineral soils 

have less than 0.45 ppm selenium, which is comparable to 

the blue-to-light-orange regions of the United States shown 

in figs. Q8A–C, and is the minimum concentration needed for 

forage to provide the necessary selenium for stock animals 

(Parkman and Hultberg, 2002, and references therein). In 

contrast, selenium-amended agricultural soils show increased 

selenium of 0.065 to 1.7 ppm. Lake waters in Sweden—both 

those affected by and those not affected by selenium effluent—

show between 43 and 209 nanograms per liter (ng/L) of 

total selenium (total = particulate + dissolved), and this 

concentration is independent of the trophic status of the lake 

(Parkman and Hultberg, 2002, and references therein).

In Finland, background concentrations of dissolved 

selenium in lake waters are estimated to be 56 to 62 ng/L, and 

although no seasonal variations were observed, researchers 

did find spatial variations caused by human activities. As 

a comparison, in lakes affected by agricultural runoff, the 

selenium concentration was 76.8 ng/L. In these lakes, the total 

selenium concentration was distributed among (1) selenium in 

organic materials, such as humic acids, which was greater than 

or equal to 50 percent of the total; (2) particulate selenium that 

was approximately 10 percent of the total selenium concentration; (3) dissolved selenite and selenate, which accounted 

for 8 to 9 percent of the total selenium concentration; and 

(4) selenium associated with neutral humic acids, which 

made up approximately 5 percent (Parkman and Hultberg, 

2002). Sediments in these lakes show an increase in selenium 

concentration during the 20th century, most likely because 

of selenium-containing fertilizers. In 1992, concentrations in 

nine lakes ranged between 0.27 and 3.64 ppm, and concentrations were higher in oligotrophic lakes than in mesotropic 

or eutrophic lakes. In stream sediments in Finland, the mean 

concentration of selenium was 0.26 ppm (Parkman and 

Hultberg, 2002, and references therein).

Above-Background Concentrations of Selenium 

in the Environment Near Major Deposits and 

Enriched Bedrock

Soils can have highly variable selenium concentrations, 

ranging from less than 0.01 ppm in deficient soils to up to 

1,200 ppm in selenium-rich soils (Winkel and others, 2012, 

and references therein). Seleniferous soils can be found overlying selenium-rich rocks, such as black shales, carbonaceous 

limestones, carbonaceous cherts, mudstones, and seleniferous 

coal (Winkel and others, 2012).
Q24 Critical Mineral Resources of the United States—Selenium

A

Base from U.S. Geological Survey digital files, Lambert Conformal Conic projection

Standard parallels 33° N. and 45° N., central meridian 96° W.

Datum is North American Datum of 1983

Selenium figure Q9A

4.9

EXPLANATION

0

50

60

70

80

90

100

0

0.2

0.3

0.4

0.5

0.6

3.3

Outlier

0 250 500 KILOMETERS

0 125 250 MILES

125°

45°

115° 105° 95° 85° 75° 65°

35°

25°

Percentile

Selenium

ppm

Figure Q8 (pages 24 and 25). Maps showing selenium concentrations in soils of the conterminous United States. The 

selenium concentrations shown are in A, the top 0 to 5 centimeters of surface soils, B, the A horizon of surface soils, and 

C, the C horizon of surface soils. A visual comparison of the soils maps with the bedrock and plant maps in fig. Q6 shows 

many correlations in the Western United States. Maps are from Smith and others (2014, p. 197–202, figs. 95, 96, and 97). 

ppm, parts per million

Some examples of seleniferous soils are those that 

develop over seleniferous marine shales in the United States. 

Presser (1994) found that salinized soils of alluvial fans in 

the San Joaquin Valley, California, could contain as much 

as 4.5 ppm of selenium, with a median of 1.8 ppm. These 

soils developed from weathering of the Upper CretaceousPaleocene Moreno Formation and the Eocene-Oligocene 

Kreyenhagen Formation, which themselves contained concentrations of selenium up to 45 ppm, with median concentrations 

of 6.5 and 8.7 ppm, respectively (table Q5). Even higher 

concentrations have been recorded in shales of the Cretaceous 

Pierre Formation and the Niobrara Formation in the Western 

United States (maximums of 103 and 113 ppm, respectively) 

(table Q5; Moxon and others, 1938; Trelease and Beath, 1949; 

Presser, 1999).

High concentrations of selenium can develop in irrigation drainage waters and wetland environments when these 

waters interact with seleniferous soils. The wetlands of the 

Kesterson National Wildlife Refuge in San Joaquin Valley, 

Calif., were the site of an ecological crisis beginning in 1983, 

when it was discovered that high concentrations of dissolved 

selenium caused massive numbers of embryonic deformities 

and deaths of aquatic wildlife (Garone, 1999; Ohlendorf, 

2002). The wetlands developed at the terminus of irrigation 

drainage within the San Joaquin Valley, and the drainage 

waters held high concentrations of selenium (in the range of 

106 to 452 parts per billion (ppb), with an average concentration of 300 ppb) leached from seleniferous, salinized soils 

that developed over high-selenium-content marine shales. 

The link between high selenium concentrations and selenium 

toxicity is discussed in the sections of this chapter on human 

and ecological health.

Soils in the vicinity of sulfide orebodies may show 

naturally high concentrations of selenium, and this association has been cited in models for exploration geochemistry 

(Koljonen, 1976). The Luikonlahti copper orebody in eastern 

Finland displays elevated selenium concentrations in soils 

and glacial deposits immediately adjacent to sulfide-rich 

rocks (0.1 to 0.4 ppm selenium), in colloid- and organic-rich 

river and lake sediments downstream from the orebody 
Environmental Considerations Q25

B

Base from U.S. Geological Survey digital files, Lambert Conformal Conic projection

Standard parallels 33° N and 45° N, central meridian 96° W

Datum is North American Datum of 1983

0 250 500 KILOMETERS

0 125 250 MILES

Selenium figure Q9B

EXPLANATION

0

50

60

70

80

90

100

0

0.2

0.3

0.4

0.5

0.6

4.7

Outlier

6.7

8.3

6.7

125°

45°

115° 105° 95° 85° 75° 65°

35°

25°

Percentile

Selenium

ppm

C

Base from U.S. Geological Survey digital files, Lambert Conformal Conic projection

Standard parallels 33° N and 45° N, central meridian 96° W

Datum is North American Datum of 1983

0 250 500 KILOMETERS

0 125 250 MILES

Selenium figure Q9C

Percentile

Selenium

EXPLANATION

0

60

70

80

90

100

0

0.2

0.3

0.4

0.6

7.5

ppm

125°

45°

115° 105° 95° 85° 75° 65°

35°

25°
Q26 Critical Mineral Resources of the United States—Selenium

(0.2 to 1.1 ppm selenium), in forest mull humus atop the 

orebody (2.3 to 5.5 ppm selenium), and in sulfate- and 

iron-oxide-rich gossans developed above the orebody 

(2.9 to 13.0 ppm selenium). In contrast, similar soils and 

sediments from nonmineralized regions in Finland showed 

background selenium concentrations of 0.0 to 0.2 ppm 

selenium, with iron-rich lake and bog ores containing 

0.4 ppm selenium (Koljonen, 1976).

Groundwaters near an undisturbed porphyry copper 

deposit in the Atacama Desert of Chile also display aboveaverage concentrations of selenium. Leybourne and Cameron 

(2008) measured naturally occurring concentrations of up to 

800 ppb selenium within 2 kilometers (km) downstream of 

the deposit. They attribute these high concentrations to Eh-pH 

conditions and pH, which put the waters at the boundary 

between Fe(III) and Fe(II) speciation. Limited quantities of 

solid iron oxyhydroxides are present in environments with 

these values of Eh-pH; therefore, adsorption reactions with 

iron oxyhydroxides did not reduce dissolved concentrations 

of selenium.

In contrast, sorption reactions may be an important 

control on selenium mobility in volcanic soils (Floor and 

Román-Ross, 2012). High concentrations of selenium have 

been recorded in volcanic soils owing to weathering of 

seleniferous parent materials and deposition of atmospheric 

selenium. Floor and Román-Ross (2012) reported 1,200 ppm 

of selenium in volcanic soils from Ireland, and 1 to 20 ppm 

of selenium in soils from Hawaii. Floor and others (2011) 

reported a range of 0.12 to 2.25 ppm selenium in soils from 

the Mount Etna volcano in Italy, with a median of 0.39 ppm 

selenium. Despite these high soil concentrations, selenium is 

not readily leached from soils. Strong sorption of selenium 

with soil compounds, such as ferrihydrite, poorly crystalline aluminosilicates, and organic matter, has been cited in 

volcanic soils from Hawaii, Mount Etna in Italy, and New 

Zealand (Floor and Román-Ross, 2012).

The strong sorption between selenium and volcanic soils 

may be site specific, as sorption is highly dependent upon soil 

chemistry and mineralogy. Whereas groundwater in volcanic 

aquifers can have low concentrations—often less than 0.1 ppb 

selenium (Floor and Román-Ross, 2012)—some sites show 

distinctly higher concentrations. Aiuppa and others (2000) 

sampled groundwaters near Mount Etna and found a range of 

0.6 to 66.8 ppb selenium, with an average of 2.9 ppb selenium. 

At Mount Vesuvius, Italy, Aiuppa and others (2005) demonstrated that selenium can be one of the most mobile elements 

during rock weathering.

Plants show elevated concentrations of selenium when 

grown in soils with high selenium concentration, although the 

accumulation can vary by more than two orders of magnitude 

for a given selenium concentration in the soil (Winkel and 

others 2012, and references therein). Selenium accumulator 

plants include Astragalus L., as well as Oonopsis (Nutt.) 

Greene, Oryzopsis Michx., Xylorhiza Nutt., Mentzelia L., 

and Stanleya Nutt. (Moxon and others, 1938; Trelease and 

Beath, 1949; Cannon, 1957; Presser, 1999). Cannon (1957) 

reported that members of Astragalus can display selenium 

concentrations of several hundreds to a thousand parts per 

million when grown in soils that contain only 2 ppm selenium. 

In addition to soil selenium, other factors that affect the plant’s 

selenium content include the age of the plant, the stage of 

growth, and the part of the plant analyzed—seed, leaf, or stem 

(Moxon and others, 1938).

Mobility in Water and Soils

The chemical form of selenium (its speciation) determines its behavior and fate in the environment. Selenium is 

a metalloid that commonly takes an anionic form in oxic to 

suboxic conditions when dissolved, and its primary species 

in water are selenate (Se(VI) or SeO4

2–), selenite (Se(IV) or 

SeO3

2–), and organo-selenide (org-Se(-II), such as selenomethionine) (Luoma and Presser, 2009). Speciation will also 

determine whether selenium can be taken up by algae or 

microbes, or whether it can transform to particulate matter in 

seawater (seston) or in sediments.

In the sections above, the sources of selenium in the environment have been identified as emissions from volcanoes, 

biogenic processes, manufacturing and power generation, 

weathering of marine shale bedrock and overlying soils, and 

selenium leachate from fly ash and other solid waste materials. 

Mobilization of selenium from these sources to aquatic 

environments and pore waters occurs under oxidizing, slightly 

alkaline weathering conditions and produces soluble selenate 

and selenite (McNeal and Balistrieri, 1989; Presser and Swain, 

1990; Piper and others, 2000).

Once selenium has transformed to aqueous selenate 

and selenite, it is readily mobile and bioavailable (meaning 

that it can be taken up by organisms). Selenate predominates 

in well-aerated surface waters, especially at alkaline pH. 

Selenite predominates in slow-moving waters, such as lakes 

and reservoirs. Chemical parameters, such as pH and redox, 

and microbial respiration, can convert selenate to selenite 

and organo-selenium. Dissolved selenium can sometimes 

be predominately organo-selenium, and this species may be 

seasonally important (Skorupa and others, 2004). The presence 

of organic carbon, ferric oxyhydroxides, and clays in soils 

affect the partitioning of these selenium species among water 

and soil and (or) sediments.

Selenate (Se (VI)) does not form strong adsorption 

complexes on mineral surfaces. It can be easily desorbed 

by a change in solution chemistry (for example pH, or ionic 

strength), and it exchanges readily with SO4

. Ready desorption 

and exchange are among the reasons for the high solubility of 

Se(VI), which enhances its mobility (Fukushi and Sverjensky, 

2007). In solutions of alkaline pH with high Eh values, Se(VI) 

is mobile and bioavailable to plants (Winkel and others, 2012). 

In contrast, selenite (Se(IV)) is stable in solutions of moderate 
Environmental Considerations Q27

to low Eh values, and it is more likely to adsorb irreversibly 

to iron oxyhydroxides, clay minerals, and organic matter than 

to selenate. When selenite is tightly held by mineral surfaces 

and organic matter, it is less mobile and less available for plant 

uptake (Winkel and others, 2012, and references therein).

In a system where mineral surfaces and organic matter 

are not present in sufficient concentrations, however, Se(IV) 

and organo-selenide may be more readily incorporated into 

biota. When selenium is taken up at the base of the food web 

by plants and microbes, it is converted to organo-selenide, 

regardless of whether it was initially present as Se(IV) 

or Se(VI) (Presser and others, 1994). Organo-selenide is 

rereleased to water as these cells die or are consumed, and it 

can be reoxidized to Se(IV). Slow reaction kinetics restrict 

reoxidation to Se(VI). This means that biologically productive 

systems will evolve to see a decrease in soluble Se(VI) and an 

increase in Se(IV) and organo-Se(-II). For example, in streams 

in the upper part of California’s San Joaquin River watershed, 

dissolved selenium is almost 100 percent Se(VI), but lower in 

the watershed and downstream of the delta of the San Joaquin 

River, Se(VI), Se(IV), and organo-Se(-II) are equal in abundance (Cutter and Cutter, 2004).

Elemental (native) selenium is not soluble in water and 

is considered unavailable to plants. Some estimates suggest 

that elemental selenium may account for 30 to 60 percent of 

the total selenium in sediments (Zhang and others, 2004). It 

is often found as nanometer-sized spherical particles that are 

produced from microbially mediated reduction of Se(VI) and 

Se (IV). The extremely small size of these particles enables 

them to remain in colloidal suspension for weeks, and they 

may be able to enter cells directly, making the particles more 

bioavailable. A high surface-to-volume ratio makes them 

highly reactive, meaning that they may reoxidize to selenite 

and selenate. Finally, an organic polymer layer of protein that 

has been observed on the surface of these spherical particles 

may alter their surface properties and affect the distribution 

of elemental selenium in the environment (Winkel and 

others, 2012).

It was noted earlier that selenium exchanges readily with 

sulfur; this is because selenium and sulfur are both members 

of Group 6b of the periodic table; hence, they exhibit similar 

chemical behavior because they have similar ionic radii and 

can have the same valence. The S(VI) competition with Se(VI) 

for adsorption sites on mineral surfaces and in biological 

uptake is a primary mechanism for Se(VI) mobilization. For 

example, in soils developed on the flanks of Mount Etna, it 

was observed that selenium is mobilized during rainwatersoil interaction. It was found that aluminum oxides and soil 

organic matter were the active phases to sequester selenate 

in soil through adsorption and complexation reactions. 

When sulfate was introduced from rainwater, however, 

sulfate exchanged with adsorbed and complexed selenate, 

transforming it to a soluble and therefore mobile selenium 

species. Yet when redox conditions allowed for selenium 

transformation to Se0, selenium would not be remobilized by 

acidic water (Floor and others, 2011).

Mine Waste Characteristics

Coal deposits and associated mine waste rock are 

common geologic sources for environmental selenium 

(fig. Q5). In the Appalachian region, coal mining companies 

often practice mountaintop mining to access underlying coal 

seams, and the waste rock is commonly placed in streams 

and valleys. According to the U.S. Environmental Protection 

Agency (EPA), an estimated 724 stream miles (1,165 stream 

kilometers), or 1.2 percent of streams, were covered by more 

than 6,800 valley fill waste piles from 1985 to 2001 in the 

States of Kentucky, Tennessee, Virginia, and West Virginia 

(U.S. Environmental Protection Agency, 2005, p. 4–5). This 

waste disposal practice serves to increase selenium input 

into aquatic systems because fresh faces on newly broken 

rock provide increased surface area for water-rock reaction 

and selenium transfer into surface waters. Bryant and others 

(2003) observed that the selenium concentration in Appalachian streams associated with valley fills (<0.3 to 37 ppb 

selenium) was greater than background concentration 

(<0.3 ppb selenium), and that the selenium concentrations 

in holding ponds at the base of valley fills ranged from 

24 to 42 ppb selenium. Surface and groundwater pH greatly 

affect selenium solubility. Unlike other trace elements associated with coal (for example, cadmium, copper, and lead), the 

oxic forms of selenium are soluble in waters with neutralto-alkaline pH. This behavior is explained by the difference 

in adsorption potential among base cations (for example, 

cadmium, copper, and lead), which adsorb strongly to oxide 

surfaces at higher pH values, and oxyanion forms of selenate 

and selenite, which adsorb more strongly at lower pH values 

(Stumm, 1992; Elzinga and others, 2009).

Although selenium can occur naturally in soils associated 

with unmined coal seams, mining is known to enhance its 

solubility in reclaimed and revegetated mine lands, which can 

potentially result in aquifer contamination and elevation of 

selenium in plant materials. Sharmasarkar and Vance (1997) 

analyzed 40 soil samples (ranging in depth from 0 to 200 cm) 

from native, reclaimed, and abandoned coal mine environments in Wyoming, and observed selenium concentrations 

of 0.26 to 26.13 ppm selenium, with a mean concentration 

of 1.73 ppm.

Another source of selenium to natural waters is ash and 

other waste products produced from coal-fired powerplants. 

The case study for this selenium pathway was Belews Lake in 

North Carolina. From 1974 to 1985, lake water was withdrawn 

and mixed with ash generated by a nearby coal-fired electricitygenerating plant. Presser (1999) noted that fly ash may contain 

even higher levels of soluble selenium than the initial, seleniferous coal; that is, 50 to 500 ppm versus 2 to 20 ppm. Solids 

were allowed to settle from the slurry before the wastewater, 
Q28 Critical Mineral Resources of the United States—Selenium

which contained 150 to 200 ppb selenium, was returned to the 

lake, creating lake water selenium concentrations of 20 ppb. 

Once this disposal practice was ended, lake water selenium 

concentrations fell to less than 1 ppb (Lemly, 1997).

High ash content in bone coals (coal with a high ash 

content) creates another source of environmental selenium. 

In Pingli County, Shaanxi Province, China, the Badao 

bone coal contains 41 to 75 ppm of selenium, with a mean 

concentration of 57.5 ppm. After combustion, the coal ash 

fraction was found to contain 28 ppm of selenium, indicating 

that the difference (approximately 30 ppm) is volatilized to 

the atmosphere. Selenium is also released to the environment 

through the use of bone coal ash as an agricultural soil amendment. The local environment in Pingli County was found to 

contain soil with a mean concentration of 7.1 ppm selenium, 

surface water with concentrations of 0.3 to 1.1 ppm selenium, 

and agricultural crops with concentrations of 0.57 to 2.75 ppm 

selenium (Fang and others, 2003).

As discussed previously, selenium is associated with 

phosphate deposits, and, quite notably, in the United States, 

with the Meade Peak member of the Permian Phosphoria 

Formation in Idaho, Montana, Nevada, Utah, and Wyoming. 

Features of the mined environment include lakes that form in 

open pit mines after mining has ended, waste rock contoured 

into hills or emplaced as cross-valley fills, wetlands that form 

at the base of waste dumps, and vegetation that is reseeded 

onto the surface of the dumps.

A selenium mass-balance study was conducted for the 

wetland that formed at the base of the Wooley Valley Unit 4 

waste dump in southeastern Idaho (Stillings and Amacher, 

2004; Stillings and others, 2007). The waste dump was created 

in the 1970s from overburden and waste shale that remained 

after the mining of the upper and lower phosphate zones of the 

Meade Peak member, and the wetland developed soon thereafter. In 2002, inlet and outlet flow volumes and chemistry 

were measured continuously from May through September. 

The calculated influx of selenium was high at the beginning of 

May (approximately 90 grams per day [g/d] of selenium), but 

dropped rapidly throughout the month and became zero when 

the seep dried up at the end of June. The outflux of selenium 

ranged from approximately 0.05 to 2.0 g/d during the entire 

season. The mass balance created for this time period showed 

that although 685 grams (g) of selenium was delivered to 

the wetland, only 84 g of selenium left the wetland, meaning 

that 601 g, or 88 percent of the influx, was retained within 

the wetland. Further work determined that selenium was 

scavenged from wetland surface waters by adsorption and 

co-precipitation with iron oxyhydroxides that precipitated on 

the surface of wetland soils (Stillings and Amacher, 2004; 

Stillings and others, 2007).

In a related study, selenium concentrations were 

measured in plants grown on the waste rock dumps at Wooley 

Valley Units 1, 3, and 4 and compared with concentrations of 

plants growing on undisturbed areas of the Phosphoria Formation. Vegetation growing in highly disturbed soils, including 

those on the surface of waste dumps, had the highest tissue 

selenium concentrations. The mean concentration for legumes 

was 80 ppm; trees, 52 ppm; grasses, 18 ppm; shrubs, 6 ppm; 

and forbs, 3 ppm. The grasses collected from the wetland 

described in the previous paragraph showed a mean selenium 

concentration of 53 ppm. This second study was used to 

argue for the removal of legumes from seed mixtures used 

to revegetate the surfaces of the cross-valley waste dumps, 

thus decreasing selenium in browse vegetation for herbivores 

(Mackowiak and others, 2004).

Roll-front uranium deposits in the Western United States 

are another deposit type known to contain selenium; hence, 

selenium is likely to be present in these mine wastes as well. 

When uranium is mined with in situ methods (which are 

commonly used with roll-front deposits), the overlying rock is 

not disturbed and thus not exposed to oxidizing conditions at 

Earth’s surface; however, the in situ mining methods used rely 

on injections of lixiviant (an oxidizing agent—either oxygen 

or hydrogen peroxide) into the uranium deposit to oxidize 

and extract the uranium ore. This fluid can leach selenium 

into the surrounding groundwater by oxidizing organic matter 

that contains selenium, and by converting insoluble organic 

selenides to soluble selenate and selenite. Selenium concentrations can exceed 1,000 ppm in areas that have been mined 

for uranium (Naftz and Rice, 1989, and references therein).

Both groundwater and surface water in areas near copper 

mines and processing facilities may show elevated selenium 

concentrations because of selenium’s association with sulfide 

minerals. Elevated selenium concentrations (>10 ppm) have 

been found in groundwater under the copper smelter and 

refinery facility near the Great Salt Lake, Utah. The facility 

is located within 1.5 km of the southern shore of the lake, 

and there is some concern that a portion of this plume could 

discharge into the lake (Naftz and others, 2009, p. 35, and 

references therein). In 2001, the facility was selected for 

an EPA selenium treatment/demonstration project, which 

field-tested three technologies for removing selenium from 

the Garfield Wetlands-Kessler Springs site. At the start of this 

project (in May 1999), the Garfield Wetlands-Kessler Springs 

water contained approximately 2,000 ppm selenium (MSE 

Technology Applications, Inc., 2001, p. 1).

Soils near a lead-zinc smelter in Kabwe, Zambia, show 

contamination from atmospheric dust fallout from smelting 

activities. Kříbek and others (2010) sampled contaminated 

topsoil and a reference (background) subsoil horizon at 80 to 

90 cm depth. The topsoil had a maximum selenium concentration of 0.01 percent (100 ppm), which was significantly higher 

than the concentrations of trace metals in the subsurface soil.

There are limited reports of selenium in leachate from 

mine waste associated with banded iron formations in the 

Upper Peninsula of Michigan. In 2008, the Michigan Department of Environmental Quality (MDEQ) began to assess selenium concentrations and water quality impacts in the vicinity 

of the Empire Mine and the Tilden Mine in the Marquette 

Iron Range near Palmer, Michigan. Results of the MDEQ 
Environmental Considerations Q29

assessment showed that selenium concentration in water at 

many lakes near the mine sites were less than 5 ppb, which is 

the chronic selenium water quality standard for Michigan. Two 

streams that flowed through mine waste facilities carried mean 

values of 9.3 and 11 ppb selenium, however, and a popular 

fishing site, Goose Lake, exhibited 8.8 to 10 ppb selenium. 

Small streams and springs buried by waste rock piles showed 

selenium concentrations that ranged from 15 to 68 ppb. 

Sediments from streams and lakes ranged from less than 

5 ppm to between 27 and 39 ppm selenium in Goose Lake 

(Selenium Monitoring Work Group, 2009, p. ii). Biological 

samples from Goose Lake Outlet showed a mean value of 8.0 

to 17.35 ppm selenium in brook trout (Salvelinus fontinalis

[Mitchill, 1814]), white suckers (Catostomus commersonii

[Lacepède, 1803]), fathead minnows (Pimephales promelas

[Rafinesque, 1820]), and creek chubs (Semotilus [Rafinesque, 

1820]) and 22.3 ppm selenium (composite sample of northern 

caddisfly [Limnephilidae (Kolenati, 1848)] tissue). Selenium 

was present in lower levels in benthic macroinvertebrates, 

such as leeches, snails, mayflies, and odonates (Selenium 

Monitoring Work Group, 2009, p. 7–8).

Human Health Concerns

Dietary selenium plays a critical role in human health 

owing to its importance to thyroid hormone metabolism, antioxidant defense mechanisms, and immune functions (Rayman, 

2000). Insufficient selenium is associated with growth retardation, impaired bone metabolism, and abnormalities in thyroid 

function (Stone, 2009). Human populations with insufficient 

selenium in their diet can be found living in geographic areas 

with low selenium contents in local bedrock and soils. One 

example is a northeast-southwest-trending belt in China and an 

area of southeastern Siberia where populations are at risk for 

Keshan disease (which causes cardiomyopathy) and Kashin 

Beck disease (a chronic degenerative osteoarthritis); both are 

caused by a deficiency of selenium.

Keshan disease was described in Chinese medical 

literature 100 years ago, but it was not until the mid-1970s 

that selenium deficiency was identified as its primary cause. 

After a widespread outbreak in 1935, it was determined to 

be endemic in children aged 2 to 10 years and in women of 

child-bearing age. Symptoms are fatigue after light exercise, 

cardiac arrhythmia and palpitations, loss of appetite, cardiac 

insufficiency, cardiomegaly, and congestive heart failure 

(World Health Organization and Food and Agricultural Organization of the United Nations, 2004). Kashin Beck disease 

was also detected in Chinese children aged 5 to 13 years and 

in southeastern Siberia. The disease causes joint necrosis in 

the form of degeneration of the arm and leg joints, which 

results in structural shortening of bones and leads to growth 

retardation and stunting (World Health Organization and 

Food and Agricultural Organization of the United Nations, 

2004).

In Finland, blood serum levels of less than 45 ppb 

selenium were believed to increase the risk of cardiovascular 

disease and cancer. Because of this, sodium selenate has been 

added to agricultural fertilizers since 1985 in order to increase 

the dietary intake. This practice has caused selenium dietary 

intake to increase from a base of 30 to 40 micrograms per day 

(μg/d) to greater than 100 μg/d by the early 1990s, and the 

daily intake is now considered safe and adequate (Varo and 

others, 1994). It should be noted, however, that some research 

recommends lower levels of dietary intake, and finds that 

doses of selenium greater than 75 μg/d, can be mildly toxic 

and related to an increased risk of diabetes (Winkel and others, 

2012, and references therein)

Somewhat analogously, human populations living in 

geographic areas with high levels of selenium in local foods 

may suffer from selenium toxicity. This has been observed 

in humans and animals in the Hubei and Shaanxi Provinces 

in China, and in Punjab, India, where selenium levels in 

local foods are high and dietary intake per person is between 

750 and 4,990 μg/d. In Punjab, the high dietary selenium 

values are the result of agricultural irrigation practices and 

high evaporation rates, which serve to concentrate selenium 

into irrigated soils and food crops. At doses of greater than 

400 μg/d, selenium is acutely toxic and may lead to acute 

respiratory distress, cardiac failure, facial flushing, hair 

loss, kidney failure, light headedness, muscle tenderness, 

myocardial infarction, severe gastrointestinal and neurological 

symptoms, tremors, and, in rare cases, death (National 

Institutes of Health, 2013, and references therein). Chronic 

exposure to subacute levels leads to brittle hair and nails, skin 

lesions, and neurological disturbances (Winkel and others, 

2012, and references therein).

One outcome of modern research into selenium and 

human health is a better understanding of its role in the 

immune system. It has been found that the cardiac lesions of 

Keshan disease are not the result solely of selenium deficiency, 

but are also related to infection by the coxsackie virus, 

which is initially nonvirulent, but which becomes virulent 

and myopathogenic in individuals with selenium deficiency. 

This process is not unique to the coxsackie virus. It has been 

determined that the early preclinical stages of human immunodeficiency virus (HIV) infection are marked by a significant 

decrease in blood plasma selenium concentrations, and that the 

magnitude of this decrease is linked to the rates of development of acquired immunodeficiency syndrome (AIDS) and to 

mortality rates associated with AIDS (World Health Organization and Food and Agricultural Organization of the United 

Nations, 2004). Finally, selenium deficiency has been linked 

to an increase in mercury toxicity (Berry and Ralston, 2008). 

Because of the high binding affinity between mercury and 

selenium, it is thought that selenium can function to chelate 

mercury and reduce its toxicity; if the molar ratio of mercury 

to selenium is high, however (that is, greater than 1:1), then 

the concentration of selenium is insufficient to provide this 

protective effect.
Q30 Critical Mineral Resources of the United States—Selenium

Pathways for Selenium Ingestion in Humans

Plant foods are the main dietary sources of selenium 

in most countries, followed by meats and seafood (National 

Institutes of Health, 2013). In all cases, the selenium content 

of food is derived from soils and the feed for livestock (World 

Health Organization, 2011). Although soil selenium is the 

source of selenium, it is important to note that total selenium 

concentration is not what controls selenium levels in food 

crops; rather, it is the concentration of bioavailable selenium 

that controls the levels. For example, the classical definition 

of selenium-deficient soils is when concentrations are less than 

0.1 ppm selenium, whereas selenium-rich soils have greater 

than 0.5 ppm selenium. These concentrations refer to total 

selenium. Soils in areas of China where Keshan disease has 

occurred have an average concentration of 0.15 ppm selenium, 

which would not be classified as selenium deficient; however, 

these soils are selenium deficient because the bioavailable 

selenium is low. High content of organic matter is the cause 

of the low bioavailable selenium concentration because it 

immobilizes selenium by adsorption and biochemical reduction, making it unavailable for uptake by food crops.

The dietary intake of selenium varies somewhat across 

the United States and Canada because of differences in local 

soils and food; however, most Americans consume adequate 

amounts of selenium. The average daily selenium intake in 

Americans over 2 years in age is 108.5 micrograms from food 

alone, and if supplements are included, this value increases to 

120.8 μg/d (National Institutes of Health, 2013).

Related Diseases, Cancers, and Human Health 

Effects From Industrial Exposures

Little has been reported regarding selenium health effects 

from human industrial exposures, and it is selenium deficiency 

from an inadequate diet that is the greatest human health 

concern. Selenium concentrations in gas emissions from 

coal combustion could be one potential cause of industrial 

exposure, as could exposure in industries with a significant 

occupational metals exposure risk, such as metals recovery 

(Fairweather-Tait and others, 2011). The manufacturing 

process for modern products, such as solar panels, might 

represent a potential risk if selenium is present in a bioavailable form. Manufactured nanomaterials may also pose health 

effects, because selenium is a key component in CdSe and 

PbSe quantum dots, which are nano-sized semiconductors 

used in light-emitting diodes and photovoltaics. Emergent 

properties of these materials could elicit toxic responses that 

may be different from those current selenium sources (Young 

and others, 2010).

Human Health Guidelines

Although water is not usually a major source of selenium 

to the human population (Fairweather-Tait and others, 2011), 

the World Health Organization (WHO) has set a provisional 

guideline value for selenium in drinking water. The WHO 

recommendation is that when selenium intake from dietary 

sources is unknown, then 40 ppb should be the upper limit 

of selenium in drinking water (World Health Organization, 

2011). In Canada, the Government has set this level at 10 ppb, 

although the Canadian Government is revisiting the criterion 

and proposing a new maximum acceptable concentration 

for human consumption of 50 ppb in response to a review 

of current scientific studies and an understanding of chronic 

selenosis (Health Canada, 2013). In the United States, the 

EPA has set the maximum contaminant level for selenium 

in drinking water at 50 ppb (U.S. Environmental Protection 

Agency 2014a).

Because there is a narrow concentration range between 

dietary deficiency (< 40 μg/d) and highly toxic levels 

(>400 μg/d) (Winkel and others, 2012), the WHO recommends dietary intakes of 15 μg/d for infants from birth to 

6 months, 30 μg/d for children aged 4 to 8 years, and 55 μg/d 

for older children and adults (World Health Organization, 

2011). Older recommendations were 0.05 to 2 milligrams per 

day (mg/d) for adults, with lower ranges for children (Health 

Canada, 1992). In the United States, the National Institutes of 

Health has published recommended daily allowances (RDAs) 

for humans; these RDAs are similar to those recommended by 

the WHO, and range from an adequate intake of 15 μg/d for 

infants from birth to 6 months, to 55 μg/d for older children and 

adults (table Q7). Pregnant and breast-feeding women have 

RDAs of 60 and 79 μg/d, respectively (table Q7; National 

Institutes of Health, 2013).

Ecological Health Concerns

Research on selenium in the environment began in 

the 1930s when livestock in the Great Plains and Western 

United States were found with the symptoms of “alkali 

disease,” including emaciation, loss of hair, deformation and 

shedding of hooves, loss of vitality, and erosion of joints in 

long bones (Science Dictionary, 2015). It became recognized 

that livestock poisoning was occurring because the livestock 

were browsing on selenium accumulator plants of the genus 

Astragalus; these plants grow in saline soils derived from 

marine Cretaceous sedimentary rocks exposed in more than 

830,000 square kilometers of the Western United States 

(Trelease and Beath, 1949).

Once the cause of selenium poisoning was recognized, 

efforts were made to prevent and control the exposure of 

livestock to seleniferous plants using range management 

techniques. In 1957, however, it was learned that selenium is 

an essential trace element (or micronutrient) in mammals. It is 

an essential component of bacterial and mammalian enzyme 

systems, and that without it, selenium deficiency disorders, 

such as white muscle disease (in sheep) and mulberry heart 

disease (in pigs), could occur (Young and others, 2010, and 

references therein).
Environmental Considerations Q31

In 1976, the sudden disappearance of popular game fish 

from Belews Lake, North Carolina, was attributed to selenium 

poisoning. It was learned that the manmade reservoir was fed 

by water from a coal-ash settling basin that had created lake 

waters with selenium concentrations of 20 ppb. This level of 

selenium caused chronic selenium poisoning in the Belews 

Lake fish populations, where the most serious symptom was 

reproductive failure and post-hatch mortality. As a result of 

the selenium toxicity, 19 of 20 species of fish in Belews Lake 

were eradicated (Lemly, 2002).

Soon after this, in 1983, another extinction of fish populations was discovered. This time it was at the Kesterson Reservoir in San Joaquin Valley, Calif., which is a wetland area fed 

by agricultural drainage from the San Joaquin Valley. Massive 

numbers of dead and deformed bird embryos and chicks were 

found to accompany the fish extinction, and the cause was 

determined to be selenium poisoning from bioaccumulation 

in the food chain (Ohlendorf, 2002). After this discovery, 

deformities of aquatic birds from selenium toxicosis began to 

be observed in similar environments; that is, wetland basins that 

received waters draining seleniferous soils, such as the Tulare 

Lake Bed Area, Calif.; the Middle Green River Basin, Utah; 

the Kendrick Reclamation Project Area, Wyoming; the Sun 

River Basin, Montana; and the Stillwater Wildlife Management 

Area, Nevada. Most of these study areas had climate, geology, 

and hydrology similar to that of the San Joaquin Valley. Results 

indicated that bioaccumulation of selenium was occurring in 

most of the wetland areas, and these data marked the initial 

elucidation of a biogeochemical pathway for selenium from 

geologic source to aquatic wildlife (Presser and others, 1994).

Pathways for Selenium Ingestion in Aquatic and 

Terrestrial Organisms

Just as in human health, the main pathway for selenium 

entry into aquatic and terrestrial organisms is through diet, and 

speciation is an important factor in determining the amount 

of selenium that enters the food chain. The primary step by 

which selenium accumulates in aquatic systems is found at the 

base of the food chain where algae and other microorganisms 

(trophic level 1) accumulate selenium from water by factors 

of up to 1 million times the aqueous concentration. Selenate 

(Se(VI)) is slow to be taken up by these organisms, especially 

in the presence of sulfate, which is more competitive. Selenite 

(Se(IV)) and organo-selenides are taken up more quickly. 

Once selenium has been accumulated by an organism, the 

selenium is reduced to selenide (Se(-II)) and ultimately 

incorporated into amino acids, primarily selenomethionine, 

which is the organo-selenide that bioaccumulates in food webs 

and produces toxicity (Luoma and Presser, 2009).

The algae and microbes, together with abiotic particlesorbed selenium, constitute the particulate selenium fraction 

in the water column. Particulate matter is ingested by aquatic 

primary consumers, such as zooplankton, insect larvae, 

larval fish, and bivalves (trophic level 2), thus transferring 

Table Q7. National Institutes of Health recommended dietary 

reference intakes (DRIs) for selenium.

[Data are from the Food and Nutrition Board of the Institute of Medicine, 

National Academy of Sciences, and published by the National Institutes of 

Health (2013). Recommendations are for men and for women who are not 

pregnant or nursing. μg/d, microgram per day]

Age

(years)

Recommended

daily allowance

 (μg/d)

Tolerable upper

intake level

(μg/d)

0 to 0.5 1

15 45

0.5 to 1 20 60

1 to 3 20 90

4 to 8 30 150

9 to 13 40 280

14 to 18 55 400

19 to 50 55 400

51+ 55 400

1

This level is called the adequate intake (AI) level because evidence is 

insufficient to develop a recommended daily allowance.

selenomethionine (a naturally occurring amino acid) up the 

food chain. Abiotic selenium, which may be ingested with 

particulate matter, is transformed to selenomethionine by 

the primary consumers, which may, in turn, be consumed by 

predators, such as fish and birds (trophic level 3 and above). 

In this manner, selenium continues to accumulate, and its 

concentrations magnify as it is transferred to higher trophic 

levels. The accumulated concentration determines whether 

toxicity is manifested (Luoma and Presser, 2009; Young and 

others, 2010; U.S. Environmental Protection Agency, 2014b).

Recycling of selenium in the aquatic environment occurs 

when organo-selenide, mostly in the form of selenomethionine,

is returned to the water upon the death and decay of aquatic 

organisms. Some organo-selenide may be oxidized to Se(IV), 

but neither organo-selenide nor Se(IV) will be converted 

to Se(VI) because the reaction is very slow, with a half-life 

of hundreds of years. The net outcome of recycling in a 

watershed is an evolution of selenium from Se(VI), which is 

the primary form of selenium from natural and anthropogenic 

sources, to Se(IV) and organo-selenide (Se-II), which are the 

more bioavailable species (Young and others, 2010).

Selenium enters terrestrial organisms by a similar 

pathway. Accumulator plants take up selenium from soil 

water, usually as Se(VI) or Se(IV), and transform it primarily 

into the amino acid selenomethionine. These accumulators are 

able to incorporate large amounts of selenium—some as high 

as 15,000 ppm (Trelease and Beath, 1949). Livestock grazing 

on these plants accumulate selenomethionine, with potentially 

fatal results.
Q32 Critical Mineral Resources of the United States—Selenium

Aquatic Toxicity and Aquatic Ecosystem Health 

Guidelines

Selenium is bioaccumulative, meaning that fish and 

other aquatic life can contain concentrations greater than 

the surrounding water. The primary source of selenium to 

aquatic life is dietary, and toxicity primarily occurs through 

transfer to eggs and subsequent reproductive effects. Because 

selenium bioaccumulation rates may vary among aquatic 

environments, the EPA has proposed a four-part freshwater 

criterion that includes elements for (1) fish eggs and ovaries, 

set at 15.2 ppm selenium dry weight; (2) whole fish body 

tissue not to exceed 8.1 ppm dry weight, or muscle tissue not 

to exceed 11.8 ppm dry weight; (3) a 30-day average selenium 

concentration not to exceed 4.8 ppb selenium in lotic (flowing) 

waters or 1.3 ppb selenium in lentic (standing) waters, more 

than 1 time in 3 years, on average; and (4) repeated short-term 

exposures where the intermittent water quality concentration 

for both lentic and lotic waters is calculated from the 30-day 

concentration; the background selenium concentration; and 

the fraction of time (over the 30-day period) during which 

elevated selenium concentrations occur. The purpose of this 

criterion with two fish-tissue-based and two water-based 

elements is to protect against adverse reproductive effects 

of selenium on aquatic life, and the criterion apply to total 

selenium concentrations; that is, the sum of all forms of 

selenium (selenate, selenite, organic selenium, and any other 

forms) (U.S. Environmental Protection Agency, 2014a).

Previous to this four-part criterion, the EPA freshwater 

criterion for protection of aquatic life was 5 ppb selenium with 

a lower criterion of 2 ppb selenium for some wetland environments in California (Luoma and Presser, 2009). In marine 

waters, the maximum allowable concentration is 290 ppb 

selenium, whereas the continuous concentration is 71 ppb 

selenium (U.S. Environmental Protection Agency, 2014c). 

It should be noted that there is no international agreement 

on these guidelines. In Canada, the marine and freshwater 

guidelines are 2 ppb selenium, and in Europe, selenium is not 

considered an ecological threat and is not subject to environmental quality standards (Luoma and Presser, 2009; European 

Commission Joint Research Centre Institute for Health and 

Consumer Protection, 2011).

Problems and Future Research

There is no “free lunch.” Selenium extraction from 

mineral resources has led to technological development and 

has greatly benefited our society. At the same time, there are 

environmental consequences of mining that must be recognized and evaluated for every mining project. This chapter 

provides an overview of the selenium life cycle. Although the 

main focus is selenium as a mineral resource and commodity, 

these selenium topics cannot be presented without acknowledging the environmental concerns that accompany the 

presence of selenium in the environment.

Challenges abound for new research on selenium in the 

21st century. Extraction of selenium from the waste products 

(anode slimes) of copper ore processing has been a proven 

technology for many decades, but changes in processing 

technology will affect selenium recovery. Will it be possible 

to recover selenium from other sources economically? 

Coal and other organic carbon sources come immediately 

to mind; these technologies exist but are not economically 

favorable at present. High-grade but low-tonnage deposits 

with selenide minerals may be another source for selenium 

extraction, contingent upon available technology and 

favorable economics.

Other research directions would contribute to the 

scientific understanding of the selenium elemental cycle in 

Earth’s crust. A very basic question concerns the amount 

of selenium in the crust. The current estimate of 0.05 ppm 

selenium in the continental crust dates to 1961. Could this 

estimate be updated and revised, given our current knowledge 

of selenium concentrations in various crustal reservoirs? 

Next, can selenium concentrations and isotopic composition 

be used as an exploration tool for the discovery of concealed 

deposits? Finally, to thoroughly understand the elemental 

cycle of selenium in the earth, it would be useful to compare 

and contrast the magmatic selenium pathway from volcanoes 

to enriched ore deposits, to the sedimentary selenium pathway 

that involves selenium uptake by algae and bacteria in productive marine basins and deposition in organic-rich sediments. 

Are there connections between these two pathways and does 

selenium cycle between them?

Research opportunities exist under the topic of selenium 

environmental concerns as well. Some of the most pressing 

topics involve the bioavailability of selenium and its incorporation into the food web and whether the understanding 

of this cycle can translate into ecosystem and human health 

guidelines. The newly proposed tissue-based criterion for 

selenium in aquatic systems is a direct result of years of 

focused scientific study. Other research might target methods 

to redirect or disrupt the process that transfers selenium from 

rock to biota. Studies that amend selenium-rich soil with 

sulfur, or plant non-selenium-accumulating plants on mine 

waste, are creative attempts to break the rock-biota selenium 

connection. Studies to limit selenium solubility from nanoparticles would aid in keeping new technologies environmentally safe. Regardless of the focus, research into the elemental 

cycle of selenium, including the manufacture and disposal of 

selenium-containing commodities, will continue to increase 

understanding of Earth processes.

Acknowledgments

The author thanks technical reviewers John DeYoung,

Daniel Edelstein, Micheal George, Sandy Neuzil, Theresa Presser,

Klaus Schulz, Bob Seal, and Michael Magyar for improving 

this manuscript with their gifts of time and insight. Bob Seal 

is especially commended for reminding the author that this 
References Cited Q33

chapter is part of a volume on mineral resources, not environmental processes. The USGS publications staff is gratefully 

acknowledged for gentle correction of the author’s mistakes 

and poor writing habits. Keith Labay helped create the deposit 

location maps, and Suzanne Nicholson provided additional 

editorial comments after the technical review was complete.

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Tables Q1 and Q2
Q44 Critical Mineral Resources of the United States—Selenium

Table Q1. Selenium minerals recognized by the International Mineralogical Association.—Continued

[Data are from Mindat.org (2015) and the International Mineralogical Association (IMA) (2014). ?, uncertain according to the IMA; NA, not available]

Mineral name Chemical formula Date of discovery 

or IMA approval Geologic setting and (or) first recorded locality

Achavalite (Fe,Cu)Se 1939 Selenide deposit, Mendoza, Argentina

Aguilarite Ag4

SeS 1891 Low-temperature hydrothermal deposits with silver and selenium 

but low amount of sulfur, Guanajuato, Mexico

Ahlfeldite (Ni,Co)SeO36

 • 2H2O 1935 Alteration product of nickel sulfides and selenides, Pacajake 

Mine, Potosí Department, Bolivia

Allochalcoselite Cu+

PbCu2+

5

(SeO3)2Cl5O2 2004 Scoria cone, Tolbachik volcano, Far Eastern Region, Russia

Antimonselite Sb2

Se3 1993 Uranium deposit No. 504, Guizhou Province, China

Athabascaite Cu5

Se4 1970 Inclusions in and replacements of umangite, as well as stringers 

and veinlets in carbonate veins cutting basalt, Martin Lake 

Mine, Saskatchewan, Canada

Babkinite Pb2

Bi2(S,Se)3 1994 Hydrothermal deposit in sediments, near contact with granite 

intrusion, Nevskoe tungsten-tin deposit, Far Eastern Region, 

Russia

Bambollaite Cu(Se,Te) 1972 Oxidized gold-tellurium ore, Moctezuma Mine, Sonora, Mexico

Bellidoite Cu2

Se 1975 Low-to-moderate temperature hydrothermal deposits, 

Habři Mine, Moravia, Czech Republic

Berzelianite Cu2

Se 1850 Hydrothermal veins, Skrikerum Mine, Östergötland, Sweden

Bohdanowiczite AgBiSe2 1967 Fluorite and quartz veins near magnetite skarn, Kletno, 

Lower Silesia, Poland

Bornhardtite Co2+Co3+

2

Se4 1955 Hydrothermal deposits, Trogtal Quarries, Lower Saxony, 

Germany

Brodtkorbite Cu2

HgSe2 1999 Telethermal selenide vein-type assemblage, Tumiñico Mine, 

La Rioja Province, Argentina

Bukovite Tl2

(Cu,Fe)4Se4 1971 Calcite veins of hydrothermal origin, Bukov Mine, Moravia, 

Czech Republic

Burnsite KCdCu2+

7

(SeO3 )

2

O2Cl9 2000 Scoria cone, Tolbachik volcano, Far Eastern Region, Russia

Cadmoselite CdSe 1957 Sedimentary rocks in an alkaline, reducing environment, 

Ust’ Uyok deposit, eastern Siberia, Russia

Carlosruizite K6

(Na,K)4Na6Mg10(SeO4 )12

(IO3 )12 • 12H2O

1994 Zapiga, Tarapacá Region, Chile

Chalcomenite CuSeO3

 • 2H2O 1881 Secondary mineral in copper-selenium deposits, Cerro de 

Cacheuta, Mendoza, Argentina

Chaméanite (Cu,Fe)4

As(Se,S)4 1980 Late-stage deposits cutting granites, Chaméane uranium deposit, 

Auvergne, France

Chloromenite Cu9

(SeO3)4O2Cl6 1996 Novaya fumarole, Tolbachik volcano, Far Eastern Region, Russia

Chrisstanleyite Ag2

Pd3Se4 1996 Gold-bearing veins, Hope’s Nose, Devon, England

Clausthalite PbSe 1832 Selenium-bearing, low-sulfur hydrothermal deposits, St. Lorenz 

Mine, Lower Saxony, Germany

Cobaltomenite CoSeO3

 • 2H2O 1882 Cerro de Cacheuta, Mendoza, Argentina

Crerarite Pt2–x(Bi,Pb)11(S,Se)11 1994 Copper–nickel–platinum-group-element deposit, Lac Sheen, 

Quebec, Canada

Crookesite Cu7

(Tl, Ag)Se4 1866 Copper-silver-lead-selenium hydrothermal deposits, 

Skrikerum Mine, Östergötland, Sweden

Demesmaekerite Pb2

Cu5(UO2 )2(SeO3 )6(OH)6 • 

2H2

O

1965 Copper-cobalt-manganese-uranium deposit, Musonoi Mine, 

Democratic Republic of the Congo (Congo [Kinshasa])

Derriksite Cu4

(UO2 )(SeO3 )2

(OH)6 • H2O

1971 Copper-cobalt-manganese-uranium deposit, Musonoi Mine, 

Democratic Republic of the Congo (Congo [Kinshasa])
Table Q1 Q45

Table Q1. Selenium minerals recognized by the International Mineralogical Association.—Continued

[Data are from Mindat.org (2015) and the International Mineralogical Association (IMA) (2014). ?, uncertain according to the IMA; NA, not available]

Mineral name Chemical formula Date of discovery 

or IMA approval Geologic setting and (or) first recorded locality

Downeyite

Drysdallite

Dzharkenite

Eldragónite

Eskebornite

Eucairite

Favreauite

Ferroselite

Fischesserite

Francisite

Freboldite

Geffroyite

Georgbokiite

Giraudite

Guanajuatite

Guilleminite

Hakite

Haynesite

Ikunolite1

Ilinskite

Jacutingaite

Jagüéite

Jolliffeite

Junoite

Kalungaite

Kawazulite

Kitkaite

SeO2

Mo(Se,S)2

FeSe2

Cu BiSe (Se ) 6 4 2 

CuFeSe2

AgCuSe

PbBiCu6 4 3 4 O (SeO ) (OH) • H O2

FeSe2

Ag AuSe 3 2

Cu Bi(SeO ) O Cl 3 3 2 2

CoSe

(Cu,Fe,Ag) (Se,S) 9 8

Cu (SeO ) O Cl 5 3 2 2 2

Cu [Cu (Fe,Zn) ]As Se 6 4 2 4 13

Bi Se 2 3

Ba(UO2 (SeO3 )3 (OH)4 )2  •  

3H O2

Cu [Cu Hg ]Sb Se 6 4 2 4 13

(UO ) (OH) (SeO ) 2 3 2 3 2 2  • 5H O

Bi (S,Se) 4 3

(Na,K)Cu5(SeO3)2 Cl O2 3

Pt2HgSe3

Cu2Pd3Se4

NiAsSe

Cu Pb Bi (S,Se) 2 3 8 16

PdAsSe

Bi2

Te2Se

NiTeSe

1971

1973

1995

2011

1949

1818

2014

1955

1971

1990

1957

1982

1996

1982

1873

1965

1970

1991

1959

1996

2011

2002

1991

1975

2004

1970

1961

Burning anthracite coal deposits, Pennsylvania, United States

Oxidation zone of uranium deposits, Kampijimpanga, 

North-Western Province, Zambia

Suluchekinskoye selenium-uranium deposit, Almaty Province, 

Kazakhstan

Hydrothermal veins with selenium > sulfur, El Dragón Mine, 

Potosí Department, Bolivia

Low-temperature hydrothermal deposits, Eskaborn Adit, 

Saxony-Anhalt, Germany

Copper-silver-lead-selenium hydrothermal deposits, Skrikerum 

Mine, Östergötland, Sweden

Hydrothermal veins with selenium > sulfur, El Dragón Mine, 

Potosí Department, Bolivia

Red bed deposits of Colorado plateau type, Ust’ Uyok deposit, 

eastern Siberia, Russia

Carbonate vein in epithermal hydrothermal gold-uranium-selenium deposit, Předbořice, Bohemia, Czech Republic

Iron Mine, Iron Monarch open cut, South Australia, Australia

Dolomite veinlets, Trogtal Quarries, Lower Saxony, Germany

Late-stage veins cutting granite, Chaméane uranium deposit, 

Auvergne, France

Scoria cone, Great Fissure eruption, Far Eastern Region, Russia

Late-stage veins cutting granite, Chaméane uranium deposit, 

Auvergne, France

Low- to moderate-temperature hydrothermal deposits, 

Santa Catarina Mine, Guanajuato, Mexico

Oxidized zone of a copper-cobalt-manganese-uranium deposit, 

Musonoi Mine, Democratic Republic of the Congo (Congo 

[Kinshasa])

In epithermal calcite veins, Předbořice, Bohemia, Czech Republic

Oxidized uranium deposits, Repete Mine, Utah, United States

Xenothermal veins with silver-gold-copper-lead-tin-tungstenzinc, Ikuno Mine, Honshu Island, Japan

Scoria cone, Tolbachik volcano, Far Eastern Region, Russia 

Iron-gold-palladium deposit, Cauê Mine, Minas Gerais, Brazil

Telethermal selenide vein-type deposit, El Chire prospect, La Rioja Province, Argentina

Uraninite vein, Fish Hook Bay, Saskatchewan, Canada

Gold-copper-bismuth deposit, Juno Mine, Northern Territory, 

Australia

Gold deposit, related to a peraluminous granite, Buraco do Ouro 

Mine, Goiás, Brazil

Epithermal gold-silver-tellurium-manganese veins, Kawazu Mine, 

Honshu Island, Japan

Low-grade uranium and selenium mineralization in carbonatebearing veinlets, Kitka River Valley, northern Finland
Q46 Critical Mineral Resources of the United States—Selenium

Table Q1. Selenium minerals recognized by the International Mineralogical Association.—Continued

[Data are from Mindat.org (2015) and the International Mineralogical Association (IMA) (2014). ?, uncertain according to the IMA; NA, not available]

Mineral name Chemical formula Date of discovery 

or IMA approval Geologic setting and (or) first recorded locality

Klockmannite

Krut’aite

Kullerudite

Kurilite

Laitakarite

Laphamite

Larisaite

Litochlebite

Luberoite

Mäkinenite

Mandarinoite

Marthozite

Mgriite

Miessiite

Milotaite

Molybdomenite

Mozgovaite

Munakataite

Naumannite

Nestolaite

Nevskite

Nicksobolevite

Nordströmite2

Olsacherite

Oosterboschite

Orlandiite

CuSe

CuSe2

NiSe2

Ag Te Se 8 3

Bi Se S 4 2

As (Se,S) 2 3

Na(H O)(UO ) (Se4+O )O  •  3 2 3 3 2

4H O2

Ag PbBi Se 2 4 8

Pt Se 5 4

γ-NiSe

Fe +2(Se 3

4+O )3 3 2  • 6H O

Cu(UO ) (SeO ) (OH)  •  2 3 3 2 2

8H O2

Cu AsSe 3 3

Pd Te Se 11 2 2

PdSbSe

PbSeO3

PbBi (S,Se) 4 7

Pb Cu (Se4+O )(SO )(OH) 2 2 3 4 4

Ag Se 2

CaSeO  •  3 2 H O

Bi(Se,S)

Cu (SeO ) O Cl 7 3 2 2 6

CuPb Bi (Se S ) 3 7 4 10 

Pb (SeO )(SO ) 2 4 4 

(Pd,Cu) Se 7 5

Pb (SeO )O 10 4 7 4 2 Cl  • H O

1928

1972

1964

2009

1959

1985

2002

2009

1992

1967

1978

1969

1982

2006

2003

1882

1998

2007

1828

2013

1984

2013

1978

1969

1970

1998

Silver-copper-selenium deposit, Las Asperezas Mine, La Rioja 

Province, Argentina

Uranium deposit, Petrovice, Moravia, Czech Republic

Alteration of wilkmanite in low-grade uranium and selenium 

mineralization, Kitka River Valley, northern Finland

Kuril islands gold deposit, Prasolovskoe gold deposit, 

Far Eastern Region, Russia

Copper-gold deposit, Orijärvi, southwestern Finland

Burning anthracite coal deposit, Burnside, Pennsylvania, 

United States

Uranium deposit, Repete Mine, Blanding, San Juan County, 

Utah, United States

Quartz-uranium-carbonate deposit, Zálesí, Moravia, 

Czech Republic

Lubero region, Kivu, Democratic Republic of the Congo (Congo 

[Kinshasa])

Low-grade uranium and selenium mineralization, Kitka River 

Valley, northern Finland

Oxidation product of penroseite and pyrite, Virgen de Surumi 

Mine (Pacajake Mine), Potosí Department, Bolivia

Copper-cobalt-manganese-uranium deposit, Musonoi Mine, Katanga, Democratic Republic of the Congo (Congo [Kinshasa])

Silver-uranium deposit, Schlema-Hartenstein District, Saxony, 

Germany

Placer deposits, Miessijoki River, Lapland Region, Finland

Gold-uranium-selenium deposit, Redbořice, Bohemia, 

Czech Republic

Cacheuta Mine, Mendoza, Argentina

Magmatic-hydrothermal system, La Fossa crater, Vulcano Island, 

Sicily, Italy

Oxide zone of aguilarite-bearing lead-zinc-gold-silver ores, 

Kato Mine, Kyushu Region, Japan

Selenium mineralization in hematite mine, Tilkerode, SaxonyAnhalt, Germany

In sandstone host, Little Eva Mine, Utah, United States

Nevskoe tungsten-tin deposit, Far Eastern Region, Russia

Scoria cone, Tolbachik volcano, Far Eastern Region, Russia

Copper-gold-silver-lead-zinc deposit, Falun Mine, Falun, 

Dalarna, Sweden

Silver-selenium deposit, Virgen de Surumi Mine (Pacajake Mine), 

Potosí Department, Bolivia

Copper-cobalt-manganese-uranium deposit, Musonoi Mine, 

Katanga, Democratic Republic of the Congo (Congo [Kinshasa])

Galena-arsenopyrite deposit, Baccu Locci Mine, Sardinia, Italy
Table Q1 Q47

Table Q1. Selenium minerals recognized by the International Mineralogical Association.—Continued

[Data are from Mindat.org (2015) and the International Mineralogical Association (IMA) (2014). ?, uncertain according to the IMA; NA, not available]

Mineral name Chemical formula Date of discovery 

or IMA approval Geologic setting and (or) first recorded locality

Padmaite

Palladseite

Parageorgbokiite

Paraguanajuatite

Pekoite1

Penroseite

Penzhinite

Permingeatite

Petrovicite

Petrovskaite1

Piretite

Plumboselite

Poubaite

Prewittite

Proudite

Sabatierite

Sarrabusite

Schlemaite3

Schmiederite

Sederholmite

Selenium

Selenojalpaite

Selenopolybasite

Selenostephanite

Skippenite

Sofiite

Součekite

PdBiSe

Pd Se 17 15

Cu (SeO ) O Cl 5 3 2 2 2

Bi Se 2 3

PbCuBi (S,Se) 11 18

(Ni,Co,Cu)Se2

(Ag,Cu) Au(S,Se) 4 4

Cu SbSe 3 4

Cu HgPbBiSe 3 5

AuAg(S,Se)

Ca(UO ) (SeO ) (OH)  •  2 3 3 2 4

4H O2

Pb O (SeO ) 3 2 3 

PbBi (Se,Te,S) 2 4

KPb ZnCu (SeO ) O Cl 1.5 6 3 2 2 10

CuPb Bi (S,Se) 7.5 9.33 22

Cu TlSe 6 4

Pb CuCl (SeO ) 5 4 3 4

(Cu,□) (Pb,Bi)Se 6 4

Pb Cu (Se +O )(Se +O ) 2 2 6 4 4 3 

(OH)4

 ß-NiSe

Se

Ag CuSe 3 2

[(Ag,Cu) (Sb,As) (S,Se) ] 6 2 7 

[Ag Cu(S,Se) Se ] 9 2 2 

Ag Sb(Se,S) 5 4

Bi TeSe 2 2

Zn (SeO )Cl 2 3 2

PbCuBi(S,Se)3

1991

1977

2006

1948

1975

1926

1984

1971

1976

1984

1996

2010

1978

2002

1975

1976

1997

2003

1962

1964

 1828?

2004

2006

1985

1987

1989

1979

Uranium-vanadium deposit, Srednyaya Padma Mine, 

Northern Region, Russia

Itabira, Iron Quadrangle, Minas Gerais, Brazil

Scoria cone, Tolbachik volcano, Far Eastern Region, Russia

Gold-sulfur deposit, Santa Catarina Mine, Guanajuato, Mexico

Gold-copper-bismuth deposit, Juno Mine, Northern Territory, 

Australia

Silver-selenium deposit, Virgen de Surumi Mine (Pacajake Mine), 

Potosí Department, Bolivia

Sergeevskoye gold-silver occurrence, Far Eastern Region, Russia

Gold-uranium-selenium deposit, Předbořice, Bohemia, 

Czech Republic

Uranium deposit, Petrovice, Moravia, Czech Republic

Maikain gold deposit, Pavlodar Province, Kazakhstan

Copper-cobalt-nickel-uranium deposit, Shinkolobwe Mine, Katanga, Democratic Republic of the Congo (Congo [Kinshasa[)

Copper-lead-zinc-silver-germanium-cadmium deposit in dolomite, Tsumeb Mine, Otjikoto Region, Namibia

Uraninite deposit with sulfur-selenium, Oldřichov, Bohemia, 

Czech Republic

In fumarole field, Tolbachik volcano, Far Eastern Region, Russia 

Gold-copper-bismuth deposit, Juno Mine, Northern Territory, 

Australia

Uranium deposit, Bukov Mine, Moravia, Czech Republic

Galena-arsenopyrite deposit, Baccu Locci Mine, Sardinia, Italy

Hydrothermal selenium-bearing dolomite-ankerite vein, 

Schlema-Hartenstein District, Saxony, Germany

Hydrothermal veinlets in sedimentary host, selenium > sulfur, 

El Cóndor Mine, Los Llantenes district, La Rioja Province, 

Argentina

In calcite veins with nickel selenides associated with low-grade 

uranium mineralization, Kitka River Valley, northern Finland

NA

Copper-silver-lead-selenium hydrothermal deposits, 

Skrikerum Mine, Östergötland, Sweden

Silver deposit, De Lamar Mine, Idaho, United States

Volcanogenic silver-gold deposit, Rudnaya Sopka, Far Eastern 

Region, Russia

Otish Mountains uranium deposit, Quebec, Canada

Scoria cone, Tolbachik volcano, Far Eastern Region, Russia

Uranium deposit with sulfur-selenium, Oldřichov, Bohemia, 

Czech Republic
Q48 Critical Mineral Resources of the United States—Selenium

Table Q1. Selenium minerals recognized by the International Mineralogical Association.—Continued

[Data are from Mindat.org (2015) and the International Mineralogical Association (IMA) (2014). ?, uncertain according to the IMA; NA, not available]

Mineral name Chemical formula Date of discovery 

or IMA approval Geologic setting and (or) first recorded locality

Stilleite

Sudovikovite

Telluronevskite

Tiemannite

Tischendorfite

Trogtalite

Trüstedtite

Tsnigriite

Tyrrellite

Umangite

Verbeekite

Vihorlatite

Watkinsonite

Weibullite

Wilkmanite

Wittite

Zincomenite

ZnSe

PtSe2

Bi TeSe 3 2

HgSe

Pd Hg Se 8 3 9

CoSe2

Ni Se 3 4

Ag SbTe (S,Se) 9 3 3

Cu(Co3+,Ni3+) Se 2 4

Cu Se 3 2

PdSe2

Bi Se Te 24 17 4

Cu PbBi (Se,S,Te) 2 4 8

Pb Bi Se S 5 8 7 11

Ni Se 3 4

Pb Bi (S,Se) 8 10 23

ZnSeO3

1956

1995

1993

1855

2001

1955

1967

1991

1952

1891

2001

1988

1985

1910

1964

1924

2014

Uranium deposit with copper, cobalt, and nickel, Shinkolobwe Mine, 

Katanga, Democratic Republic of the Congo (Congo [Kinshasa])

Uranium-vanadium deposit, Srednyaya Padma Mine, 

Northern Region, Russia

Selenium and tellurium mineralization in hydrothermal veins, 

in argillic andesite host, Poruba pod Vihorlatom, Michalovce 

County, Košice Region, Slovakia

Ancient lead and zinc mine, working veins of the Burgstadt 

vein system, St. Lorenz Mine, Lower Saxony, Germany

Selenium mineralization in hematite mine, Eskaborn Adit, 

Tilkerode, Saxony-Anhalt, Germany

Abandoned greywacke quarries, Trogtal Quarries, 

Lower Saxony, Germany

Low-grade uranium and selenium mineralization in an albite 

diabase, Kitka River Valley, northern Finland

Black-shale-hosted Vysokovol’tnoye silver-gold deposit, 

Bel’tau Mountains, Kyzylkum Desert, Uzbekistan

Ato Bay, Beaverlodge Lake, Goldfields District, Beaverlodge Region, 

Saskatchewan, Canada; and Eagle Claims uranium-copperselenium occurrence, Eagle shaft area, Melville Lake, Goldfields District, Beaverlodge Region, Saskatchewan, Canada

Las Asperezas silver mine (Sarmiento deposit, Veta Sarmiento), 

Sierra de Cacho, Villa Castelli, General Lamadrid Department, 

La Rioja Province, Argentina; and abandoned greywacke quarries, 

Trogtal Quarries, Lautenthal, Harz, Lower Saxony, Germany

Copper-cobalt-manganese-uranium deposit. Known for rare assemblage of uranyl selenites, Musonoi Mine, Kolwezi, Kolwezi 

District, Katanga Copper Crescent, Katanga (Shaba), Democratic Republic of the Congo (Congo [Kinshasa])

Poruba pod Vihorlatom, Michalovce County, Košice Region, 

Slovakia

Otish Mountains uranium deposit, Otish Mountains, Baie-James, 

Nord-du-Québec, Quebec, Canada

Copper mine with minor amount of gold, Falun Mine, Falun, 

Dalarna, Sweden

Low-grade uranium and selenium mineralization in an albite 

diabase, Kitka River Valley, Kuusamo, northern Finland

Falun Mine, Falun, Dalarna, Sweden

Northern fumarole field, first scoria cone, Northern Breakthrough 

(North Breach), and Great Fissure eruption (Main Fracture), 

Tolbachik volcano, Far Eastern Region, Russia

1

The IMA chemical formula does not show selenium substitution for sulfur.

2

The IMA chemical formula is Pb3CuBi (S,Se) 7 14.

3

The symbol □ in the chemical formula denotes an empty atomic site that can be filled (or not), depending upon compositional variation.
Table Q2 Q49

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Seleniumconcentration Notes Reference(s)

Australia, eastern

Volcanic-hosted massive sulfideAgincourt deposit Pyrite <6 to 9 ppm — Huston and others (1995)

 Do. Dry River South 

deposit

 do. <6 to 760 ppm — Do.

 Do. Mount Chalmers 

deposit

 do. <5 to 200 ppm — Do.

 Do. Rosebery deposit do. <6 to 33 ppm — Do.

 Do. Waterloo deposit do. <6 to 180 ppm — Do.

Australia, southern

Protoerozoic Olympic

Dam IOCG 

Olympic Dam Bornite 18.8 ppm — Cook and others 

(2011)

 Do. do. Chalcopyrite 8.7 ppm — Do.

Australia, western

Au-Ag-PGE-Hg-Se 

polymetallic veins

Pilbara region,

Copper Hills

Chrisstanleyite 

(Ag2

Pd2Se4)

38.5 wt. % — Nickel (2002)

 Do. do. Cu-chrisstanleyite 

(Cu2Pd2Se4)

41.33 wt. % — Do.

 Do. do. Naumannite (Ag2Se) 28.8 wt. % — Do.

 Do. do. Oosterboschite 

(Pd4Cu3Se5)

45 wt. % — Do.

 Do. do. Umangite (Cu3Se2) 13.7 wt. % — Do.

Brazil, northern

Au-Pd-Pt Serra Pelada Bulk rock of bonanza 

grade Au-Pd-Pt

21 to 36 ppm Disaggregated, clayey material Cabral and others (2002)

 Do. do. Palladseite (Pd17Se15) n.r. — Do.

 Do. do. Sudovikovite (PtSe2) 11.5 to 12.2 wt. % Pd-Pt-Se inclusions in palladian 

gold aggregates

 Do.

Canada, British Columbia

Au-pyrite quartz 

veins

Caribooa Pyrite <15 ppm Low to negligible Se content of Au-quartz 

deposits is attributed to Se-poor magmatic sources and (or) wall rocks.

Hawley and 

Nichol 

(1959)

Canada, Manitoba

Massive Cu-Fe-(Zn)-

S replacement 

Flin Flon Chalcopyrite 145 to 470 ppm Major concentration of Se found in the 

Canadian Precambrian deposits occurs 

in the massive Cu, Fe S replacement 

types of deposits, hydrothermal in origin.

Hawley and 

Nichol 

(1959)

 Do. do. Pyrite <15 to 220 ppm — Do.

 Do. do. Pyrrhotite 375 ppm — Do.

Canada, New Brunswick

Cu-Fe-Zn-S banded 

ores

Heath Steele Pyrite 20 ppm — Hawley and 

Nichol 

(1959)

Massive Cu-Fe-(Zn)-

S replacement, 

banded

 do. Pyrrhotite 21 ppm — Do.

Canada, Newfoundland and Labrador

Ni-Cu-Co Nain Plutonic Suite, 

Voisey’s Bay, 

Ovid ore zone

Chalcopyrite 38 to 70 ppm — Ripley and others (2002)

 Do. do. Pentlandite 53 to 93 ppm — Do.
Q50 Critical Mineral Resources of the United States—Selenium

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Selenium concentration Notes Reference(s)

Canada, Northwest Territories

Au-pyrite quartz 

veins

Yellowknife, Discovery 

Mine

Pyrrhotite 26 to 30 ppm — Hawley and 

Nichol 

(1959)

Canada, Ontario

Sudbury Ni-Cu-PGE 

sulfide deposit

Sudbury Complex, 

Creighton deposit, 

Western Branch

Chalcopyrite 49.43 to 124.59 ppm — Dare and others 

(2010)

 Do.

 Do.

 Do.

 Do.

 do.

 do.

 do.

Creighton, Frood, Garson, Levack, Murray 

Mines

Pentlandite

Pyrite

Pyrrhtotite

Cu-Ni ores, no specific 

minerals

49.34 to 113.66 ppm

60.56 to 151.83 ppm

63.52 to 133.98 ppm 

20 to 82 ppm

—

—

—

—

 Do.

 Do.

 Do.

Nriagu and 

Wong (1983)

Au-pyrite quartz 

veins

 Do.

 Do.

 Do.

 Do.

 Do.

Banded sideritepyrite

Cathroy-Larder, Hardrock,

Howey-Hasaga

(Red Lake), Renabi

Hollinger

 do.

McIntyre Mine

 do.

McIntyre Mine

Michipicoten Iron 

Formation, Rand 

No. 2 Mine

Pyrite

Chalcopyrite

Pyrrhotite

High-temperature pyrite

Low-temperature pyrite

Pyrite

Pyrrhotite

<15 ppm

23 ppm

26 to 30 ppm

<15 to 83 ppm

<15 to 110 ppm

<15 to 110 ppm

<15 ppm

Low to negligible Se content of Au-quartz 

deposits is attributed to Se-poor magmatic sources and (or) wall rocks.

—

—

—

In the McIntyre Mine, the authors iden -

tified low-temperature pyrite with a 

higher range of Se concentrations than 

high-temperature pyrite.

—

—

Hawley and 

Nichol 

(1959)

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do. Michipicoten Iron Frm, 

Rand No. 2 Mine

Pyrite <15 ppm — Do.

Diabase dikes Algoma U district, 

Can-Met Mine

Pyrrhotite <15 ppm — Do.

Magnetite (replacement)

Massive Cu-Fe-(Zn)-

S replacement 

 Do.

 Do.

Ni-Cu-S ore

 Do.

 Do.

 Do.

 Do.

Pyritiferous U banded 

or bedded deposits

Marmora

Manitouwadge, 

Geco Mine,

 do.

 do.

Porcupine, Alexo Mine

Sudbury Complex 

 do.

 do.

 do.

Algoma, Blind 

River area

Pyrite

Chalcopyrite

Pyrite

Pyrrhotite

 do.

Chalcopyrite

Pentlandite

Pyrite

Pyrrhotite

Pyrite

<15 ppm

130 to 540 ppm

44 to 310 ppm

<15 to 595 ppm

17 ppm

11 to 165 ppm

30 to 160 ppm

25 to 130 ppm

17 to 230 ppm

<15 to 79 ppm

Author noted that Se showed no order of 

preference among individual sulfide, 

phases, but showed a tendency for progressive concentration of Se with 

the order of crystallization.

—

—

—

—

—

—

—

—

—

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.
Table Q2 Q51

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Selenium concentration Notes Reference(s)

Canada, Quebec

Au-pyrite quartz 

veins

 Do.

Cu-Fe-Zn-S banded 

Massive Cu-Fe-(Zn)-

S replacement 

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

Granada

Powell Rouyn

Suffield

Aldermac

 do.

 do.

CampbellChibougamau

 do.

 do.

Noranda

 do.

 do.

Normetal

 do.

 do.

Quemont

 do.

 do.

Waite Amulet

Pyrite

 do.

 do.

Chalcopyrite

Pyrite

Pyrrhotite

Chalcopyrite

Pyrite

Pyrrhotite

Chalcopyrite

Pyrite

Pyrrhotite

Chalcopyrite

Pyrite

Pyrrhotite

Chalcopyrite

Pyrite

Pyrrhotite

Chalcopyrite

<15 ppm

23 to 24 ppm

<15 ppm

135 ppm

56 ppm

74 ppm

25 to 94 ppm

51 ppm

<15 to 82 ppm

360 to 535 ppm

33 to 1,000 ppm

36 to 555 ppm

23 to 260 ppm

<15 to 320 ppm

15 to 235 ppm

40 to 280 ppm

10 to 250 ppm

54 to 425 ppm

395 ppm

Low to negligible Se content of Au-quartz 

deposits is attributed to Se-poor magmatic sources and (or) wall rocks.

—

—

—

—

—

—

In pyrite, pyrrhotite, and chalcopyrite, 

the Se content is greatest when the mineral phase is located in a massive Cu-S 

replacement body.

—

—

—

—

—

—

—

—

—

—

—

Hawley and 

Nichol 

(1959)

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

Canada, Saskatchewan

Au-pyrite quartz 

veins

Foudulac, Algold Pyrite 23 to 24 ppm — Hawley and 

Nichol 

(1959)

Chile, northern

High sulfidation, 

epithermal AuAg-Cu

Volcanic-hosted 

strata-bound Cu

 Do.

Pascua deposit

Mantos Blancos 

deposit

 do.

Main stage pyrite

Chalcopyrite

Bornite

0.01 to 0.06 wt. %

0.34 to 4.76 ppm

0.16 to 5.11 ppm

—

—

—

Chouinard and 

others (2005)

Reich and others (2013)

 Do.

Democratic Republic of the Congo (Congo [Kinshasa])

U-Cu-Co

 Do.

 Do.

 Do.

Katanga Copperbelt, 

Musonoi Mine

 do.

 do.

 do.

Berzelianite (Cu Se) 2

Digenite (Cu S) 1.8

Athabascaite (Cu Se ) 5 4

Trogtalite (CoSe )

2

28.65 to 40.40 wt. %

0.63 to 19.05 wt. %

12.48 to 34.37 wt. %

69.43 to 70.23 wt. %

—

—

—

—

Pirard and Hatert (2008)

 Do.

 Do.

 Do.
Q52 Critical Mineral Resources of the United States—Selenium

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Selenium concentration Notes Reference(s)

Europe, southeastern

Cretaceous skarn, 

epithermal and 

porphyry deposits

Baita Bihor, Chelopeck, 

Elasite, Ocna de Fier,

Radka, Sasca Montana; 

Banatitic magmatic 

and metallogenetic 

belt (BMMB)

Bornite 1.5 to 1,046 ppm — Cook and others 

(2011)

 Do. Baita Bihor, Ocna 

de Fier; Banatitic 

magmatic and 

metallogenetic belt 

(BMMB)

Chalcopyrite 8.7 to 538 ppm — Do.

 Do. do. Cu S sulfides 2–x 151 to 383 ppm — Do.

Italy, Sardinia

Pb-As Baccu Locci Mine Cupropavonite 0.26 wt. % — Pirri (2002)

 Do. do. Galena 0.64 wt. % — Do.

 Do. do. Greenockite 0.13 wt. % — Do.

 Do. do. Se-rich galena (clausthalite?)12.99 wt. % — Do.

 Do. do. Sphalerite 0.12 wt. % — Do.

 Do. do. Wittichenite 0.22 wt. % — Do.

Japan, Hokkaido

Ag-Bi-Te-Se-S

 Do.

 Do.

 Do.

Suttsu Mine

 do.

 do.

 do.

Kawazulite (Bi (Te 2.01 1.42 

Se S ) ) 1.15 0.42 2.99

Selenian bismuthinite 

(Bi (S Se ) ) 2.00 2 0.25 3.01

Selenian pavonite 

(Ag (Bi Pb ) 1.00 2.86 0.18 3.04

(S Se Te ) ) 3.54 1.36 0.06 4.96

Unnamed (Bi (Te, Se, 3

S) )4

12.9 wt. %

3.7 wt. %

11.1 wt. %

9.5±4 wt. %

—

—

—

—

Shimizu and 

others (1995)

 Do.

 Do.

 Do.

Kazakhstan

Porphyry Cu-Mo Boshchekul’ deposit Chalcopyrite II 64 to 123.7 ppm — Filimonova 

(1972)

 Do. do. Chalcopyrite III 111 to 225 ppm — Do.

 Do. do. Chalcopyrite IV 20.6 to 44.4 ppm — Do.

 Do. do. Magnetite 5 to 7 ppm — Do.

 Do. do. Molybdenite I 120 to 781 ppm — Do.

 Do. do. Molybdenite II 235 to 800 ppm — Do.

 Do. do. Pyrite I 42 to 105 ppm — Do.

 Do. do. Pyrite III 50 to 203 ppm — Do.

 Do. do. Pyrite IV 16 to 100 ppm — Do.

 Do. do. Sphalerite 29.5 ppm — Do.
Table Q2 Q53

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Selenium concentration Notes Reference(s)

Pacific Ocean, 13° North, East Pacific Rise

Cu sulfides, Black 

smoker active 

chimneys

Fe-Cu massive sulfides,

mature hydrothermal edifice

 Do.

Hi-T hydrothermal 

sulfide deposits: 

axial graben

Off-axis seamounts

 do.

Chalcopyrite

Chalcopyrite

Pyrite

100 to 2,000 ppm

0 to 2,500 ppm

0 to 1,500 ppm

High Se values occur in high-temperature 

mineral assemblages, found in the inner 

part of the deposit in equilibrium with 

pure, unmixed hydrothermal water; 

lowest Se values in outer part of deposit, 

influenced by mixing of hydrothermal 

waters and seawater.

—

—

Auclair and others (1987)

 Do.

 Do.

Portugal

Cu-Zn ore

Pb-Zn ore

Neves Corvo deposit, 

Lombador orebody

Neves Corvo deposit, 

Zambugal orebody

Se-galena, up to 30% 

of clausthalite 

(PbSe), and junoite 

(Pb Cu Bi [S,Se] ) 3 2 8 16

 do.

2,560 ppm (avg)

3,220 ppm (avg)

—

—

Pinto and others 

(2013)

 Do.

Russia, Siberia

Cu-Ni-PGE

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

Noril’sk, Oktyabr 

deposit

 do.

 do.

 do.

 do.

 do.

 do.

Noril’sk, Talnakh 

deposit

 do.

 do.

 do.

 do.

Chalcopyrite

Cubanite (CuFe S ) 2 3

Millerite, NiS

Moikhukite

Pentlandite

Pyrrhotite

Troilite (FeS)

Chalcopyrite

Cubanite

Pentlandite

Pyrrhotite

Taknakhite

44 to 148 ppm

92 to 94 ppm

496 ppm

94 ppm

56 to 213 ppm

50 to 98 ppm

87 ppm

69 to 152 ppm

85 to 181 ppm

139 to 190 ppm

70 to 74 ppm

115 to 190 ppm

—

—

—

—

—

—

—

—

—

—

—

—

Kovalenker and 

others (1975)

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

Russia, Ural Mountains

Cu-S

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

Karabash group

 do.

 do.

 do.

 do.

 do.

 do.

Bornite

Chalcopyrite (from Cusphalerite ores)

Chalcopyrite (from late 

quartz-carbonate 

veinlets)

Galena 

Pyrite

Sphalerite

Tetrahedrite-tennantite 

(from cupreous 

pyrite)

18 ppm

3.8 to 216 ppm

11 to 16 ppm

0.5 ppm

3 to 110 ppm

16 ppm

81 ppm

—

—

—

—

—

—

—

Yushoko-Zakharova and 

others (1978)

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.
Q54 Critical Mineral Resources of the United States—Selenium

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Selenium concentration Notes Reference(s)

Russia, Ural Mountains—Continued

Cu-S

 Do.

 Do.

 Do.

 Do.

Cu-Zn-S

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

FeS2

 Do.

 Do.

Karabash group

Mauk deposit

 do.

 do.

 do.

Gay deposit

 do.

 do.

 do.

 do.

 do.

III Internatsional 

deposit

 do.

 do.

 do.

 do.

 do.

 do.

Levikha XIV deposit

 do.

Sibay deposit

 do.

 do.

 do.

 do.

Zyuzel’skoye deposit

 do.

 do.

Tetrahedrite-tennantite 

(from quartz-carbonate

veins and veinlets)

Chalcopyrite (from Cupyrrhotite ores)

Pyrite

Pyrrhotite

Sphalerite

Bornite

Chalcopyrite (from Cusphalerite ores)

Galena 

Pyrite

Sphalerite

Tetrahedrite-tennantite 

(from cupreous pyrite)

Bornite

Chalcocite

Chalcopyrite (from Cusphalerite ores)

Chalcopyrite (from late 

quartz-carbonate 

veinlets)

Pyrite

Sphalerite

Tetrahedrite-tennantite 

(from cupreous 

pyrite)

Chalcopyrite (from Cusphalerite ores)

Tetrahedrite-tennantite 

(from quartz-carbonate 

veins and veinlets)

Chalcopyrite (from Cupyrrhotite ores)

Chalcopyrite (from Cusphalerite ores)

Pyrite

Pyrrhotite

Sphalerite

Chalcopyrite (from Cusphalerite ores)

Chalcopyrite (from late 

quartz-carbonate 

veinlets)

Pyrite

45 to 100 ppm

180 to 325 ppm

n.d. to 119 ppm

21 to 145 ppm

140 to 187 ppm

10 to 120 ppm

5 to 150 ppm

30 ppm

5 to 550 ppm

8 to 84 ppm

10 ppm

Trace to 10 ppm

n.d. to 200 ppm

33 to 220 ppm

Trace to 30 ppm

Trace to 146 ppm

2 to 16 ppm

9 ppm

5 to 210 ppm

13 to 20 ppm

80 to 265 ppm

10 to 130 ppm

1 to 88 ppm

53 to 126 ppm

0.5 to 13 ppm

160 ppm

Trace to 54 ppm

Trace to 495 ppm

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

—

Yushoko-Zakharova and 

others (1978)

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.

 Do.
Table Q2 Q55

Table Q2. A summary of selenium concentrations in various selenides and sulfides from deposits around the world.—Continued

[avg, average; b.d.l., below detection limit; Do., do., ditto; IOCG, iron oxide-copper-gold; n.d., not detected; n.r., not reported; ppm, part per million; 

wt. %, weight percent; —, none. Abbreviations for elements: Ag, silver; As, arsenic; Au, gold; Bi, bismuth; Co, cobalt; Cu, copper; Fe, iron; Hg, mercury; 

Mo, molybdenum; Ni, nickel; Pb, lead; Pd, palladium; PGE, platinum-group element; Pt, platinum; S, sulfur; Se, selenium; Te, tellurium; Ti, titanium; 

Tl, thallium; U, uranium; Zn, zinc]

Deposit type and (or) 

chemistry Mine/deposit name Mineral phases Selenium concentration Notes Reference(s)

Sweden and Norway

Late Proterozoic 

metamorphic Cuvein systems

 Do.

 Do.

Glava, Grusen, 

Moberg, Tinnsjå, 

Tjøstølflaten

Tinnsjå

Grusen, Moberg, 

Tinnsjå

Bornite

Chalcopyrite

Cu S sulfides 2–x

2.8 to 4,617 ppm

20.7 ppm

b.d.l. to 1,291 ppm

Se concentrations in bornite show promise 

as a vector for exploration and may be 

able to track proximal to distal zonation.

—

—

Cook and others 

(2011)

 Do.

 Do.

Sweden, southeastern

Cu-Ag-Tl-Se Skrikerum deposit, 

Småland

Selenojalpaite 

(Ag CuSe ) 3 2

29.01 wt. % — Bindi and Pratesi (2005)

Spain

Carbonate-hosted 

epithermal gold

Providencia Mine, 

Leon

Villamaninite 

((Cu,Ni,Co,Fe)

(S,Se,Te) )2

1.89 to 9.12 wt. % — Paniagua 

(1995)

Turkey, northeastern

Cu

 

 

 

Do.

Do.

Do.

Murgul deposit

 do.

 do.

 do.

Aikinite )Pb Cu 0.96 1.10 

Bi S Se Te ) 0.98 2.73 0.20 0.03

Galena (Pb Ag 1.00 0.01

Bi S Se ) 0.01 0.95 0.03

Tennantite (Cu Zn 9.66 2.02 

Fe As Bi Te 0.03 3.90 0.35 0.08 

S Se and 12.95 0.01

Cu Zn Fe 9.59 1.97 0.03 

As Bi Te 3.76 0.44 0.11 

S Se ) 13.03 0.07

Tetradymite (Bi Pb 1.76 0.22 

Cu Te S Se ) 0.06 1.72 0.68 0.56

2.75 wt. %

1 wt. %

0.05 and 

0.36 wt. %

6.33 wt. %

—

—

—

—

Willgallis and 

others (1990)

 Do.

 Do.

 Do.

United States, Alaska

Porphyry Cu-Au-Mo Bristol Bay, Pebble 

deposit

Pb-Zn Southern Brooks 

Range, Red Dog

Sample pyrite-4

Galena and sphalerite 

(ore minerals), and 

associated pyrite

142 ppm

<1 ppm

—

Se distributed evenly among these 

mineral phases

Gregory and 

others (2013)

Baker and others (2000)

United States, Minnesota

Cu-Ni-PGE

 Do.

 Do.

Duluth Complex, Babbitt deposit

 do.

 do.

Cubanite

Pentlandite

Pyrrhotite

6 to 157 ppm

24 to 363 ppm

24 to 169 ppm

—

—

—

Ripley (1990)

 Do.

 Do.

United States, Nevada

Au-Ag veins

 Do.

Humboldt Coounty, 

National District, 

Buckskin Mountain

 do.

HgSeSCl phases

Stibnite

2.12 to 21.91 wt. %

0.10 to 3.62 wt. %

—

—

Vikre (2007)

 Do.

Tellurium

By Richard J. Goldfarb, Byron R. Berger, Micheal W. George, and Robert R. Seal II

Chapter R of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–R

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter R

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................R1

Introduction..................................................................................................................................................R1

Uses and Applications ......................................................................................................................R3

Demand and Availability of Supply .................................................................................................R3

Geology.........................................................................................................................................................R3

Geochemistry......................................................................................................................................R3

Mineralogy..........................................................................................................................................R5

Deposit Types .....................................................................................................................................R6

Magmatic Copper–Nickel–Platinum-Group-Metal Sulfide Deposits ...............................R6

Iron Oxide-Copper-Gold Deposits ..........................................................................................R7

Volcanogenic Massive Sulfide Deposits ..............................................................................R7

Porphyry Deposits.....................................................................................................................R7

Skarn Deposits ..........................................................................................................................R8

Epithermal Deposits .................................................................................................................R9

Orogenic Gold Deposits .........................................................................................................R11

Carlin-Type Gold Deposits .....................................................................................................R12

Resources and Production ......................................................................................................................R12

Identified Resources and Reserves ..............................................................................................R12

Undiscovered Resources ...............................................................................................................R14

Exploration for New Deposits .................................................................................................................R14

Environmental Considerations ................................................................................................................R14

Sources and Fate in the Environment ..........................................................................................R15

Mine Waste Characteristics ..........................................................................................................R15

Human Health Concerns .................................................................................................................R16

Ecological Health Concerns ...........................................................................................................R17

Carbon Footprint...............................................................................................................................R17

Mine Closure.....................................................................................................................................R17

Problems and Future Research ..............................................................................................................R18

Acknowledgments....................................................................................................................................R18

References Cited.......................................................................................................................................R18

Suggested citation:

Goldfarb, R.J., Berger, B.R., George, M.W., and Seal, R.R., II, 2017, Tellurium, chap. R of Schulz, K.J., DeYoung, J.H., Jr., 

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. R1– R27, http://dx.doi.org/

10.3133/pp1802R.
Tellurium

By Richard J. Goldfarb, Byron R. Berger,1 Micheal W. George, and Robert R. Seal II

1

Deceased.

Abstract

Tellurium (Te) is a very rare element that averages only 

3 parts per billion in Earth’s upper crust. It shows a close 

association with gold and may be present in orebodies of most 

gold deposit types at levels of tens to hundreds of parts per 

million. In large-tonnage mineral deposits, such as porphyry 

copper and sea-floor volcanogenic massive sulfide deposits, 

sulfide minerals may contain hundreds of parts per million 

tellurium, although the orebodies likely have overall concentrations of 0.1 to 1.0 parts per million tellurium. Tellurium is 

presently recovered as a primary ore from only two districts 

in the world; these are the gold-tellurium epithermal vein 

deposits located adjacent to one another at Dashuigou and 

Majiagou (Sichuan Province) in southwestern China, and 

the epithermal-like mineralization at the Kankberg deposit in 

the Skellefteå VMS district of Västerbotten County, Sweden. 

Combined, these two groups of deposits account for about 

15 percent (about 70 metric tons) of the annual global 

production of 450 to 470 metric tons of tellurium. Most of the 

world’s tellurium, however, is produced as a byproduct of the 

mining of porphyry copper deposits. These deposits typically 

yield concentrations of 1 to 4 percent tellurium in the anode 

slimes recovered during copper refining. Present production 

of tellurium from the United States is solely from the anode 

slimes at ASARCO LLC’s copper refinery in Amarillo, Texas, 

and may total about 50 metric tons per year. The main uses of 

tellurium are in photovoltaic solar cells and as an additive to 

copper, lead, and steel alloys in various types of machinery. 

The environmental data available regarding the mining of 

tellurium are limited; most concerns to date have focused 

on the more-abundant metals present in the large-tonnage 

deposits from which tellurium is recovered as a byproduct. 

Global reserves of tellurium are estimated to be 24,000 metric 

tons, based on the amount of tellurium likely contained in 

global copper reserves and on a 50 percent recovery rate from 

refinery anode slimes during the commonly used electrolytic 

process, also known as solvent extraction-electrolytic refining. 

If the more economical solvent-leach process—a process that 

does not recover tellurium—is increasingly used in the future 

to recover lower grades of copper from porphyry and other 

large-tonnage deposits, then additional high-grade telluriumrich gold deposits may become new primary sources for 

tellurium, particularly epithermal vein deposits associated with 

alkaline magmatism.

Introduction

Tellurium (Te) is an element for which there is little 

reliable information about specific deposits and (or) primary 

production sources. Two vein deposits in southwestern China 

that are adjacent to one another and a possible deformed 

ancient sea-floor volcanogenic massive sulfide (VMS) deposit 

in Sweden may represent the only two locations in the world 

where tellurium is a primary resource (fig. R1).

Most tellurium is recovered as a byproduct in the anode 

slimes from the electrorefining of copper and other polymetallic sulfide ores, which are typically mined from porphyry, 

VMS, magmatic nickel–copper– platinum-group-metal (PGM), 

and (or) skarn deposits The tellurium received for processing 

at a refinery may have come from several different types of 

mineral deposits within a given country and (or) from another 

country or countries. If the tellurium is to be recovered as 

a final tellurium metal product at a refinery, it commonly 

requires upgrading to tellurium metal at a second industrial 

facility, which may be located in the same or a different 

country. In the United States, for example, all tellurium 

production is from a single refinery in Texas, but it is impossible to identify the specific deposits that have contributed the 

ore from which the tellurium is produced, although the sources 

are likely some of the large porphyry copper deposits scattered 

across the southwestern and western parts of the country.

The refined anode slimes from which tellurium is extracted

are typically dominated by copper and silver, although most 

global refineries report contents of 1 to 4 percent tellurium 

in the anode slimes, and some copper refineries in central 

Asia and Russia report as much as 8 to 9 percent tellurium 

(Moats and others, 2007). A minor part of the world’s 

tellurium is also recovered from the dust and gases produced 

during the smelting of sulfide-rich ores and from the skimmings (when a distillation process is used) at lead refineries.
R2 Critical Mineral Resources of the United States—

Tellurium

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Tellurium, Fig 1.

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Commoner

Bushveld

Great Dyke

Rodalquilar

Bingham Canyon

Joplin

Ducktown

Elatsite

Skouriés

MetaliferiMountains

Cyprus

Porgera

Mt. Kare

Golden Mile

Olympic Dam

Stormont

Baguio

Lepanto

Lihar

Ertsberg

Kankberg Kola Peninsula

Ortosa

Kumtor Almalyk

Kochbulak

Sandaowanzi

Geodo

Emperor

Stillwater

Golden Sunlight

Black Hills

Kensington

Willow

Creek

Pebble

Sapo Alegre

Pueblo Viejo

El Indio

Ashanti

Dashuigou/Majiagou

Zarshuran

Sentyabr'sky

Sukhoi Log

Ural Mountains

Bereznyakovskoe

Leadville

Ortiz Mountains

Cripple Creek

Telluride

LaPlata

La Bambolla

Hedley

Zortman-Landusky

Mount Milligan

Deer Horn

Lone Pine

Getchell, Meikle

Carlin, Deep Star

Fortitude

Grass Valley

Mother Lode

Bodie Hills

Jefferson

Red Lake

Kalamazoo

Patagonia

Mountains

Ely

Noril’sk

VMSIOCGOrogenicPorphyryEpithermalSkarnCarlinMagmatic 

Cu-Ni-PGM 

Other

Deposit type

EXPLANATION

Figure R1. World map showing locations of selected tellurium-enriched mineral occurrences discussed in the text, by deposit type. Tellurium is produced as a primary commodity from the Kankberg deposit located in Västerbotten County, Sweden, and the Dashuigou and Majiagou deposits located in Sichuan and Shaanxi Provinces, respectively, in China. The deposit types shown here and discussed in the text are classified as follows: volcanogenic massive sulfide (VMS), iron oxide-copper-gold (IOCG), orogenic gold, porphyry, epithermal, skarn, Carlin-type gold, magmatic copper–nickel–platinum-group metal (Cu-Ni-PGM), and other deposits. 
Geology R3

Uses and Applications

The main use of tellurium is for development of 

cadmium-tellurium film in photovoltaic solar cells, which is 

considered to be an emerging application, particularly in the 

growing economies of China and India. In addition, tellurium 

is used as an additive to copper, lead, and steel alloys to 

improve machine efficiency, particularly in thermoelectric 

cooling products. Together, the photovoltaic solar and thermoelectric uses account for more than two-thirds of the world’s 

tellurium usage (fig. R2; Anderson, 2015). About 5 percent 

of tellurium usage is for improving heat resistance during 

vulcanization of rubber. Tellurium is also used as an alloy 

with selenium in photoreceptors in copying machines and as a 

coloring agent in ceramics and glass. It is, in addition, applied 

in integrated circuits, laser diodes, and medical instrumentation; all these uses have had robust growth in recent years.

Demand and Availability of Supply

The price of tellurium has fluctuated significantly since 

2010. According to the U.S. Department of Energy, the supply 

of tellurium is not critical, although it is projected to be “nearcritical” through 2025 (U.S. Department of Energy, 2010). 

Significant amounts of tellurium are available globally from 

the mining of major base-metal-rich deposits. Enrichments in 

precious-metal deposits are well known, although these deposits 

have generally not produced significant amounts of tellurium.

The solar energy industry represents the largest share of 

the tellurium market today, accounting for nearly 40 percent of 

annual tellurium consumption. Some consumption percentages 

have changed since 2010, as metallurgy use has decreased 

to about 15 percent from approximately 30 percent of total 

consumption, and thermoelectric products presently account 

for about 30 percent (fig. R2). Because of the demand for solar 

cells, an estimated 30 percent of the world’s produced tellurium 

is used by First Solar Inc., which in 2016 continued to be the 

world’s leading producer of cadmium-tellurium solar panels.

Geology

Geochemistry

Owing to its high volatility in the hydride phase and 

probable evaporative loss during Earth’s early low-oxygen 

period, tellurium is an exceptionally rare chalcophile element 

in Earth’s upper crust. Tellurium is in the family of Group 6A 

elements in the periodic table, which also includes oxygen, 

sulfur, and selenium, and its chemical behavior is similar to 

that of selenium and sulfur. Native tellurium typically appears 

as a brittle, silver-white substance. It has an atomic number 

of 52 and an atomic weight of 127.6. Tellurium combines 

with elements that are incapable of reducing their volume 

upon losing valence electrons (Sindeeva, 1964). The energy 

required to detach tellurium’s first electron is much higher 

than for the alkaline and alkaline earth metals, and, therefore, 

tellurium does not tend to form ionic bonds; rather, it forms 

covalent and metallic bonds. Tellurium tends to combine with 

metals of high atomic number and similar atomic radii; for 

example, it typically forms minerals that also contain gold, 

silver, and platinum.

McDonough (2003) estimated an overall bulk Earth 

abundance of 300 parts per billion (ppb) tellurium, but 

96 percent of Earth’s tellurium remains in the core, which 

averages 885 ppb tellurium. Earth’s mantle may have an 

average concentration of about 8 ppb tellurium, although 

local concentrations in peridotite can be twice that amount 

because residual sulfide phases in these rocks are commonly 

enriched in tellurium (Palme and O’Neil, 2003). Tellurium 

is preferentially retained in refractory sulfides in harzburgite 

in the mantle during melting processes (Hattori and others, 

2002). Crustal tellurium values are extremely low, and Li 

and Schoonmaker (2003) estimate an average upper crustal 

abundance value of 3 ppb tellurium, which is about eight times 

less than that for gold and makes tellurium one of the least 

common elements in crustal rocks. Only a limited amount 

of abundance data for tellurium is available, and, in fact, 

tellurium is not listed in many modern-day comprehensive 

compilations of global element abundances (for example, 

table 2 in Rudnick and Gao, 2003). In marine shale, however, 

Li and Schoonmaker (2003) indicate much higher average 

concentrations of 70 ppb tellurium, which likely reflects the 

abundance of disseminated oceanic pyrite that has elevated 

levels of tellurium. Hein and others (2013) note mean values 

of 30 to 60 parts per million (ppm) tellurium in ironmanganese crusts formed on the floor of several oceans owing 

to the oxidation of tellurite (Te(IV)) to tellurate (Te(VI)).

Figure R2. Pie chart showing major end uses of tellurium as a 

percentage of world consumption in 2010. Compiled using data 

from Selenium Tellurium Development Association, Inc. (2010) and 

Anderson (2015).

Rubber

formulation

5% Other

10% 

Metallurgy

15% 

Solar cells

40% 

Thermoelectric

products

30% 
R4 Critical Mineral Resources of the United States—Tellurium

The solubility and transport of tellurium in magmatic and 

hydrothermal fluids are incompletely understood, particularly 

at the temperatures and pressures of igneous melt reservoirs 

and in high-grade metamorphic terranes. Taken together with 

the intrinsic low abundance of the element, the heterogeneity 

of the concentrations of tellurium within any mineral-deposit 

type and the very fine granularity of telluride minerals limit 

the ability to predict tellurium resources that might occur in 

a specific type of mineral deposit. Tellurium has a relatively 

low solubility in aqueous solutions; fields for the dominant 

aqueous tellurium species are shown as a function of pH and 

oxygen fugacity in figure R3. Given the high abundance of 

tellurium in some deposits, the low solubility suggests that 

tellurium may be transported predominantly in the vapor 

phase (for example, Cooke and McPhail, 2001), probably 

as solvated molecular species. Tellurium occurs in volcanic 

fumaroles, such as the high-tellurium amorphous arsenic-rich 

sulfurite at Vulcano, Sicilia, Italy (Garavelli and others, 2013) 

and in volcanic condensate sulfur from the Mendeleev volcano 

in the Kuril Islands, Sakhalinskaya Oblast’, Russia, and the 

Mutnovskaya volcano on Russia’s Kamchatka Peninsula 

(Sindeeva, 1964). Volatile transport of tellurium could also be 

enhanced owing to the suggested relatively high anhydrous 

volatility of the element, along with that of arsenic, mercury, 

and selenium (for example, Saunders and Brueseke, 2012), 

where Hg>As>Se>Te>Tl>>Sb>Pb>Ag>Cu>Au. In addition, the common enrichment of tellurium in nonmagmatic 

orogenic gold systems may indicate that carbon dioxide (CO2

) 

immiscibility also is an important factor leading to the gaseous 

transport and concentration of tellurium.

Although tellurium varies in oxidation state from 

–2 (telluride) to +4 (tellurite) to +6 (tellurate), tellurium is 

most soluble as tellurite in hydrothermal fluids. Hydrothermal 

waters in the areas of epithermal gold deposits may contain 

as much as 4 ppb tellurium, and some sea-floor hot springs 

have as much as 18 ppb tellurium (Grundler and others, 

2013). In typical reduced hydrothermal fluids (for example, 

log f  O2 <–30), Te2

2– is the dominant tellurium-bearing ion 

(Brugger and others, 2012), although under slightly more 

acidic or very reduced conditions HTe–

 is dominant and 

H2TeO3 could be important in very acidic and less reduced 

solutions (fig. R3; McPhail, 1995; Grundler and others, 2013). 

The Te2

2– activity is controlled in aqueous solutions by the 

solubility of native tellurium, which is commonly deposited 

along with many telluride phases. The tellurate species are 

dominant under only very oxidizing conditions, such as in 

near-surface weathering of telluride-rich ores (fig. R3).

Some researchers have suggested that the common 

occurrence of gold-bearing telluride minerals implies the 

transport of gold as a soluble tellurium complex (for example, 

Jensen and Barton, 2000; Cooke and McPhail, 2001; Brugger 

and others, 2012), although most experimental evidence at 

magmatic conditions indicates the transport of gold either 

as a sulfide complex (Loucks and Mavrogenes, 1999) or as 

both chloride and sulfide complexes (Frank and others, 2011). 

Hydrothermal fluids with anomalous tellurium concentrations, 

based on the abundance of telluride minerals, include both 

CO2-rich fluids that form orogenic gold and alkalic-related 

epithermal gold deposits, and CO2-poor fluids that typify most 

low- and high-sulfidation precious-metal-bearing epithermal 

ores. Therefore, it is unlikely that CO2 plays a significant role 

in the transport or concentration of tellurium. Furthermore, 

although typical tellurium enrichments in ore deposits are 

assumed to reflect magmatic-hydrothermal activity, the 

tellurium enrichment in many orogenic gold deposits, which 

are widely accepted to have formed from nonmagmatic 

metamorphic fluids, suggests that such a magmatic association 

is not necessary.

–35

–30

–25

–20

–15

–10

–5

0

pH

Log oxygen fugacity

H2TeO3(aq)

HTeO3

–

TeO3

2 –

Te2–

HTe–

Tellurium

Py

He

Mt

Po

0 2 4 6 8 10 12 14

Te2

2–

H3TeO+

3

Figure R3.

EXPLANATION

Stability field for native tellurium

Stability field for aqueous tellurium species

Boundary of stability field for tellurium species

Boundary of stability field for iron-bearing minerals

Figure R3. Phase diagram showing speciation calculations 

for tellurium in a hydrothermal fluid at 300 degrees Celsius (°C), 

as a function of pH and oxygen fugacity, with the following 

activities: tellurium=10−10, iron=0.1, sulfur=10−4, and chlorine=0.5. 

At an activity of tellurium of 10−10, a typical near-neutral to acidic, 

reduced aqueous fluid cannot carry much tellurium, and native 

tellurium will be stable (shaded area in lower left side of diagram). 

The low solubility of tellurium in aqueous solutions suggests 

that concentration of tellurium in a vapor may be important for 

mineral formation unless the fluid has a very high concentration of 

tellurium (that is, at an activity of Te=10−8 or greater). The oxygenbearing tellurate species, although with large stability fields, 

are only significant under very oxidizing conditions above a log 

oxygen fugacity of about 10–29 for any pH value. After Grundler 

and others (2013). aq, aqueous; He, hematite; Mt, magnetite; 

Po, pyrrhotite; Py, pyrite; and Te, tellurium
Geology R5

Studies of both liquid and vapor fluid inclusions from 

porphyry and epithermal deposits have indicated concentrations of hundreds of parts per million tellurium in the fluids 

that formed some deposits (Wallier and others, 2006; Pudack 

and others, 2009). For example, fluid inclusions in highsulfidation epithermal veins in Romania contained as much as 

3 weight percent tellurium, which appeared to correlate with 

enrichments in dissolved silver and gold (Kouzmanov and 

others, 2010). Grundler and others (2009, 2013) concluded 

that these unusually elevated tellurium concentrations indicate 

that aqueous tellurite was transported in a highly oxidized 

magmatic fluid and (or) in a fluid with basic pH (approximately greater than pH 8), with tellurite reduction by fluidrock interaction or by fluid mixing that caused precipitation 

of telluride minerals. Available information suggests that 

any drastic change in fluid conditions, including a decrease 

in temperature, wallrock sulfidation, phase separation, pH 

shifts, and (or) redox change, can destabilize telluriumbearing complexes (for example, Cooke and McPhail, 2001; 

Ciobanu and others, 2006; Maslennikov and others, 2013). 

Most hypogene telluride minerals probably were deposited 

at temperatures of between 300 and 100 degrees Celsius (oC) 

(for example, Zhang and Spry, 1994; Shackleton and others, 

2003), although in some magmatic copper-nickel-PGM 

deposits, palladium- and (or) lead-bearing tellurides are stable 

at temperatures of 400 o

C (Vymazalová and Drábek, 2011).

Mineralogy

The most common tellurium-bearing minerals are listed 

in table R1. These minerals occur in many different types of 

ore deposits, but the minerals themselves are typically too low 

in abundance to be sources for recoverable tellurium ore.

Table R1. Tellurium-bearing minerals, many of which are mentioned in this chapter.

Mineral name Chemical formula

Altaite PbTe

Benleonardite Ag8

(Sb,As)Te2S3

Bezsmertnovite Au4Cu(Te,Pb) 

Bilibinskite PbCuAu3

Te2

Blakeite (Fe2(TeO3)3

Bogdanovite Au,Te,Pb)3(Cu,Fe)

Buckhornite AuPb2

BiTe2S3

Calaverite AuTe2

Cervelleite Ag4TeS

Csiklovaite Bi2

TeS2

Coloradoite HgTe

Emmonsite Fe2

(TeO3)3 • 2H2O

Empressite Ag5–xTe3

Frohbergite FeTe2

Goldfieldite Cu12(Te, Sb, As)4S13

Grunlingite Bi4TeS3

Hedleyite Bi7Te3

Henryite Cu4Ag3Te4

Hessite Ag2Te

Imgreite NiTe

Joseite Bi4

(Te, S)3

Kawzulite Bi2(Te, Se, S)3

Keithconnite PdTe3

Kostovite CuAuTe4

Kotulskite PdTe 

Krennerite Au4

AgTe10

Mackayite Fe2

(TeO3)3nH2O

Mineral name Chemical formula

Maslovite PtBiTe 

Melonite NiTe2

Merenskyite PdTe2

Michenerite PdBiTe 

Moncheite PtTe2

Montainite Bi3

TeO4(OH)4

Montbrayite (Au,Sb)2Te3

Museumite [Pb2(Pb,Sb)2S8][(Te,Au)2]

Muthmannite (AuAg)Te

Nagyágite Pb5

Au(Te,Sb)4S5–8

Native tellurium Te

Pašavaite Pd3

Pb2Te2

Petzite Ag3AuTe2

Poughite Fe2(TeO3)2SO4 • 3H2O

Rickardite Cu4 –xTe2

Rodalquilarite H3Fe2(TeO3)4Cl

Rucklidgeite (Bi, Pb)3

Te4

Stutzite Ag7Te4

Sylvanite AuAgTe4

Teineite Cu(Te, S)O4 • 2H2O

Tellurantimony Sb2

Te3

Tellurobismuthite Bi2Te3

Telluropalladinite Pd9Te4

Tetradymite Bi2Te2S

Tsumoite Bi4

Te2

Volynskite AgBiTe2

Weissite Cu2–xTe
R6 Critical Mineral Resources of the United States—Tellurium

Tellurium is most strongly concentrated in telluride 

minerals, although some tellurium is also concentrated in 

sulfide minerals, particularly in sulfosalts, such as the mineral 

goldfieldite. In the gold-silver-tellurium system alone, 

11 different tellurium-bearing phases have been identified 

(Zhang and Spry, 1994). Grundler and others (2013) indicate 

that calaverite typically coexists with native gold or the silverbearing tellurides krennerite and petzite, whereas native tellurium is more commonly associated with other silver-bearing 

tellurides, such as hessite, stutzite, and sylvanite. Where the 

ratios of tellurium to gold plus silver are high, bismuth-bearing 

tellurides, coloradoite, or goldfieldite have been shown to be 

the dominant tellurium-bearing phases (Grundler and others, 

2013). There are numerous palladium-bearing tellurides in 

the lead-palladium-tellurium system, including keithconnite, 

kotulskite, merenskyite, pašavaite, and telluropalladinite, and 

these are sometimes common in magmatic copper-nickelPGM deposits (Vymazalová and Drábek, 2011). Moncheite is 

a common platinum host in the deposit type.

In most ore deposits, telluride minerals consistently are 

deposited in the later stages of paragenetic sequences; that 

is, they form after the deposition of many iron- and basemetal-bearing sulfides. At the tellurium-rich Cripple Creek 

district gold deposits in Teller County, Colorado, Lovering and 

Goddard (1950) and Saunders (1986) documented a relative 

chronology of telluride mineral deposition. Generally, with 

time, telluride composition shifts from gold-rich to silver-rich 

during these late hydrothermal events (for example, Dye 

and Smyth, 2012). There are exceptions, however, such as 

at the Tínos Island precious-metal deposit in Notío Aigaío, 

Greece, where a silver-to-copper-to-gold paragenetic sequence 

characterizes telluride deposition (Tombros and others, 2010).

Native tellurium is reported from many precious-metal 

deposits, including numerous gold deposits in Japan. Nakata 

and Komuro (2011) indicate that quartz, quartz-goldfieldite, 

quartz-stutzite-hessite-sylvanite-tetradymite, and quartzhessite-sylvanite-tellurantimony assemblages all contain 

native tellurium.

On the active sea floor along the Kermadec-Tonga trench 

in the Pacific Ocean, telluride minerals and native tellurium 

have been identified in massive chalcopyrite occurrences and 

in areas between sulfide chimneys of chalcopyrite and pyrite 

(Berkenbosch and others, 2012). Most of the telluride minerals 

are less than 5 micrometers in diameter, but some reach 

lengths of 80 micrometers. These are dominantly gold-rich 

(calaverite) and bismuth-rich (kawazulite, tellurobismuthite, 

and tetradymite) telluride minerals.

Oxidation of gold-rich tellurides can form tellurite and 

tellurate phases within the weathering zone. It can also lead 

to the formation of low reflectivity “mustard gold” through 

dissolution of calaverite and other gold-rich tellurides and 

subsequent reprecipitation processes (Zhao and others, 2009). 

In addition, gold pseudomorphs after tellurides may form. In 

the weathering environment, tellurium most likely undergoes 

sorption onto, and thus concentration within, clay-sized soil 

particles (Hayes and others, 2012)

Deposit Types

Tellurium is present in many different types of mineral 

deposits. The diversity of mineral deposit types in which it 

occurs may reflect enrichment of the element in the structure 

of sulfide grains or the formation of discrete tellurium-bearing 

minerals. The deposit types listed here are only those for 

which there is, has been, or is projected to be some economic 

interest in their tellurium content.

Magmatic Copper–Nickel–Platinum-GroupMetal Sulfide Deposits

Magmatic sulfide deposits are mafic and ultramafic 

igneous bodies that are enriched in copper, nickel, and (or) 

platinum-group metals (PGMs) (Naldrett, 2011). Little or 

no published information is available regarding tellurium 

production from and the concentration of tellurium in studied 

deposits of this type, but it is highly likely that these ores 

contribute to global tellurium production, as the tellurium is 

produced as a byproduct of the milling of the major mineral 

commodities. For example, some amount of tellurium is 

recovered in Russia from magmatic sulfide deposits. Telluride 

minerals, which contain the greatest concentrations of 

tellurium of any minerals, are present in many deposits of 

this type, but only in subeconomic quantities.

At the Sudbury district in Ontario, Canada, where Paleoproterozoic ore-related magmas were produced by shock waves 

from a meteor impact, hydrothermal activity in the footwall 

of the emplaced igneous complex has enriched the rocks with 

gold and PGMs. Much of the PGM mineralization is hosted in 

tellurium-rich phases that include kotulskite, merenskyite, and 

monocheite. In addition, bismuth-, mercury-, and silver-rich 

telluride phases have been noted (Péntek and others, 2013).

At Noril’sk in Krasnoyarskiy Kray, Russia, the magmatic 

copper-nickel-PGM deposits associated with the Triassic 

continental flood basalts have yielded minor amounts of 

byproduct tellurium (in the range of 5 metric tons per year) 

(Safirova, 2012). Tellurium of potential economic interest 

at Noril’sk is confined to sulfide minerals (Sindeeva, 1964). 

Chalcopyrite, for example, contains from 2 to 72 ppm 

tellurium; pyrrhotite, from 4 to 45 ppm; pentlandite, 13 ppm; 

and bornite, from 10 to 15 ppm tellurium (Sindeeva, 1964). 

Massive vein-type ores of copper contain the highest concentrations of tellurium, whereas disseminated ores contain lesser 

amounts of tellurium (from trace amounts to 12 ppm).

Some of the platinum and palladium resource in the 

Paleoproterozoic Merensky Reef in the eastern Bushveld 

Complex of Limpopo Province, South Africa, is contained 

in michenerite and maslovite, which are PGM-bearing 

bismuthotellurides that were suggested to have formed by 

remobilization of magmatic PGMs during an overprinting 

hydrothermal event (Rose and others, 2011). The minerals 

that host palladium and platinum most commonly in the Pana 

and Fedorov massifs of the Kola Peninsula, Russia, and the 
Geology R7

Great Dyke, Zimbabwe, are arsenides and bismuthotellurides 

(Oberthür, 2002; Mitrofanov and others, 2005). In the Late 

Archean Stillwater Complex in Montana, mineralized samples 

with as much as 182 ppm palladium and 52 ppm platinum 

also contain as much as 17 ppm tellurium and significant 

amounts of arsenic and bismuth, reflecting a complex telluride 

mineralogy (Zientek and others, 1990).

Iron Oxide-Copper-Gold Deposits

Iron oxide-copper-gold (IOCG) deposits represent a 

poorly defined and understood group of deposits of mainly 

Precambrian age (Groves and others, 2010). They are generally defined as sulfur-poor magmatic-hydrothermal deposits 

with large resources of copper, gold, iron, and (or) uranium in 

extensional cratonic settings. There are few data on tellurium 

in IOCG deposits. The anode slime at the refinery of the giant 

Mesoproterozoic Olympic Dam deposit in South Australia, 

Australia, which is the world’s largest identified IOCG deposit, 

averages about 3.5 percent tellurium (Moats and others, 2007) 

and is a source of some tellurium production. This indicates that 

the tellurium produced from Olympic Dam is recovered from 

the bornite, chalcocite, and chalcopyrite ore minerals. Discrete 

telluride phases at Olympic Dam have not been reported.

Volcanogenic Massive Sulfide Deposits

The volcanogenic massive sulfide (VMS) deposits consist 

of strata-bound sulfide minerals that formed at or near the 

sea floor in association with contemporaneous volcanism. 

The leading producers of tellurium in Russia are the VMS 

deposits in the Ural Mountains, which yield about 35 metric 

tons per year of tellurium (Safirova, 2012). Fine-grained 

telluride minerals are abundant in the Silurian and Devonian 

copper-zinc VMS deposits of the Urals (Vikentyev, 2006). 

Maslennikov and others (2013) described a variety of 

Ag-Pb-Bi-Te and Cu-Ag-Te-S solid-solution series phases and 

tellurium-rich sulfosalts in ancient sea-floor chimney structures. They suggest, however, that chalcopyrite is the most 

significant host for tellurium. In the Sibay VMS deposit in the 

Urals region of Bashkortostan Republic, Russia, tellurium is 

predominantly associated with copper ores (which average 11 

ppm tellurium), whereas zinc-rich ores less frequently have 

detectable levels of tellurium (Sindeeva, 1964). In the Uchaly 

deposit, which is also located in the Urals region of Bashkortostan Republic, pyrite and pyrite-copper ores contain from 22 

to 190 ppm tellurium (Sindeeva, 1964). No production data 

are available for individual deposits; tellurium is recovered 

from electrolytic refining of the copper-rich VMS ores, but 

the discrete telluride phases are unlikely to be important 

contributors.

Some VMS deposits are transitional with telluride-bearing 

epithermal precious-metal mineral occurrences, including 

those as old as Archean (for example, Côté-Mantha and 

others, 2012). The poorly understood Kankberg deposit in the 

Paleoproterozoic (circa 1,890 mega-annum [Ma]) Skellefteå 

district in Västerbotten County, Sweden, was originally mined 

as a copper-zinc deposit into the 1990s, and it was classified by 

some workers as a VMS deposit. The deposit has been reopened 

as a gold-tellurium mine that is projected to produce approximately 10 percent of the world’s tellurium, or about 41 metric

tons per year, from 2012 to 2020 (Boliden Group, 2011). Thus, 

Kankberg represents one of the few presently active mining 

operations in which tellurium is a primary product and telluride 

minerals are the source of the tellurium. The ore averages 

approximately 4.1 grams per metric ton (g/t) gold and 186 g/t 

tellurium. The sea-floor deposit is thought to be rich in tellurium 

and is possibly a hybrid VMS-epithermal system that formed 

on a shallow part of the ocean floor. If this is correct, it would 

indicate that significant amounts of tellurium are available 

from sea-floor deposits and are not solely associated with the 

more classic subaerial epithermal ores described below (for 

example, Mercier-Langevin and others, 2013). Numerous VMS 

deposits—both ancient ones exposed on continents and modern 

ones presently forming on the sea floor—may thus be associated with an epithermal suite of trace elements (for example, 

Hannington and others, 2005; Berkenbosch and others, 2012) 

and potentially could be highly enriched in tellurium.

Porphyry Deposits

Tellurium is enriched in sulfide and (or) telluride 

phases in most copper porphyry deposits and thus may be 

a recoverable byproduct during milling of some of these 

copper ores. Porphyry deposits are shallowly formed 

magmatic-hydrothermal deposits typically developed in the 

cupolas to porphyritic intrusions within active continental 

margins above subduction zones. The deposits are generally 

characterized by stockwork veinlets rich in copper- and 

(or) molybdenum-bearing minerals. The principal source of 

tellurium in the United States and the world is as a byproduct 

of copper refining of ores from large-tonnage, low-grade 

copper and copper-gold porphyry deposits. For example, the 

giant Carboniferous copper porphyry deposits of the Almalyk 

district in Toshkent Province, Uzbekistan, are a source of 

recovered tellurium; the deposits were estimated to contain 

1,098 metric tons of tellurium in 1996 (Levine, 2011, citing 

report from United States Trade and Development Agency, 

1996). In the Almalyk district, average concentrations in 

ore were 16 ppm tellurium at the Kal’makyr deposit and 

7 to 8 ppm tellurium at the Dalnee deposit. In the world-class 

Cretaceous Pebble copper-gold-molybdenum deposit in 

Alaska, calaverite and petzite have been observed to contain 

some of the gold. Gregory and others (2013) estimate that 

2.5 to 3.0 percent of the gold in the deposit is hosted by the 

two telluride phases. These phases occur as inclusions in chalcopyrite, which is the main copper ore mineral at the Pebble 

deposit. The majority of what is conjectured to be an enormous tellurium endowment at the Pebble deposit is potentially 

hosted in the structure of chalcopyrite and pyrite, however, 

assuming these phases contain tellurium concentrations similar 
R8 Critical Mineral Resources of the United States—Tellurium

to sulfide minerals in many porphyry deposits, which provide 

much of the byproduct tellurium produced in the United States 

and the world. It is not clear whether tellurium recovery is 

planned if and when the Pebble deposit goes into production. 

Palladium is enriched in many porphyry copper-gold deposits, 

reflecting the presence of palladium-bearing telluride minerals 

as distinct grains or as inclusions in bornite and chalcopyrite. 

For example, merenskyite is a common tellurium-rich phase 

that is associated with lesser amounts of hessite and kotulskite 

in many porphyry deposits (Economou-Eliopoulos, 2005).

Eilers (1914) reported that the Tertiary Bingham Canyon 

deposit in Utah contained greater than 23 to 27 ppm tellurium 

in blister copper (the almost pure copper produced during 

refining), although tellurium was not detected in the partially 

refined copper from the porphyry deposits at Ely, Nevada. 

High-grade copper-molybdenum-gold ore from Bingham 

Canyon has been estimated to average 4.8 ppm tellurium 

(Austen and Ballantyne, 2011). Economou-Eliopoulos and 

Eliopoulos (2000) reported whole-rock concentrations of 

tellurium ranging from 0.33 to 2.7 ppm in mineralized samples 

from the Skouriés deposit and the adjacent Fissoka prospect 

in the Peloponnísos region of Greece. Tokmakchieva (2002) 

and Tarkian and others (2003) reported as much as 106 ppm 

tellurium in ore samples from the Elatsite deposit in Sofiya, 

Bulgaria. Economou-Eliopoulos and Eliopoulos (2000) also 

estimated concentrations of 4.0 and 18.5 ppm tellurium in two 

samples of chalcopyrite from the Skouriés deposit. Reported 

differences in concentrations from the deposits likely reflect 

sampling variability; higher concentrations result from 

analyses of samples of mainly copper-bearing sulfide minerals 

that may contain hundreds of parts per million tellurium, 

whereas samples of sulfide minerals in wallrock would have 

much lower concentrations.

Although the main copper±gold ores in porphyry 

deposits at Ely, White Pine County, Nevada, are not enriched 

in tellurium, mineralized rock in the silver-rich halo is highly 

anomalous in tellurium, averaging about 100 ppm tellurium 

(Gott and McCarthy, 1966). Watterson and others (1977) also 

found high concentrations of tellurium in silicified rocks in a 

halo surrounding the copper ores. Chaffee (1982) identified two 

zones with anomalous whole-rock tellurium at the Kalamazoo 

deposit in Pinal County, Arizona. The highest concentrations of 

tellurium were present in the outer pyritic halo of the porphyry 

deposit, whereas lower anomaly values corresponded with 

the copper-gold ore zone. Cox and others (1975) identified a 

similar zonation at the Sapo Alegre deposit in Puerto Rico. 

Therefore, although much of the present tellurium production 

is a byproduct from copper production from porphyry copper 

ores, future primary resources may be considered in the alteration halos of many porphyry deposits if tellurium demand 

increases. Because tellurium is relatively volatile compared 

with many other elements, it is not surprising that there are 

high tellurium concentrations in alteration zones surrounding 

the main bulk-minable copper resources.

Many of the porphyry systems with abundant telluride 

minerals are specifically associated with subduction-related 

alkalic porphyritic intrusions. These include those 

of the Late Triassic-Early Jurassic continental arc in 

British Columbia, Canada. The Mount Milligan copper-gold 

deposit, which is one of the larger deposits of the group, is 

characterized by a late-stage “subepithermal” or intermediatesulfidation overprinting event that deposited an Au-PGM-AsSb-Bi-Te-Hg assemblage, perhaps by way of vapor transport 

(LeFort and others, 2011). Many gold-rich epithermal 

deposits, particularly those related to alkalic igneous rocks, 

contain significant tellurium concentrations and may reflect 

late-stage events in an evolving porphyry-epithermal 

magmatic-hydrothermal system. The La Plata deposit in 

La Plata County, Colo., may be such an example, where goldtelluride vein and replacement deposits surround an alkaline 

intrusion that is rich in gold and copper (Jones, 1992).

Some workers have suggested that epithermal 

gold-silver-tellurium mineralization is closely linked to the 

more molybdenum-rich alkalic porphyry deposits (Rice 

and others, 1985; Saunders, 1991). Spry and others (1996), 

for example, described the Golden Sunlight deposit in 

Jefferson County, Montana, as an epithermal gold deposit that 

developed in alteration surrounding an alkaline, low-grade 

molybdenum-bearing Late Cretaceous porphyry intrusion. 

They described bismuth-, gold-, and silver-rich telluride 

minerals in a breccia pipe that overprints the earlier stockwork 

molybdenum mineralization and represents the later stages of 

a magmatic-hydrothermal system. Whether or not tellurium is 

also concentrated in early-stage molybdenite and is recovered 

as a byproduct from milling of molybdenum porphyry deposits 

is uncertain. In the United States, all tellurium production is 

from ASARCO LLC’s copper refinery in Amarillo, Texas, 

which processes copper ore but not molybdenum ore 

(David A. John, Research Geologist, U.S. Geological Survey, 

written commun., 2013); consequently, porphyry molybdenum 

deposits located in the Rocky Mountains of Colorado are not 

a present-day source of domestic tellurium production.

Skarn Deposits

Copper-bearing skarn deposits that were formed in areas 

of replacement of carbonate rock by garnet, pyroxene, and 

other calc-silicate minerals commonly are associated with 

porphyry copper deposits where carbonate rocks are part of the 

country-rock assemblage. Fuertes-Fuente and others (2000) 

report hedleyite in the Ortosa deposit in Asturias, Spain, and 

Prendergast and others (2005) report altaite, hessite, petzite, 

and tetradymite in the Big Gossan and Wanagon skarn deposits 

in the Ertsberg district of Papua Province, Indonesia. In the 

Darwin district of Inyo County, California, bismuth tellurides, 

rickardite, and tellurium-rich sulfosalts occur (Afifi and others, 

1988). In many gold-rich skarns, bismuth and tellurium are 

both enriched (for example, at the Hedley deposit in British 

Columbia, Canada [Meinert, 2000]; the Fortitude deposit in 

Lander County, Nev. [Meinert, 2000]; the Stormont deposit 

in Tasmania, Australia [Cockerton and Tomkins, 2012]; and 

the Geodo deposit in Gangwon, Republic of Korea [Kim and 
Geology R9

others, 2012]). It is unknown whether skarns contribute to 

any of the world’s tellurium production. If large refineries that 

recover copper from numerous porphyry deposits also handle 

some ores from surrounding skarns, then potentially some 

byproduct tellurium may have been produced from the skarns.

Epithermal Deposits

Epithermal gold and silver deposits may or may not be 

associated with porphyry deposits. They are more shallowly 

formed deposits commonly consisting of veins associated 

with volcanic activity. Tellurium typically is more abundant 

and tellurium-bearing minerals are more diverse in epithermal 

deposits than in porphyry and other types of magmatic 

deposits. Tellurium in epithermal deposits occurs in telluride 

minerals, as native tellurium, and as tellurium-bearing sulfosalts in unoxidized ores, as well as in the form of tellurites in 

secondary ores. Tellurium is enriched in both main subclasses 

of the epithermal type; that is, the so-called high-sulfidation 

and low-sulfidation epithermal varieties (for example, White 

and Hedenquist, 1990). The high-sulfidation subclass, which is 

spatially associated in many instances with porphyry deposits 

(Simmons and others, 2005), typically is characterized by an 

alteration assemblage that includes alunite, kaolinite, quartz, 

and pyrophyllite. The low-sulfidation subclass is typically 

characterized by an alteration assemblage that includes 

adularia, quartz, and sericite (illite). Some epithermal districts 

contain both high-sulfidation and low-sulfidation types of ores.

Low-sulfidation epithermal deposits occur with a wider 

variety of igneous rock compositions than do high-sulfidation 

deposits, but they also occur in some hydrothermal systems 

that lack any obvious magmatic input. Tellurium-bearing 

minerals may constitute economically significant gold ores, 

such as at deposits in the Cripple Creek district of Teller County, 

Colo., or they may be minor accessories, such as at deposits in 

the Telluride district of San Miguel County, Colo. Although not 

in all cases, economically significant tellurium-bearing ores 

occur in epithermal-like veins with a polymetallic character, 

which are sometimes referred to as intermediate-sulfidation 

epithermal deposits; that is, they contain significant amounts 

of copper, lead, and zinc sulfide and sulfosalt minerals, and 

are sometimes considered a transitional deposit type between 

epithermal and porphyry ores (Simmons and others, 2005).

Most epithermal deposits are associated with subductionrelated oceanic or continental magmatic arcs. One subtype 

of the low-sulfidation-type gold-silver epithermal deposit 

is associated with alkaline to subalkaline, light-rare-earthelement-enriched magmatism in back-arc and post-subduction 

settings, and is notably enriched in tellurium (Jensen and 

Barton, 2000; Kelley and Spry, 2016). These alkaline gold 

deposits include those of (a) the Rocky Mountain System 

of the United States (for example, the Zortman-Landusky 

district in Phillips County, Mont.; the northern Black Hills 

of South Dakota; the Cripple Creek district in Teller County, 

Colo.; and the Ortiz Mountains of New Mexico), (b) the 

Himalayan foreland of southwestern China (for example, 

the Dashuigou deposit and the Majiagou deposit [Sichuan 

Province]) (Mao and others, 2002), and (c) Papua New Guinea 

and Fiji (for example, in Papua New Guinea, the Porgera 

deposit in Enga Province, the Mt. Kare deposit in Enga 

Province, and the Lihir deposit in New Ireland Province; and 

in Fiji, the Emperor deposit in Viti Levu). In the Cripple Creek 

district, Gott and others (1967) reported the district-wide 

background as 2.7 ppm tellurium in volcanic rocks, with an 

average value of 10.7 ppm tellurium in areas of anomalous 

gold. In China, the adjacent Dashuigou and Majiagou 

deposits are rare examples in which tellurium is the main ore 

commodity; it is recovered mainly from tetradymite. Many 

deposits in this alkaline subtype are typically characterized by 

a fluorine-gold-tellurium-vanadium magmatic-hydrothermal 

geochemical signature (Jensen and Barton, 2000; Simmons 

and others, 2005).

Saunders and Brueseke (2012) suggested that alkaline, 

tellurium-rich low-sulfidation epithermal deposits of the 

Western United States are preferentially located over 

Precambrian basement rocks east of the initial 87Sr/ 86Sr = 0.706 

isopleth. They show a distinct break in epithermal ore 

geochemistry, as tellurium is more abundant than selenium 

to the east and selenium is more abundant than tellurium to 

the west of the ancient cratonic margin. The tellurium-rich 

epithermal province, which extends from Montana to New 

Mexico, has mainly alkalic magmatism, whereas the seleniumrich ores occur with more typical calc-alkaline rocks. Saunders 

and Brueseke (2012) related the different patterns to differences in the volatility of the metalloids within a subducting 

Laramide slab; less volatile tellurium is released later and at 

greater depths, and thus further inboard than the selenium. 

There are, however, notable exceptions to this pattern, 

including tellurium-rich low- and high-sulfidation gold ores 

at Bodie in eastern California (David A. John, Research 

Geologist, U.S. Geological Survey, written commun., 2013).

About 5 metric tons of tellurium were produced in the 

early 1960s from the Lone Pine deposit in the Wilcox district 

in Catron County, New Mexico, where fluorite, gold, and 

silver have also been recovered in minor amounts (Lueth and 

others, 1996). Low-sulfidation ores are hosted by Oligocene 

rhyolitic intrusions in a caldera ring complex. Telluride 

minerals (krennerite?) and native tellurium are reported. The 

weathering of pyrite and native tellurium led to formation of 

many different iron tellurate minerals. The original hypogene 

telluride mineralization was best developed at the pyritefluorite transition, where the deposit was upwardly zoned from 

pyrite to pyrite-tellurides to fluorite (Lueth and others, 1996).

Pals and Spry (2003) report that telluride minerals associated with alkaline magmatism account for 10 to 50 percent of 

the gold in the Emperor deposit at Vatukoula in Fiji’s Western 

Division. Sylvanite is the main precious-metal-hosting phase, 

although the epithermal ores also include benleonardite, 

coloradoite, and native tellurium. Grab samples of these ores 

contain as much as 345 ppm tellurium, which correlates with 

very high concentrations of fluorine, gold, selenium, and silver 

(Kwak, 1990).
R10 Critical Mineral Resources of the United States—Tellurium

The Dashuigou deposit in the alkalic magmatic province 

of southwestern China (Sichuan Province) hosts a reported 

9,400-metric-ton indicated resource averaging 1.40 percent 

tellurium and a 20,800-metric-ton inferred resource averaging 

0.51 percent tellurium, for a total resource of 240 metric 

tons of tellurium. The indicated and inferred resources at the 

similar nearby Majiagou deposit in Sichuan Province contain 

an estimated 272 metric tons of tellurium; the average grade 

of the resources is about 2 percent tellurium. The ores contain 

even larger bismuth resources, reflecting the composition 

of the tetradymite (Bi2Te2S)—59.27 percent bismuth, 

36.19 percent tellurium, and 4.55 percent sulfur, by weight 

(Behre Dolbear Asia Inc., 2009). The tellurium resource 

occurs in low-sulfidation epithermal veins with highly variable 

grades that locally reach 8 percent tellurium (Behre Dolbear 

Asia Inc., 2009). Many parts of the orebodies consist of 

massive pyrrhotite, which were previously mined for iron. The 

ores also contain anomalous gold, which is interpreted to have 

been deposited after both the tellurium-bearing phases and 

native tellurium (Mao and others, 2002). The current annual 

production of these two deposits is unknown, but Anderson 

(2015) estimated it to be between 10 and 30 metric tons per 

year of tellurium.

Cook and Ciobanu (2005) importantly stressed that 

although there is definitely a common genetic link between 

tellurium-bearing gold deposits and alkalic magmatism, many 

low-sulfidation epithermal systems with significant volumes of 

telluride minerals are also associated with normal calc-alkaline 

magmatism. For example, calc-alkaline-related telluride 

ores are diverse and were historically economic across a 

600-square-kilometer area in the Munţii Metaliferi (Metaliferi 

Mountains) in western Romania, which is one of Earth’s 

richest epithermal ore regions (Alderton and Fallick, 2000). 

The western Romania telluride minerals include hedleyite, 

rucklidgeite, and tsumoite, which are not common in most 

tellurium-rich epithermal deposits and tend to occur in the 

shallowest parts of the deposits (Cook and Ciobanu, 2004). 

Over several centuries, until 1941, the Sacarimb deposit, 

which is located in Hunedoara County in the Metaliferi 

Mountains, is estimated to have cumulatively produced 

30 metric tons of gold, 85 metric tons of silver, and 60 metric 

tons of tellurium from ores that averaged 20 g/t tellurium and 

10 g/t gold. Ore minerals were mainly nagyagite and sylvanite 

that were deposited in Neogene calc-alkaline volcanic rocks 

(Popescu and Neacsu, 2004). This deposit is also the type 

locality for krennerite, museumite, muthmannite, nagyágite, 

petzite, and stutzite. The telluride minerals were deposited 

at temperatures of between 200 and 300 oC from very dilute 

hydrothermal solutions (Cook and others, 2005).

Other low-sulfidation epithermal deposits with an 

abundance of telluride minerals could be future sources of 

tellurium resources. For example, the Jefferson gold-silver 

project, which is located a few kilometers east of the giant 

Round Mountain low-sulfidation epithermal deposit in 

Nye County, Nev., has also been evaluated as a primary tellurium resource (Mexivada Mining Corp., 2008). This evaluation 

is still very speculative, however, as such a possibility is based 

solely upon a small number of drill intersections with assays 

of tens to hundreds of parts per million tellurium.

High-sulfidation epithermal deposits are a significant 

gold source, and some have produced appreciable amounts of 

copper (for example, the Lepanto deposit in Benguet Province, 

Philippines, and the El Indio deposit in the Coquimbo region 

of Chile). The deposits are synvolcanic in the centers of 

andesite-dacite volcanic complexes. Copper-arsenic-antimony 

sulfosalts are the principal copper-bearing minerals. Tellurium 

commonly substitutes for arsenic and antimony in tennantite 

locally and in sufficient abundance to form the tellurian 

endmember goldfieldite. In all well-studied high-sulfidation 

epithermal deposits, precious-metal telluride minerals have 

been documented, although they are generally less economically important than native gold and electrum. At the El Indio 

deposit, native tellurium is associated with high-sulfidation 

alteration (Siddeley and Araneda, 1986).

Kesler and others (2003) documented the distribution 

of tellurium in the high-sulfidation epithermal deposits of the 

Pueblo Viejo district in Sánchez Ramírez Province, Dominican 

Republic. They concluded that tellurium appears to increase 

steadily upward in the deposit. In the district’s Monte Negro 

orebody, higher tellurium concentrations generally correspond 

to higher gold concentrations, and both increase steadily 

upward in the deposit. Lehmann and others (1999) discussed 

the abundance of telluride minerals in the Bereznyakovskoe 

high-sulfidation epithermal deposit in Chelyabinskaya Oblast’ 

in the Urals. They noted that precious-metal mineralization 

was associated with a tetrahedrite-telluride hydrothermal 

stage, and that mineralized and altered rock varied from tens 

to hundreds of parts per million tellurium. At the Lepanto 

deposit, Benguet Province, Philippines, enargite grains were 

noted to contain many hundreds of parts per million tellurium 

(Deyell and Hedenquist, 2011). Chaffee and others (1981) 

analyzed stream sediment samples from the Patagonia Mountains in Arizona and found tellurium anomalies (>399 ppb) 

consistently associated with high-sulfidation alteration, as well 

as a high local background of 100 to 399 ppb tellurium and 

a broader regional background of 15 to 99 ppb tellurium in 

samples more distal to the mineralization.

The variety of telluride minerals in high-sulfidation 

deposits is not extensive. The common telluride minerals are 

calaverite, hessite, krennerite, petzite, and sylvanite. Arribas 

and others (1995) reported the occurrence of the secondary 

tellurite minerals blakeite, emmonsite, and rodalquilarite at 

Rodalquilar, Andalucia, Spain.

The Deer Horn intermediate-sulfidation epithermal 

deposit in west-central British Columbia, Canada, contains 

high gold and silver grades with abundant base-metal 

sulfides and telluride minerals. A National Instrument 

(NI) 43–101 resource evaluation (Deer Horn Metals Inc., 

2012) suggested resources of 66 metric tons of tellurium, 

as well as 2,100 kilograms (kg) (68,000 troy ounces) of 

gold and 65,000 kg (2.1 million troy ounces) of silver. 

Tellurium concentrations in drill intercepts with the highest 
Geology R11

precious-metal grades are generally 150 to 250 ppm, and some 

of the highest values were in the thousands of parts per million 

tellurium; analyses of these zones indicate a silver-to-tellurium 

ratio that is generally about 1:1.

The Sandaowanzi intermediate-sulfidation epithermal 

deposit in Heilongjiang Province is one of the many 

extension-related Early Cretaceous epithermal and porphyry 

deposits of northeastern China. The deposit contains more 

than 15 metric tons of gold and has been mined since 2006. It 

is almost entirely characterized by micrometer- to millimetersize disseminated and quartz vein-hosted, coarse-grained 

silver- and gold-bearing telluride minerals, with lesser 

amounts of lead- and mercury-bearing telluride minerals; the 

largest grains are 3 centimeters in diameter (Liu and others, 

2011, 2013). Tellurium is likely being recovered, although 

the amount being recovered is not available.

Some tellurium-rich deposits have descriptive attributes 

of both low- and high-sulfidation deposits; for example, the 

deposit at Kochbulak in Toshkent Province, Uzbekistan. 

Here, the telluride mineral assemblage is more diverse than 

is typical of high-sulfidation deposits, and the tellurium 

content of ores averages 101.6 ppm (Plotinskaya and others, 

2006). In the Baguio district, Benguet Province, Philippines, 

both intermediate- and high-sulfidation epithermal mineral 

deposits are present, whereas abundant telluride minerals are 

associated only with the advanced argillic alteration in the 

high-sulfidation ores.

Orogenic Gold Deposits

Structurally controlled gold deposits in metamorphic 

rocks, or orogenic gold deposits, may contain high abundances 

of tellurium-bearing mineral phases. High concentrations 

of tellurium are present in both Precambrian deposits in 

greenstone belts and in Phanerozoic gold deposits hosted 

in metasedimentary rock-dominant orogens. Similar to the 

epithermal deposits, these concentrations of tellurium have so 

far not been recovered during gold mining, although goldbearing tellurides can constitute a significant amount of the 

recovered gold ore. In many districts, specific deposits hosted 

by local intrusions, rather than those in widespread metavolcanic or metasedimentary country rocks, tend to contain 

a high abundance of telluride minerals (see the discussion 

of Kirkland Lake and Kensington below), suggesting that 

oxidation of ore fluids by the felsic to intermediate igneous 

rocks provides, in some cases, a major control on destabilizing 

hydrothermal tellurium complexes. Abundant tellurides and 

the presence of hydrothermal hematite in other deposits, such 

as the Golden Mile deposit in Western Australia, Australia, and 

the Commoner deposit in Mashonaland West, Zimbabwe, also 

suggest some degree of oxidation of the ore fluids.

The world’s largest Archean orogenic gold deposit, the 

Golden Mile, which is hosted in mainly mafic metavolcanic 

rocks of the Yilgarn Craton of Western Australia, contains 

1,500 metric tons of gold. Although the majority of the 

ore is hosted in arsenic-rich pyrite, about 20 percent of the 

resource is hosted in late-stage telluride minerals (Lindgren, 

1906; Mueller and Muhling, 2013). Many of the higher-grade 

gold zones contain hundreds of parts per million tellurium. 

Recognized minerals that concentrate gold, mercury, and (or) 

silver, as well as the tellurium, include calaverite, coloradoite, 

hessite, and petzite, with a total of 19 distinct telluride phases 

identified. Mueller and Muhling (2013) described a zoning of 

telluride minerals that progressed from a core of calaverite to 

hessite to an outer, very late coloradoite zone. Shackleton and 

others (2003) noted the telluride minerals occur as fine-grained 

composite aggregates, massive segregations, composite grains 

with native gold, inclusions in and intergrowths with pyrite 

and tetrahedrite-group minerals, or inclusions in carbonates, 

quartz, and tourmaline. Eilu and others (1998) measured 

tellurium anomalies of greater than 10 ppm in mafic volcanic 

rocks for more than 100 meters (m) beyond the Western 

Australia ore zones and suggested that tellurium is a very 

important geochemical pathfinder for orogenic gold in the 

Yilgarn Craton.

Similarly, many of the world-class Archean gold deposits 

of the Abitibi and Yellowknife greenstone belts in Canada may 

have elevated concentrations of tellurium, typically in goldbearing telluride minerals. Telluride minerals are common 

in deposits in the Kirkland Lake and Red Lake districts of 

Ontario, Canada. As in other deposits, the tellurides, as well 

as much of the gold, are typically paragenetically late stage 

(for example, Thomson, 1928). Native gold in the syenitehosted gold-bearing veins at the Kirkland Lake deposit occurs 

with altaite, calaverite, coloradoite, and petzite, as well as 

sericite. Highest grade ore samples of about 200 to 400 ppm 

gold also contain about 200 to 500 ppm molybdenum and 

300 to 700 ppm tellurium, with a typical gold-molybdenumtellurium signature interpreted as representing an alkali 

magmatic ore fluid source that is unusual in the district 

(Ispolatov and others, 2008). Boyle (1979) reported that the 

tellurium in the ores of the Yellowknife mining district in 

Northwest Territories, Canada, contained, in the main sulfide 

stage, as much as 1,000 ppm tellurium in galena, 100 ppm 

tellurium in stibnite, and 100 ppm tellurium in lead-bearing 

sulfosalt minerals.

A strong gold-tellurium association is also recognized 

at Ashanti in Western Region, Ghana, which hosts the 

world’s largest Paleoproterozoic orogenic gold deposit. Four 

microscopic telluride minerals (altaite, coloradoite, hessite, 

and petzite) and seven submicroscopic telluride minerals have 

been analyzed at Ashanti within a hematite-quartz-telluride 

assemblage; the assemblage was deposited with minor 

amounts of native gold that formed after an earlier event that 

deposited the bulk of the gold ore (Bowell and others, 1990). 

The telluride mineral distribution is zoned from an inner 

coloradoite-rich assemblage in the veins to an assemblage 

dominated by altaite along the vein walls and in the country 

rocks. The late telluride event was shown to reflect a cooling 

of fluid temperature and associated oxidation, because 

hypogene hematite and goethite were deposited with the 

telluride minerals.
R12 Critical Mineral Resources of the United States—Tellurium

The Phanerozoic deposits in accreted metasedimentary 

rock belts are also commonly characterized by an abundance 

of gold- and silver-rich telluride minerals. Altaite, calaverite, 

melonite, petzite, and sylvanite are associated with the main 

mineralization event that formed the gold veins, stockworks, 

and breccias in the early Paleozoic clastic rocks at the Kumtor 

deposit, Ysyk-Kol Oblusu, Kyrgyzstan (Ivanov and others, 

2000). The Cretaceous California Mother Lode deposits 

have been long recognized for their common gold-, lead-, 

nickel-, and silver-bearing telluride minerals that are associated with native gold (Turner and Becker, 1894). As noted 

above, intrusions emplaced into these terranes, which are 

often structurally favorable hosts for the vein deposits, tend 

to have greater abundances of telluride minerals. The Eocene 

Kensington gold deposit in the Juneau gold belt, southeastern 

Alaska, is dominated by veins within an intermediate stock 

with native gold, calaverite, and petzite, and lesser amounts 

of alatite, hessite, sylvanite, tellurobismuthite, and volynskite 

(Casey, 2000). Much of the gold endowment is hosted in the 

two main telluride minerals. Similarly, the tonalite-hosted 

Willow Creek district in the Matanuska-Susitna Borough of 

south-central Alaska contains abundant telluride minerals in 

the main ore assemblage; coloradoite is the dominant mineral, 

and altaite and nagyagite occur in lesser amounts (Burleigh, 

1987). Interestingly, the main gold ores of the Grass Valley 

district of Nevada County, Calif., are hosted by north-striking 

veins cutting granodiorite, but these lack telluride minerals; 

only auriferous east-west-striking veins in adjacent mafic 

and ultramafic country rocks contain lead- and silver-bearing 

telluride minerals (Johnston, 1940).

Many Russian and central Asian Phanerozoic gold 

deposits report significant platinum and palladium grades. 

Whereas some of this may reflect high backgrounds of PGMs 

in black shale host rocks, some PGMs are clearly associated 

with numerous tellurium-bearing minerals. For example, at the 

Sukhoi Log gold deposit in Irkutskaya Oblast’, eastern Russia, 

palladium-bearing telluride minerals are present, as well as 

many bismuth-, gold-, lead-, and silver-rich telluride minerals 

(Distler and Yudovskaya, 2005).

Low-grade (<1 g/t gold) disseminated-style mineralization is associated with the bulk-tonnage open pit mining 

operations along the Mojave-Sonora megashear zone in 

northern Sonora, Mexico. A small, recently worked underground operation along the southeastern side of the shear 

zone, the La Bambolla deposit in the Moctezuma district 

(Mexivada Mining Corp., 2008; E.O. Monteagudo, unpub. 

data, 2009) developed ores from gold-hematite-quartzpyrite-telluride veins, which averaged, in order of value, 

4 g/t gold, 5 g/t silver, and 2.5 g/t tellurium; it is not clear, 

however, whether any of the tellurium was recovered along 

with the precious metals. Some mineralized grab samples 

were reported to contain as much as 3.2 percent tellurium. 

Although classification of this deposit is not clear, it occurs 

in a belt of orogenic gold deposits, and La Bambolla would 

be one rare example of an orogenic gold deposit from which 

tellurium is produced as a primary commodity.

Carlin-Type Gold Deposits

The Carlin-type gold deposits in Nevada are Tertiary 

replacement ores in carbonate-dominant rock sequences that 

have undergone decarbonization and sulfidation, and ores 

occur as submicron-sized gold in disseminated pyrite. In 

contrast to the epithermal and orogenic gold ores, Carlin-type 

ores are not widely recognized for containing abundant 

telluride minerals. In part, this may reflect the exceptional 

fine-grained nature of the ores, so that the telluride minerals 

are rarely visible. Nevertheless, Harris and Radtke (1976), in 

one of the first geochemical studies of the Carlin gold deposit 

itself (which is located in Eureka County, Nev.), indicated 

a strong gold-tellurium correlation, which would favor the 

presence of telluride minerals in this deposit type. Emsbo 

and others (2003) noted minor amounts of telluride minerals 

and as much as 130 ppm tellurium in ore-related pyrite in the 

Meikle deposit of Elko County, Nev. Cail and Cline (2001) 

report values for ores of as much as 180 ppm tellurium in the 

Getchell trend in Humboldt County, Nev., and Heitt and others 

(2003) indicate as much as 84 ppm tellurium at the Deep Star 

deposit in Eureka County, Nev. Ore-stage pyrites in the Carlin 

ores may contain thousands of parts per million tellurium 

(Muntean and others, 2011).

Whereas telluride minerals are typically not reported 

from the classic Carlin ores in Nevada, the Carlin-like 

Zarshuran deposit in West Azerbaijan Province, Iran, contains 

coloradoite and tellurium-bearing orpiment that were 

deposited with arsenian pyrite during the main gold event 

(Asadi and others, 2000). This could suggest either variability 

in the metal source or hydrothermal fluid chemistries among 

Carlin-type gold deposits, or more-significant differences 

in deposit type than originally thought between the defined 

Carlin deposits of Nevada and the Zarshuran ores.

Resources and Production

Identified Resources and Reserves

Specific global and domestic sources for tellurium are 

very poorly known because tellurium is mostly a byproduct 

of the mining of a number of different base-metal-rich deposit 

types. Tellurium concentrations are highest in precious-metalbearing deposits that contain abundant telluride minerals; 

these deposits, however, currently are not significant sources 

of tellurium because the ore-processing method to extract 

gold and silver is not amenable to tellurium recovery. Instead, 

tellurium concentrated in sulfide minerals in base-metal-rich 

deposits is the current main source of tellurium because the 

refining process for such base-metal-rich ores is amenable 

to tellurium recovery. The only certain present-day primary 

sources for tellurium ore are the Kankberg VMS deposit 

in Västerbotten County, Sweden, which produces about 

10 percent of the world’s tellurium, and the Dashuigou and 

Majiagou epithermal deposits in China, which together 

produce from 2 to 7 percent of the world’s tellurium (fig. R1). 
Resources and Production R13

Whether the orogenic gold-tellurium deposits in Sonora, 

Mexico, are also active primary tellurium producers is not 

clear (for example, La Bambolla with a possible eastern 

extension at AuroTellurio). Importantly, prior to leaching and 

recovery, most gold ores are roasted or subjected to other 

pre-leaching processes. Tellurium is not recovered during 

this pretreatment and is typically lost from later recovery.

The availability of economically recoverable tellurium 

depends on the production of copper as a (or the) primary 

product. Reliance on mining tellurium as a primary material 

would drastically increase its cost. At least 75 percent of 

the world’s recovered tellurium is produced as a byproduct 

of copper mining, although not all the mined tellurium 

is recovered at the copper refineries (Anderson, 2015). 

Tellurium can only be extracted from copper that is refined 

by the electrolytic process, a technique that is cost-effectively 

applied to high-grade copper ores (Moss and others, 2011). 

In such a process, tellurium and other impurities, such as gold, 

silver, and platinum, are concentrated at and removed from 

the crude copper anodes. Typically, the slimes contain from 

1 to 4 percent tellurium, although as much as 8 to 9 percent 

tellurium in the slimes has been reported at some refineries. 

These numbers reflect a recovery rate of a little less than 

50 percent of the available tellurium within the slimes. For 

refining lower grade copper ores, the most economical process 

is the solvent-leach refining process, which is currently not 

capable of recovering tellurium. The global decrease in the 

supply of high-grade copper ores may constrain the future 

supply of tellurium from copper (Moss and others, 2011). 

In addition to copper refineries, as much as 8 to 10 percent 

of the world tellurium supply is recovered from lead refinery 

skimming and from the flue dust and gases of copper smelting 

(Moss and others, 2011). Not only are data from many of these 

milling operations unavailable, but also little information 

about the source of the ores is available. Concentrates are 

commonly shipped to a mill in a different country than that in 

which the ore is mined and, therefore, the original source of 

the recovered tellurium is not known.

A small amount of tellurium is currently recovered from 

recycled sources, such as selenium-tellurium solar cells. The 

leading cadmium-tellurium solar cell manufacturer has a freeof-charge recycling program for discarded photovoltaic solar 

cell panels, which could become more important in the future. 

When large volumes of currently installed panels reach their 

end-of-life stage, they will be recycled for secondary tellurium 

or cadmium-tellurium. Pilot recycling facilities recover as 

much as 95 percent of semiconductor material and can provide 

feedstock for future production of photovoltaic solar panels 

(Moss and others, 2011).

In the United States, tellurium demand is met partly 

domestically and partly by imports. Almost all the domestic 

commercial-grade tellurium production is from ASARCO’s 

copper refinery in Texas, where it is recovered as a byproduct 

of copper anode slimes and lead refinery skimming. Resulting 

exports from the United States in 2012 were 47 metric tons of 

tellurium, whereas 36 metric tons of tellurium was imported 

from refineries abroad. About 66 percent of the tellurium was 

imported from China and Canada, and lesser amounts were 

imported from the Philippines and Belgium (George, 2014b).

Global reserves of tellurium were 24,000 metric 

tons in the beginning of 2014, which includes estimates 

of 3,600 metric tons in Peru, 3,500 metric tons in the 

United States, and 800 metric tons in Canada (George, 

2014b). The lack of reserves in Chile is surprising, given 

that its porphyry deposits make Chile the world’s leading 

copper producer; this indicates the lack of publicly available 

data for Chile, which also is characteristic of much of the 

global tellurium supply. Refineries around the world are 

estimated to produce 450 to 470 metric tons of tellurium 

annually as tellurium metal, crude tellurium dioxide, or crude 

copper telluride; the latter two commodities must be further 

processed to produce tellurium metal. Data for tellurium metal 

recovered in specific countries in 2010 include 41 metric 

tons from Sweden, 40 metric tons from Japan, 35 metric tons 

from Russia, 30 metric tons from Peru, more than 30 metric 

tons from China, and 6 to 10 metric tons from Canada. 

Data from the United States are proprietary, but the British 

Geological Survey estimates that 50 metric tons of tellurium 

are recovered in the United States (Brown and others, 2013). 

Other production of tellurium metal comes from refineries in 

Australia, Belgium, Chile, Colombia, Germany, Kazakhstan, 

Mexico, the Philippines, and Poland. Estimated refinery 

production by world region is shown in figure R4.

China

19%

Europe

29%

Japan

16%

Other Asia

16%

North America

7%

Commonwealth

of Independent

States

11%

Latin America

2%

Figure R4. Pie chart showing percentage of estimated world 

tellurium refinery production in 2013, by country or region. 

Except for two primary producers in Sweden and China, which 

may contribute 15 percent of the annual global production of 

approximately 450 to 470 metric tons of tellurium, the majority of 

global production is as a byproduct from the refining of copper 

and other base-metal ores. Compiled using data from various 

sources, including George (2014a).
R14 Critical Mineral Resources of the United States—Tellurium

Undiscovered Resources

There are no available estimates of undiscovered 

resources of tellurium. Hypothetically, existing waste piles 

from the milling of large-tonnage copper, iron, and other 

massive base-metal orebodies may contain economically 

recoverable amounts of tellurium. Maness (2010) suggests 

a potential for economic amounts of tellurium, as well as 

for other critical and near-critical elements, could exist in 

waste from massive sulfide ores in locations such as the 

Ducktown (Tennessee), Leadville (Colorado), and Joplin 

(Missouri) districts, as well as high-sulfur coal waste 

dumps in the United States. He pointed out that the world’s 

best potential source for such elements could be the giant 

copper-slag heaps on the island of Cyprus from centuries 

of mining. Detailed analytical studies of sulfide-rich 

waste material in such old mining camps could fingerprint 

areas in the United States and elsewhere with the highest 

potential as future sources of readily recoverable tellurium 

(Cone, 2013).

Whereas only a few gold deposits worldwide are 

producers of tellurium, abundant gold-telluride minerals 

are the one viable source for the mining of primary 

tellurium ores. As noted previously, most (likely as much 

as 80 to 85 percent) of the world’s tellurium is produced 

from anode slimes collected during the electrolytic 

refining of copper. If, however, the more economical 

solvent-leach procedure is increasingly used in the future 

to recover lower grades of copper from porphyry and other 

large-tonnage deposits, which is a process that does not 

recover tellurium, then additional high-grade, telluriumrich gold deposits may become new primary sources for 

tellurium. This is likely particularly true of the epithermal 

gold deposits associated with alkaline magmatism, such 

as those in the Cripple Creek district of Teller County, 

Colo. Improved extraction techniques for the recovery of 

tellurium from telluride-mineral-rich ores, coupled with 

lesser amounts of tellurium recovered from electrolytic 

refining, could lead to the production of tellurium from ore 

milled at deposits such as Cripple Creek.

Unconventional resources of tellurium in the world’s 

oceans may become economic in the future should sea-floor 

mining become feasible. As pointed out by Hein and others 

(2013), iron-manganese crusts that precipitate onto the 

surface of seamounts, ridges, and plateaus at depths of 

400 to 7,000 m can contain tens of parts per million tellurium. This is about three orders of magnitude greater than 

concentrations found in tellurium-enriched marine black 

shales that are now preserved on continents. It is possible 

that, in future decades, the global tellurium demand could 

be met by recovery of tellurium-rich crusts from some of 

the shallower marine operations.

Exploration for New Deposits

Watterson and others (1977) point out that tellurium is 

a useful pathfinder element to aid in the exploration for the 

many types of ore deposits described above. They define 

well-developed tellurium dispersion halos associated with 

porphyry, epithermal, and polymetallic vein deposit types. In 

the weathering zone, tellurites (for example, emmonsite and 

poughite) become enriched relative to sulfates because sulfates 

are much more soluble. They note that average concentrations 

of about 2 ppm tellurium occur in gossans surrounding the 

porphyry copper ores in Ely, Nev. Anomalous tellurium was 

detected surrounding silver-lead-zinc veins in central Colorado 

more than a few hundred meters beyond any visible alteration 

(Watterson and others, 1977). Values of as much as 1 ppm 

tellurium in stream sediments were noted to be excellent 

indicators of upstream gold-bearing telluride ores in the 

southwestern Pacific epithermal deposits (Rytuba and Miller, 

1990). Tellurium gave the broadest halos of any pathfinder 

surrounding orogenic gold deposits in Western Australia, 

Australia (Eilu and others, 1998).

These tellurium anomalies obviously do not indicate 

ore deposits with tellurium as a main commodity, but rather, 

deposits of a variety of types where tellurium could be 

present at levels that would make it an important recoverable byproduct. Because tellurium as a primary commodity 

is extremely rare, exploration for “tellurium deposits” is 

essentially almost never undertaken by the private sector. 

Geochemical ore pathfinder elements and geophysical 

exploration techniques will differ by mineral deposit type, and 

because tellurium can be enriched in many common varieties 

of hydrothermal ore deposits, specific exploration guides for 

potential tellurium deposits are not definable.

Environmental Considerations

The environmental characteristics of tellurium extraction 

are closely linked to those associated with the mining of the 

two types of copper deposits that represent its main sources: 

porphyry copper and VMS deposits. These two deposit types 

differ from one another significantly in terms of their geology 

and their environmental characteristics. Tellurium also occurs 

at elevated concentrations in some precious-metal deposits, 

but tellurium is presently produced from these in only a 

few locations.

The environmental geology of porphyry copper 

deposits has recently been reviewed by John and others 

(2010). Because the deposits are large in size and of low 

grade, they are commonly mined by open pit methods. The 

solid mine wastes have variable potential to generate minor 

amounts of acid drainage or neutralize it. The environmental 

geology of VMS deposits has been reviewed by Seal and 

Hammarstrom (2003), Seal and Piatak (2012), and Shanks and 
Environmental Considerations R15

Thurston (2012). The deposits are typically of moderate size, 

having lens-like or tabular ore geometries, high grades of base 

metals, and a high potential of solid mine wastes to generate 

acid drainage.

Sources and Fate in the Environment

The abundance of tellurium in the upper continental 

crust is quite low (3 ppb) (Li and Schoonmaker, 2003). 

The low environmental concentrations of tellurium in all 

environmental media are a reflection of its low crustal 

abundance. Because of its low abundance, tellurium is rarely 

included in environmental surveys unless there is a specific 

interest. Therefore, data on its concentration in environmental 

media are limited. Andreae (1984) reported concentrations of 

0.51 and 3.32 parts per trillion (ppt) for rainwater. Seawater 

concentrations of tellurium are significantly lower and range 

between 0.03 and 0.87 ppt (Andreae, 1984; Lee and Edmond, 

1985). Under oxygenated conditions in surface water, 

tellurium is present as the oxidized species tellurate (Te(VI)); 

the solubility of tellurium at 25 °C at pH 7 is approximately 

25 ppb (McPhail, 1995).

For solid environmental media, data are available for 

tellurium in soils and plants. In the national geochemical 

survey of soil in the United States, Smith and others (2013) 

reported concentrations of tellurium ranging between less than 

0.1 and 50.5 ppm, with most values less than 0.1 ppm. Cowgill 

(1988) documented concentrations in plants with mean values 

(dry basis) that ranged between 4.2 and 25.9 ppb, and found a 

strong correlation between soil and plant concentrations.

Information on pre-mining environmental concentrations 

of tellurium is lacking owing to its low concentrations. From 

an environmental perspective, baseline concentrations of other 

elements of environmental interest associated with porphyry 

copper or VMS deposits are of greater interest than those of 

tellurium because of the potential during mining to significantly elevate their concentrations in the environment. Few 

environmental baseline studies of porphyry copper deposits 

have been conducted, particularly with respect to the diversity 

of climatic settings in which they are found; however, some 

pre-mining soil and sediment information is available in mine 

permit applications for recently proposed mines. Studies from 

diverse climates, such as the cold climates of southwestern 

Alaska (Anderson and others, 2011); tropical climates of 

Puerto Rico (Learned and Boissen, 1973; Plaza-Toledo, 2005); 

and hot, arid climates of Arizona and Chile (Chaffee, 1976, 

1977; Chaffee and others, 1981; Leybourne and Cameron, 

2006, 2008) show similar geochemical features. Soils and 

sediments show elevated concentrations of iron, arsenic, 

cadmium, copper, molybdenum, lead, and zinc.

Pre-mining surface water and groundwater data 

are also limited. Plaza-Toledo (2005) and Anderson and 

others (2011) conducted pre-mining water-quality studies 

in the humid settings of Puerto Rico and southwestern 

Alaska (Pebble deposit), respectively, and Leybourne and 

Cameron (2006, 2008) investigated groundwater quality 

in the arid Atacama Desert of Chile. In humid settings, the 

pH ranged between 4.1 and 8.6. Alkalinity ranged between 

0 and 100 ppm calcium carbonate (CaCO3) equivalent; 

sulfate, between 1 and 85 ppm; and hardness, between 

2 and 130 ppm CaCO3 equivalent. Dissolved trace element 

concentrations showed a range of values, including 

arsenic (<1 to 36.2 ppb), cadmium (<0.2 to 11.6 ppb), 

copper (<0.5 to 688 ppb), iron (<20 to 15,900 ppb), lead 

(<0.05 to 18.8 ppb), molybdenum (<2 to 21.9 ppb), and zinc 

(<0.5 to 68 ppb). Leybourne and Cameron (2006, 2008) 

documented high-salinity groundwater associated with the 

undeveloped Spence copper deposit in the Atacama Desert, 

Chile, reaching salinities of 1.0 to 5.5 percent, with one outlier 

at 14.5 percent. The pH of the groundwaters varied widely 

between 4.7 and 9.2. Dissolved sulfate concentrations in 

groundwaters were between 5,000 and 10,000 ppm. Dissolved 

trace element concentrations showed a range of values 

for arsenic (<10 to 160.9 ppb), copper (9 to 28,991 ppb), 

iron (4 to 54,454 ppb), molybdenum (2 to 475 ppb), 

lead (0.025 to 23.8 ppb), and zinc (5 to 1,344 ppb).

The geologic and geochemical features of VMS deposits 

are manifested in the surrounding environmental media, such 

as soil, stream sediment, surface water, and groundwater. 

Soils and stream sediments surrounding undeveloped 

deposits can have elevated concentrations of arsenic, barium, 

cadmium, copper, iron, lead, nickel, and zinc. The surface 

water and groundwater data associated with undisturbed 

VMS deposits span a range in pH from 3 to 10; a range of 

dissolved iron (a dominant cation) from <0.02 to 300 ppm; 

and dissolved sulfate (also the product of pyrite or 

pyrrhotite weathering) from <0.1 to 4,000 ppm (Seal and 

Hammarstrom, 2003; Eppinger and others, 2007; Seal and 

Piatak, 2012;). Other dissolved constituents surrounding 

undeveloped deposits span a range of values, including for 

aluminum (<0.002 to 100 ppm), arsenic (<0.1 to 430 ppb), 

cadmium (<0.1 to 600 ppb), copper (<0.003 to 20 ppm), 

lead (<0.1 to 500 ppb), and zinc (<3 to 200 ppm).

Mine Waste Characteristics

The amount of mine waste associated with porphyry 

copper and VMS deposits varies with the size of the deposit. 

Porphyry copper deposits range from 30 million metric tons to 

greater than 20 billion metric tons of “ore” with a median size 

of 250 million metric tons (John and others, 2010). Because 

copper grades are typically less than 1 percent copper (median 

grade is 0.44 percent copper), more than 95 percent of the 

material mined will end up as solid mine waste of one form 

or another. The VMS deposits are variable in size, ranging 

from 1 to 600 million metric tons; the copper, lead, and 

zinc ore grades are also very variable, but typically average 

2 to 5 percent (Seal and Hammarstrom, 2003).
R16 Critical Mineral Resources of the United States—Tellurium

Solid mine waste at porphyry copper and VMS deposits 

typically falls into two main types: tailings and waste rock. 

A third type, leach pad waste, is present only at those mines 

where the ores are amenable to copper extraction by the 

electrolytic refining process, or where a leached cap may 

have residual gold grades that are amenable to extraction. 

The electrolytic process is not amenable to tellurium recovery. 

Tailings are produced when the ore is crushed to a sand or 

silt size to facilitate the separation of the copper sulfide and 

other metal sulfide ore minerals using a technique known as 

froth flotation. This technique yields a “concentrate” of sulfide 

minerals, and a waste material, known as tailings, which can 

be pumped as a slurry or trucked dry to a tailings storage 

facility on site. The copper concentrate is sent to a smelter and 

refinery complex, where copper and other mined commodities, 

such as tellurium, are recovered. Tailings storage facilities are 

typically impoundments surrounded by a retaining dam.

Waste rock is uneconomic rock that must be removed 

to access the ore. Waste rock is disposed of on site. For open 

pit porphyry copper mines, waste-to-ore (stripping) ratios 

commonly exceed 2:1, which means that for each ton of ore 

mined, two tons of waste rock must be removed (Porter and 

Bleiwas, 2003). Porphyry copper deposits at depth can also 

be mined by a method known as block caving. A vertical shaft 

or spiral decline is built to the base of the orebody, and the 

ore is mined from below, leaving a large, unsupported cavity 

from which the ore was removed. With this type of mining, 

the amount of waste rock is minimal and the waste material is 

dominated by tailings. As with porphyry copper deposits, the 

amount of waste rock associated with VMS deposits depends 

on the mining method, which in turn depends on the geometry 

of the orebody.

The mineralogy of the ore and waste in porphyry copper 

and VMS mines serves as the foundation for understanding 

the environmental characteristics of these deposits. In terms 

of both ecological risk and human health risk, the mineralogy 

influences the acid-generating potential of the ores and 

wastes and the mobility of trace elements. In porphyry copper 

deposits, ore minerals such as bornite, chalcopyrite, enargite, 

galena, molybdenite, pyrite, and sphalerite are the main hosts 

of trace elements of environmental concern; in total, they typically constitute less than 5 percent of the ore (John and others, 

2010). Tellurium is typically hosted by trace amounts of gold-, 

silver-, or PGM-bearing tellurides. In VMS deposits, the ore 

minerals are dominated by pyrite and pyrrhotite, with varying 

amounts of chalcopyrite, galena, and sphalerite, and a variety 

of accessory minerals (Shanks and Thurston, 2012). The 

trace-element geochemistry of tailings reflects the mineralogy 

of the waste material. For porphyry copper deposits, tailings 

typically contain significant concentrations of arsenic, copper, 

manganese, molybdenum, and zinc (John and others, 2010). 

Tailings from copper-dominant VMS deposits can contain 

elevated concentrations of aluminum, arsenic, cadmium, 

copper, iron, lead, sulfur, and zinc (Seal and Hammarstrom, 

2003; Seal and Piatak, 2012).

Approaches for managing solid waste depend upon its 

acid-generating potential. Metals and many other trace elements 

tend to be more soluble at low pH rather than at neutral or 

high pH. Therefore, the acid-generating or acid-neutralizing 

potentials of the waste rock and tailings are of prime importance 

in identifying the potential environmental risks associated with 

mining and ore beneficiation. The acid-generating potential 

of mine waste is expressed in terms of the amount of calcium 

carbonate it would take to neutralize it. Therefore, the units are 

kilograms of calcium carbonate per metric ton of mine waste 

(kg CaCO3

/t) (Price, 2009; International Network for Acid 

Prevention, 2011). The acid-generating potential primarily 

resides in pyrite. Mine waste can also have acid-neutralizing 

potential that resides in carbonate minerals, such as calcite, 

and some silicate minerals, such as feldspars.

The rocks associated with porphyry copper deposits tend, 

in general, to straddle the boundary between having net acidgenerating potential and having net acid-neutralizing potential, 

but the amount of either is relatively small compared with the 

high acid-generating potential associated with VMS deposits. 

For example, at the Pebble porphyry copper deposit in Alaska, 

the various mine waste materials are expected to span a range 

from a net acid-generating potential of 110 kg CaCO3/t to 

a net acid-neutralizing potential of 27 kg CaCO3/t (Pebble 

Partnership, 2011, written commun.). In contrast, mine waste 

from VMS deposits typically has net acid-generating potential, 

commonly ranging between 100 and 350 kg CaCO3/t, although 

some waste can have net neutralizing potential locally 

(Lindsay and others, 2009; Seal and Piatak, 2012).

Human Health Concerns

Tellurium does not appear to be an essential trace element 

for human health (Kobayashi, 2004; Ba and others, 2010). 

No information is available on the toxic effects of tellurium 

on humans, presumably in part because of its low natural 

abundance. The Agency for Toxic Substances and Disease 

Registry does not have a toxicological profile for tellurium, 

nor does the U.S. Environmental Protection Agency have 

drinking water standards or soil guidelines for tellurium. 

Animal studies, using large doses of tellurium administered 

either through ingestion, inhalation, or intravenously, indicated 

the target sites for tellurium toxicity are the kidneys, nervous 

system, skin, and brain in a fetus (Taylor, 1996). The reduced 

species, tellurite, is more toxic than either arsenite or selenite, 

and, tellurite is 10 times more toxic than the oxidized species 

tellurate (Taylor, 1996).

The more significant human health risks associated with 

tellurium production are found with the broader aspects of 

mining porphyry copper and VMS deposits. Contaminated 

groundwater plumes associated with tailings impoundments 

may pose threats to drinking-water supplies, depending upon 

the geologic and hydrologic setting and the engineering 

aspects of the mine and the waste piles for both deposit types. 

Host rocks with higher neutralization potentials, such as 
Environmental Considerations R17

carbonate rocks, tend to limit the mobility of metals and 

related compounds. Hydrologic and climatic settings that 

have net evaporative loss of water may cause evaporative 

concentration of solutes that may enter groundwater used as 

drinking-water supplies in the vicinity of mines and waste 

piles. Improperly constructed water-containment structures 

may allow contaminated mine waters to enter surrounding 

groundwater. For both porphyry copper deposits and VMS 

deposits, elements or compounds with the greatest likelihood 

of causing problems for drinking-water sources include 

aluminum, arsenic, cadmium, copper, iron, manganese, sulfate, 

and zinc. The potential for groundwater contamination depends 

upon the waste- and water-management practices in place at 

any given mine.

Ecological Health Concerns

As with the human health guidelines, no information 

is available for tellurium and its toxic effects on aquatic or 

terrestrial ecosystems, whether through surface water, soil, or 

sediment. The U.S. Environmental Protection Agency does 

not have surface water or sediment guidelines for tellurium. 

Thus, the more significant ecosystem risks associated with 

tellurium production are found with the broader aspects of 

mining porphyry copper and VMS deposits. Many of the 

ecological risks associated with both deposit types focus on the 

ability of mine wastes to generate acid, and the ability of the 

resulting acid mine drainage to carry metals and other inorganic 

contaminants. As described above with respect to mine waste 

characteristics, the acid-generating potential of mine waste is 

controlled by the pyrite content, which may be offset by the 

acid-neutralizing potential of carbonate or silicate minerals. 

Tailings are prone to water transport, particularly in the case of a 

tailings dam failure, and wind transport, because of the fine-size 

sand and silt grains. Thus, tailings present additional potential 

risks to aquatic organisms through sediment contamination.

Mine drainage data are available for porphyry copper 

deposits in British Columbia, Canada (Day and Rees, 2006); 

the Globe district in Arizona (Eychaner, 1991; Stollenwerk, 

1994; Brown and others, 1998; Lind and others, 1998; Conklin 

and others, 2001); the Morenci district in Greenlee County, 

Ariz. (Enders and others, 2006); and the Sar Cheshmeh 

deposit in Iran (Khorasanipour and others, 2011). Many of 

these values exceed relevant water-quality guidelines for the 

protection of aquatic ecosystems for aluminum, copper, iron, 

manganese, and zinc (U.S. Environmental Protection Agency, 

2009). The pH can commonly range from 2 to 8.5, and sulfate, 

which is the dominant anion, can locally reach 30,000 ppm.

Pit lakes, particularly in porphyry copper mining districts 

that are riddled with historical underground mine workings, 

such as at Butte, Mont., which has a 140-year mining history, 

can be problematic. The Berkeley Pit lake at Butte contains 

more than 100 billion liters of pH-2.5 mine water (Gammons 

and Duaime, 2005; Gammons and others, 2005). The pit 

lake has high dissolved solids and elevated concentrations 

of copper, iron, sulfate, and zinc (Gammons and Duaime, 

2005). In contrast, pit lakes in the Yerington (Lyon County) 

and Robinson (White Pine County) districts, Nevada, have 

pH values that typically range from 7.0 to 8.5 with a few 

outliers near 4.7, which have been attributed to discharge from 

solvent-extraction operations rather than groundwater-rock 

interactions in the vicinity of the pits (Shevenell and others, 

1999). In Nevada, total dissolved solids are generally less 

than 6,000 ppm; concentrations of iron reach a maximum of 

4.5 ppm; arsenic, less than 0.05 ppm; manganese, less than 

5 ppm; and selenium, less than 0.14 ppm.

The VMS mines are noted for their significant impacts 

on the aquatic ecosystem associated with abandoned historical 

mines, such as at Iron Mountain, Calif. Iron Mountain is 

unique among abandoned mines in that its effluent and seeps 

within the mine workings have extremely low pH (–4 to 1.5) 

and dissolved copper, iron, sulfate, and zinc are present in 

concentrations reaching hundreds of grams per liter (Alpers 

and others, 2003). The unusual chemistry of mine drainage 

at Iron Mountain is the result of an atypical combination of 

factors related to its geologic, hydrologic, and climatic setting. 

In contrast, the pH values from mine drainage from a broader 

spectrum of VMS mines are typically much lower than 

the values recorded at Iron Mountain, ranging from 2 to 8, 

including concentrations of dissolved copper (<0.1 ppb to 

3 ppm), iron (1 ppb to 3 grams per liter [g/L]), sulfate (1 ppm 

to 10 g/L), and zinc (<1 ppb to 1 g/L) (Seal and Hammarstrom, 2003). Other trace elements found in environmentally 

significant concentrations associated with abandoned VMS 

mines include arsenic, cadmium, lead, and nickel.

Carbon Footprint

Tellurium extraction through mining does not have a 

unique carbon footprint beyond the energy requirements of 

mining in general. In terms of its uses, the most direct link of 

tellurium to the global carbon cycle is in its use in cadmiumtelluride photovoltaic cells (Anderson, 2015).

Mine Closure

The nature of mine closure for porphyry copper and 

VMS deposits depends primarily on the method of mining and 

the characteristics of the waste material. Open pit mines for 

either deposit type display at least three different features after 

mining: the open pit, tailings storage facilities, and waste-rock 

piles. Because of the scale of mining, backfilling of the pits is 

typically not practical. If the water table is above the bottom 

of the pit, then it will become a lake. The water quality of 

the lake will depend upon a number of factors, including the 

characteristics of the wallrocks, the extent of interconnected 

underground mine workings that open to the pit, the water 

level in the pit, the local hydrology, and climate (Castendyk 

and Eary, 2009).
R18 Critical Mineral Resources of the United States—Tellurium

The long-term fate of tailing storage facilities will depend 

upon the nature of the tailings and the method of construction 

of the facility. Some facilities can be regraded, capped, and 

revegetated. Others are designed to have a water cover in 

perpetuity to limit sulfide oxidation. Either type may have 

seepage that may require some form of water treatment. The 

long-term fate of waste-rock piles may include regrading, 

capping, and revegetation. Depending upon the acidgenerating potential of the material, some piles may require 

some form of water treatment.

For underground mines, tailings and waste rock may 

be handled in a similar fashion to the approaches used at 

open pit mines. The potential exists to dispose of some of the 

tailings back into the mined out workings, depending upon 

how the orebody is mined; however, the entire volume of 

tailings cannot be placed back in the mine workings because 

of the volume expansion associated with crushing and 

milling the ore.

Problems and Future Research

In the near future, recovery of lower copper grades may 

force recovery of oxidized copper ores by solvent extraction; 

in this process, it is not possible recover the tellurium as a 

byproduct, as is possible through the smelting of higher grade 

copper sulfide concentrates. This change in method could 

potentially lead to a greater emphasis on precious-metal 

deposits to meet global and domestic tellurium demands. 

Demand has decreased for tellurium in recent years, but as 

China and other developing countries adopt more solar technologies, increased demand in the near future is a possibility.

Future research may focus on methods to recover 

tellurium from two new sources. First, future advances in 

processes for sea-floor mining will be critical if tellurium is 

recovered from extreme enrichments in iron-manganese crusts 

on the ocean floor. Second, investigations of old mine and mill 

tailings from historic large-scale operations may fingerprint 

deposits and deposit types with previously unrecognized, 

extreme abundances of tellurium in old waste piles.

Acknowledgments

David John and Karen Kelley are thanked for their 

careful and valuable reviews of this report. Erin Marsh’s 

help with the graphics is greatly appreciated, and input 

from John DeYoung on an earlier version was very helpful. 

Keith Labay helped create the deposit location map, and 

Suzanne Nicholson provided additional editorial comments 

after the technical review was complete.

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Tin

By Robert J. Kamilli, Bryn E. Kimball, and James F. Carlin, Jr.

Chapter S of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–S

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter S

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................S1

Introduction..................................................................................................................................................S2

Strategic and Critical Resource Issues .........................................................................................S3

Uses and Applications ......................................................................................................................S4

Substitutes for Tin .....................................................................................................................S4

Tin Prices and Pricing ..............................................................................................................S4

Geology.........................................................................................................................................................S5

Geochemistry......................................................................................................................................S5

Mineralogy..........................................................................................................................................S6

Deposit Types .....................................................................................................................................S6

Placer Tin Deposits ...................................................................................................................S6

Granite-Related Tin Deposits ..................................................................................................S7

Classifications of Granite-Related Tin Deposits .........................................................S7

Similarities With Granite-Related Deposits of Other Metals ....................................S8

Effect of Cupola Size and Shape on Tin-Greisen Mineralization .............................S9

Tectonic Settings of Granite-Related Tin Deposits ....................................................S9

Resources and Production ........................................................................................................................S9

Globally Important Mining Areas ..................................................................................................S12

Tin Deposits in the United States ..................................................................................................S14

Undiscovered Resources ..............................................................................................................S14

Exploration for New Deposits .................................................................................................................S15

Environmental Considerations ................................................................................................................S16

Sources and Fate in the Environment ..........................................................................................S16

Mine Waste Characteristics ..........................................................................................................S18

Human Health Concerns .................................................................................................................S19

Ecological Health Concerns ...........................................................................................................S20

Mine Closure.....................................................................................................................................S20

Problems and Future Research ..............................................................................................................S21

Supply Issues....................................................................................................................................S21

Research Issues...............................................................................................................................S21

Acknowledgments....................................................................................................................................S21

References Cited.......................................................................................................................................S22

Suggested citation:

Kamilli, R.J., Kimball, B.E.., and Carlin, J.F., Jr., 2017, Tin, chap. S of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, 

and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and 

prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. S1– S53, http://dx.doi.org/10.3133/

pp1802S.
Tin

By Robert J. Kamilli, Bryn E. Kimball, and James F. Carlin, Jr.1

1

Deceased.

Abstract

Tin (Sn) is one of the first metals to be used by humans. 

Almost without exception, tin is used as an alloy. Because of 

its hardening effect on copper, tin was used in bronze implements as early as 3500 B.C. The major uses of tin today are 

for cans and containers, construction materials, transportation 

materials, and solder. The predominant ore mineral of tin, by 

far, is cassiterite (SnO2

). Tin has not been mined or smelted 

in the United States since 1993 and 1989, respectively.

In 2015, the world’s total estimated mine production of tin 

was 289,000 metric tons of contained tin. Total world reserves 

at the end of 2016 were estimated to be 4,700,000 metric tons. 

China held about 24 percent of the world’s tin reserves and 

accounted for 38 percent of the world’s 2015 production of tin.

The proportion of scrap used in tin production is between 

10 and 25 percent. Unlike many metals, tin recycling is relatively efficient, and the fraction of tin in discarded products 

that get recycled is greater than 50 percent.

Only about 20 percent of the world’s identified tin 

resources occur as primary hydrothermal hard-rock veins, or 

lodes. These lodes contain predominantly high-temperature 

minerals and almost invariably occur in close association with 

silicic, peraluminous granites. About 80 percent of the world’s 

identified tin resources occur as unconsolidated secondary 

or placer deposits in riverbeds and valleys or on the sea 

floor. The largest concentration of both onshore and offshore 

placers is in the extensive tin belt of Southeast Asia, which 

stretches from China in the north, through Thailand, Burma 

(also referred to as Myanmar), and Malaysia, to the islands 

of Indonesia in the south. Furthermore, tin placers are almost 

always found closely allied to the granites from which they 

originate. Other countries with significant tin resources are 

Australia, Bolivia, and Brazil.

Most hydrothermal tin deposits belong to what can 

be thought of as a superclass of porphyry-greisen deposits. 

The hydrothermal tin deposits are all characterized by a 

close spatial, temporal, and genetic association with highly 

differentiated, peraluminous porphyritic granite intrusions. 

The intrusions form pegmatites; disseminated ore; parallel or 

subparallel, greisen-bordered sheeted veins that either cross-cut 

the intrusion or are peripheral to it; skarns; and (or) limestone 

replacements that contain different amounts of cassiterite, 

molybdenite, and wolframite.

The tectonic settings of tin-bearing granites are relatively 

well understood and of limited variety. Tin and tungsten 

deposits and their associated igneous rocks are found mainly 

in continental settings.

Historically, prospecting for tin has been carried out by 

the time-honored methods of panning, drilling, trenching, and 

assaying. Geophysical and geochemical surveys have been 

employed to cover large areas more rapidly, isolating areas 

of possible tin deposits so that drilling can be more effective 

and less costly. Elemental concentrations and relationships 

of the lithophile elements, especially barium, lithium, 

niobium, potassium, rubidium, and zirconium, are the most 

reliable chemical indicators of ore-forming processes and 

tin-bearing potential.

The average human diet includes an intake of about 

10 milligrams per day of tin. Ingestion of tin in significantly 

greater amounts than 10 milligrams per day may lead to a 

stomach ache, anemia, and liver and kidney problems. Exposure to some organo-tin compounds can interfere with brain 

and nervous system function and, in severe cases, can cause 

death. Extended inhalation of tin oxide—an issue mainly for 

those people who work in the tin industry—results in a higher 

potential to develop stannosis, which is a mild disease of the 

lungs caused by the inhalation of tin-bearing dust. Inorganic 

tin is poorly absorbed by the body, and no evidence exists 

for the carcinogenicity of metallic tin and tin compounds 

in humans.

Most placer tin deposits are mined by open pit and (or) 

dredging methods. Mining of alluvial placers in modern 

streambeds and riverbeds is likely to increase the amount 

of sediment delivered downstream. This, combined with 

potential diversion of rivers and streams, may negatively 
S2 Critical Mineral Resources of the United States—Tin

affect downstream ecosystems. Many of the placer deposits 

located in Burma, Indonesia, Malaysia, and Thailand are 

located offshore. Most offshore placer tin deposits are mined 

by dredging methods, which have the potential to negatively 

affect benthic, midwater, and pelagic ecosystems.

In a congressionally mandated U.S. Department of 

Defense study of strategic minerals published in 2013, tin 

has the greatest shortfall amount (insufficient supply to meet 

demand) at $416 million; this amount is more than twice that 

of antimony ($182 million), which is the strategic mineral with 

the next largest shortfall amount (U.S. Department of Defense, 

2013). The United States imported 75 percent of its tin supply 

in 2015. During the period 2012–15, these imports were from, 

in descending order of amount imported, Peru, Indonesia, 

Malaysia, and Bolivia.

A promising advancement concerning research into the 

origin of tin deposits is the recent development of a reliable 

method of analyzing tin isotopes in cassiterite. Although the 

mechanism of transport and deposition of tin is fairly well 

understood, the means by which tin is incorporated into the 

parent magma at the points of magma generation and ascent 

needs further investigation.

Tin metallogenic provinces worldwide are well known. 

Consequently, any undiscovered tin deposits will likely be 

spatially close to known deposits or extensions of the same.

Introduction

Tin was one of the earliest metals known to humanity. 

Aside from native gold and copper, which are found in nature, 

tin and lead were the first metals to be smelted. Because of 

its hardening effect on copper, tin was used in bronze implements as early as 3500 B.C., although the pure metal was 

not commonly used until about 600 B.C. Bronze, which is 

a copper-tin alloy that can be sharpened and is hard enough 

to retain a cutting edge, was used in construction tools as 

well as in weapons. Arsenic-containing bronzes, which 

were probably accidentally created by smelting arsenic- and 

copper-containing minerals, preceded tin bronzes by about 

2,000 years (fifth millennium versus third millennium B.C.). 

Oven-fired Mesopotamian pottery from the Ubaid culture 

(about 6500 to 3800 B.C.) was contemporaneous with the 

origins of metallurgy. The origins of metallurgy, therefore, 

may have been the serendipitous outcome of firing of ceramics 

when ore-bearing rocks were accidentally placed in the kiln 

(Cowen, 1999).

Tin placer deposits commonly form in streambeds where 

running water removes lighter material and leaves the heavier 

cassiterite (SnO2), which has a specific gravity of 6.8 to 7.1, 

behind. Cassiterite was probably first picked up in streambeds, 

as its density and translucent brown-black color make it 

conspicuous.

The area of Cornwall, England, was historically a leading 

world producer of tin. Tin deposits in Cornwall had been 

worked since at least Roman times, although the mining of 

tin likely began there as far back as the early Bronze Age. 

Most mining up to the 17th century was from alluvial sources. 

Cornwall continued to be a major world producer of tin 

up through the 19th century (Penhallurick, 1986, p. 148). 

Although the area still has identified resources of tin, the last 

working tin mine in Cornwall (the South Crofty Mine) was 

closed in 1998 (Carlin, 2000).

The predominant ore mineral of tin is cassiterite, 

although small amounts of tin are recovered from other 

tin-bearing minerals, including stannite (Cu2FeSnS4), cylindrite (Pb3

Sn3,Sb2S14), franckeite (Pb5Sn3Sb2S14), canfieldite 

(Ag8

SnS6), and teallite (PbSnS2). Rock-forming minerals 

that may contain significant amounts of tin (that is, tens to 

hundreds of parts per million) are, in order of increasing 

tin content, muscovite, biotite (annite), titanite, and titanomagnetitite (Wang and others, 2013). Table S1 includes 

a more comprehensive, although not exhaustive, list of 

tin-bearing minerals. The chemical symbol for tin, Sn, comes 

from the Latin stannum; hence, the name stannite for one of 

its minerals. After the Bronze Age and up to the 20th century, 

pewter, which is an alloy of 85 to 90 percent tin coupled with 

copper, antimony, and (or) lead, was the leading material for 

kitchenware, tableware, and other items, such as candlesticks, 

in households that could not afford sterling silver or gold. 

Tin cans, which are made mostly of steel, are so named 

because the steel is coated with tin to form tin plate, which 

has low toxicity.

Tin is silvery, malleable, and not easily oxidized in 

air. The metallic form of tin, or beta-tin (ß-tin, also known 

as white tin), is stable at and above room temperature, 

is malleable, and has an unusual, tetragonal crystal 

structure. Tin becomes a superconductor below 3.72 Kelvin 

(−269.43 degrees Celsius [°C]) (De Haas and others, 1935); 

in fact, tin was one of the first superconductors to be studied. 

In contrast, the nonmetallic form of tin, or alpha-tin (α-tin, 

also known as gray tin), which is stable below 13.2 °C, is 

brittle and has a diamond cubic crystal structure, similar 

to diamond, germanium, and silicon. The α-tin has no metallic 

properties at all because its atoms form a covalent structure 

where electrons cannot move freely. It is a dull-gray powdery 

material with no common uses, other than a few specialized 

semiconductor applications (Holleman and others, 1985). 

The spontaneous transformation of tin from the β form to the 

α form, a phenomenon known as “tin pest,” takes place at 

temperatures of 13.2 °C and below. The presence of impurities (for example, aluminum, zinc, and others) lowers the 

transition temperature to below 0 °C; upon the introduction of 

antimony or bismuth, the transformation may not take place at 

all. The addition of these elements thus increases the durability 

of tin (Schwartz, 2002).
Introduction S3

Table S1. List of selected tin-bearing minerals.

[Data are from Fleischer (1983). NA, not available]

Mineral name Chemical formula Location of geographic occurrence

Berndtite SnS2 Cerro de Potosí, Bolivia

Canfieldite Ag8

SnS6 Bolivia; Tasmania, Australia

Cassiterite SnO2 Worldwide

Colusite Cu3

(As,Sn,V,Fe)S4 Butte, Montana; Australia

Cylindrite Pb3

Sn4FeSb2S14 Bolivia; Tasmania, Australia

Franckeite Pb5

Sn3Sb2S14 Bolivia; Tasmania, Australia

Herzenbergite SnS Bolivia

Hulsite (also called Paigeite) (Fe2+,Mg)2

(Fe3+,Sn)BO5 Brooks Range and Seward Peninsula, Alaska; eastern 

Siberia Russia

Malayaite CaSnSiO5 Malaysia; Cornwall, United Kingdom

Mawsonite Cu6

1+Fe23+Sn4+S8 New South Wales; Tasmania, Australia

Native tin Sn Native tin is found in placer deposits and in unusual 

igneous intrusions. Australia has the recognized 

type locality.

Nigerite (More correctly, 

Ferronigerite–2N1S)

(Zn,Mg,Fe2+)(Sn,Zn)2

(Al,Fe3+)12O22(OH)2

Nigeria

Nordenskiöldine CaSnB2

O6 Norway; Namibia

Ottemannite Sn2

S3 Bolivia

Renierite Cu3

(Fe,Ge,Zn)(S,As)4 Democratic Republic of the Congo (Congo [Kinshasa]); 

Namibia

Sorensenite Na4

SnBe2Si6O16(OH)4 Greenland

Stannite Cu2

FeSnS4 Worldwide

Stannomicrolite (Sukulaite) Sn2

Ta2O7 NA

Stokesite CaSnSi3

O9∙2H2O Cornwall, United Kingdom

Teallite PbSnS2 Bolivia

Thoreaulite SnTa2

O6 Democratic Republic of the Congo (Congo [Kinshasa])

Varlamoffite (?) (Sn,Fe)(O,OH)2 Democratic Republic of the Congo (Congo [Kinshasa]); 

Bolivia

Strategic and Critical Resource Issues

The United States has no tin reserves and is dependent on 

imports to meet its tin requirements (Anderson, 2016b). Tin is 

considered a critical and strategic metal because of its importance to national security (McGroarty and Wirtz, 2012). In a 

congressionally mandated U.S. Department of Defense (DOD) 

study of strategic minerals published in 2013, tin was shown 

to have the greatest shortfall amount (insufficient supply to 

meet demand measured in monetary terms—$416 million) in 

a list of strategic minerals, and more than twice the shortfall 

amount of the next strategic mineral, antimony ($182 million). 

The shortfall amounts result from a modeling process that, 

given a 4-year scenario that assumes 1 year of conflict and 

3 years of recovery and regeneration, computes material 

demand at all levels of the U.S. economy and then compares 

the quantity of material supply with the quantity demanded, 

taking a number of conflict-related factors into account. These 

factors include unavailability of supply from adversaries, 

war damage, shipping losses, infrastructure and (or) ability 

degradation, anti-U.S. orientation, and foreign competition 

(that is, market share) (U.S. Department of Defense, 2013).
S4 Critical Mineral Resources of the United States—Tin

Artisanal and small-scale mining of tin in the Democratic 

Republic of the Congo (Congo [Kinshasa], formerly Zaire) and 

neighboring countries has been identified as a potential source 

of funding for armed groups engaged in civil unrest. This has 

led the United States, through passage of Section 1502 of the 

Dodd-Frank Wall Street Reform and Consumer Protection 

Act (Pub.L. 111–203, H.R. 4173; commonly referred to as 

Dodd–Frank) in 2010, to make it a statutory obligation for all 

companies registered with the U.S. Securities and Exchange 

Commission (SEC) to perform due diligence to determine 

whether the products or components they manufacture contain 

tantalum, tin, tungsten, and (or) gold (3TG) minerals and, 

if so, to determine whether these minerals were sourced 

from Congo (Kinshasa) and (or) its neighboring countries. 

In addition, under rules issued by the SEC, publicly traded 

companies are required to report the sources of 3TG materials 

used, starting in 2014.

In 2015, the United States relied on foreign sources for 

75 percent of its annual apparent consumption of tin. The 

major foreign sources of tin, other than scrap, during the 

period 2011–15 were, in descending order of amount imported, 

Peru, Indonesia, Malaysia, and Bolivia (Anderson, 2017). 

Nearly 30 percent of the U.S. tin supply from 2010 through 

2014 was produced from scrap (Papp, 2016).

Uses and Applications

Almost without exception, tin is used in alloys. Although

it is a metallic mineral, it is rarely used in the elemental form

because it decomposes at temperatures below 13.2 °C (Sainsbury 

and Reed, 1973, p. 637). Up to 28 percent tin can be combined 

with copper, but 10 percent produces an optimum-strength alloy, 

and as little as 2 percent tin will harden copper (Wheeler and 

Maddin, 1980). U.S. consumption of tin by end-use category 

in 2014 was in cans and containers, 23 percent; construction, 

18 percent; transportation, 17 percent; electrical, 12 percent; and 

other, 30 percent (Anderson, 2015). U.S. consumption of tin 

by end use finished product is shown in figure S1.

Substitutes for Tin

Common substitutes for tin in cans and containers 

are aluminum, glass, paper, plastic, or tin-free steel. Other 

materials that are used to substitute for tin are epoxy resins 

for solder; aluminum alloys, copper-base alloys, and plastics 

for bronze; plastics for metal bearings that contain tin; and 

compounds of lead and sodium for some tin chemicals 

(Anderson, 2016b).

Tin Figure S1

Chemicals

15%

Solder

15% Tinplate

24%

Babbitt

(bearing surfaces)

3%

Other 3%

Bronze and brass

8%

Tinning 1%

Alloys (misc.)

31%

Figure S1. Pie chart showing major end uses of tin as a 

percentage of total consumption in the United States in 2014. 

Data are from Anderson (2016a, table 3). misc., miscellaneous

Tin Prices and Pricing

During the 20th century, the pricing of tin was 

somewhat “regulated” by a series of agreements structured 

between producer countries and consumer countries 

dating back to 1921 (Carlin, 2013). The International Tin 

Council, an association of leading world tin-producing and 

-consuming countries, was established in 1956 to try to 

stabilize both tin prices and the supply of and demand for 

tin until the organization was disbanded in 1989. The Association of Tin Producing Countries (ATPC) was chartered 

on August 16, 1983, to perform activities complementary 

to those of the International Tin Council. The five charter 

members of the ATPC were Bolivia, Congo (Kinshasa), 

Indonesia, Malaysia, and Thailand, with Australia and 

Nigeria joining later. The ATPC was disbanded in 2001. 

Figure S2 shows the average annual prices of tin metal 

from 1970 to 2010 (Carlin, 2013).
Geology S5

Dollars per pound

4

2

6

8

12

14

10

Year

1970

16

0

1975 1980 1985 1990 1995 2000 2005 2010

Tin Figure S2

EXPLANATION

1992 dollars

Price

Significant events affecting tin prices since 1970

1956–85 International Tin Agreements (a continuous 

series of complex, global, 4-year pacts)

1973–80 Rampant inflation

1981–82 Sharp recession

1980–2010 Increased toxicity issues with lead, leading to 

increased tin consumption as a lead 

replacement

2002–08 A period of world steel company (tinplate 

producers) consolidations and mergers that 

often resulted in lower available world 

tinplate and increased tin pricess

2002–09 Tin mine and smelter production declined in 

Indonesia owing to Government closures of 

illegal tin smelters, smuggling, and severe 

weather

2008–09 Global financial crisis

Figure S2. Graph showing the average annual prices of tin metal from 1970 to 2010. Both the nominal (unadjusted) 

prices and real prices (in 1992 constant dollars) are shown. Graph is from Carlin (2013).

Geology

Geochemistry

Tin has an atomic number of 50 and an atomic weight of 

118.69. It has two oxidation states: +2 and +4; the +4 oxidation 

state occurs in the mineral cassiterite. Tin is the 49th most 

abundant element, and it has the largest number of stable 

isotopes (10) in the periodic table of elements. The average 

abundance of tin in Earth’s crust is low at about 2.3 parts per 

million (ppm) compared with 70 ppm for zinc, 60 ppm for 

copper, and 14 ppm for lead. Tin is both siderophile and, to a 

lesser extent, lithophile; thus, it has an affinity for both iron 

and oxygen and is associated with boron, fluorine, lithium, 

and rubidium in the late-stage differentiates of granites or their 

extrusive equivalents. In pegmatitic settings, tin is associated 

with beryllium, boron, lithium, niobium-tantalum, tungsten, and 

the rare-earth elements (Rose and others, 1979; Carlin, 1985).

According to the study of tin-bearing and tin-barren 

granites in northern Nigeria by Olade (1980), highly evolved 

biotite granites are genetically and spatially related to tin 

mineralization (also Falconer, 1912; Williams and others, 

1956; MacLeod and others, 1971; Bowden and Kinnaird, 

1978). Beus (1969) also showed that statistically significant 

variations are found between granites of nonstanniferous 

regions and granites of stanniferous regions, as well as 

among biotite granites in different areas of Russia. Part of the 

explanation for this association lies in that, in less differentiated granites, tin is locked up in early crystallized silicate 

minerals, such as fayalite, hornblende, and riebeckite. This 

conclusion is confirmed by the high contents of tin and other 

elements, such as niobium, in these rocks, particularly the 

peralkaline riebeckite granites (Bowden and Van Breemen, 

1972). In contrast, biotite crystallized late within the biotite 

granites; consequently, tin and other elements were able to 

be concentrated into the volatile-rich fraction of the melt 

and subsequently in exsolved mineralizing fluids. Alexsiyev 

(1970) and Bowden and Van Breemen (1972) suggested that 

the enrichment of tin and related elements in tin-bearing 

granites is connected with albitization and the trace-element 

overprinting by post-magmatic fluids. Albitization is a product 

of subsolidus recrystallization (hydrothermal alteration) and 

is accompanied by the enrichment of the magma in trace 

elements, such as niobium, rubidium, and tin, and its depletion 

in barium, strontium, and zirconium. This pattern of traceelement association may be attributed to the development 

of an alkali-volatile-rich phase during the evolution of the 

granitic magmas.

Tin-bearing granites tend to show a similar pattern of 

enrichment in trace elements. According to Bowden and 

Kinnaird (1978), this enrichment may be the result of a 

continuum of ore-forming processes: that is, an early dispersed 

phase of mineralization associated with alkali-volatile-rich, 

early post-magmatic fluids, and a later phase of fracturecontrolled mineralization associated with hydrothermal 

solutions from which sulfides are commonly deposited. This is 

consistent with the occurrence of both types of mineralization 

within the same intrusive body (Olade, 1980).

Three mechanisms have been proposed by Heinrich 

(1995) for cassiterite precipitation. The first is acid neutralization by feldspar hydrolysis, which takes place during the 

formation of alteration assemblages. The second is vapor 

separation (boiling), which effectively removes hydrogen 

chloride (HCl) and hydrogen gas (H2) and causes cassiterite 

deposition. The third is mixing of hot, saline magmatic fluid 

with a nonmagmatic fluid.
S6 Critical Mineral Resources of the United States—Tin

Mineralogy

Cassiterite is a member of the rutile group of metal 

oxides that includes rutile (TiO2), pyrolusite (MnO2 ), and 

plattnerite (PbO2 ); it is tetragonal and belongs to the space 

group P42/mnm. The structure of the conventional unit cell 

of SnO2, which is depicted in figure S3, is characterized by 

two lattice parameters, a and b, and an internal parameter u

associated with the O (oxygen) atomic position. There are 

two formula units per primitive unit cell, with the threefoldcoordinated oxygen atoms forming distorted octahedral 

configurations around the tin atoms (Floriano and others, 

2010). Members of the rutile group show little variation in 

composition and little tendency to form isomorphous series 

between the species. Cassiterite has an adamantine, greasy, 

submetallic luster (fig. S4); it ranges from transparent to 

opaque, and can be black, yellow, brown, red, or white. It has 

a Mohs hardness of 6 to 7, imperfect cleavage ({100} and 

{110}), and a density range of from 6.98 to 7.01 grams per 

cubic centimeter (g/cm3

). Optically, it has very high relief 

and weak pleochroism and is anomalously biaxial (Hudson 

Institute of Mineralogy, 2016).

A radially fibrous colloform variety of cassiterite (“wood 

tin”) (fig. S5) can form by secondary processes in the zone 

of oxidation of tin deposits. It can also form as a hypogene 

mineral both in deep-seated veins and, in particular, as 

colloform masses associated with rhyolite flow-dome deposits 

of lavas and associated tuff (Palache and others, 1944).

Deposit Types

Placer Tin Deposits

Placer deposits have traditionally been an important 

source of tin; in 2012, they accounted for about 70 percent 

of the world output of cassiterite concentrates, principally 

from, in order of output, Malaysia, Indonesia, China, Burma 

(also referred to as Myanmar), Thailand, Nigeria, and Congo 

(Kinshasa) (Taylor, 1979; Carlin, 2014). Because cassiterite 

is both heavy and chemically inactive, it is concentrated in 

residual placers that form over or adjacent to bedrock source 

areas where weathering and erosional processes remove lighter 

rock materials. Furthermore, gravity assists in downslope 

movement of heavy minerals that are released by weathering 

to form eluvial placers (that is, deposits that remain relatively 

close to the primary deposit from which they are derived). 

The richest placers are found in stream deposits where flowing 

water has concentrated heavy minerals generally derived 

from residual or eluvial placers. Ocean beach sands may also 

contain placer accumulations of tin, such as those mined 

offshore in Indonesia and Thailand. No extensive placer 

deposits have been found in the conterminous United States 

(Carlin, 1985).

b a2

a1

Sn

O

Tin Figure S3. Add colors to caption. Figure S3. Ball-and-stick model of part of the crystal structure 

of cassiterite (SnO2). Oxygen (O) atoms are shown in red, and tin 

(Sn) atoms, in gray. Courtesy of Ben Mills.

Figure S4. Photograph of cassiterite crystals. Copyright © 

Dakota Matrix Minerals; used with permission.

Figure S5. Photograph of wood tin cassiterite. Copyright © 

R. Weller, Cochise College; used with permission.
Geology S7

Granite-Related Tin Deposits

Granite-related tin deposits are another important source 

of tin. Lehmann (1990) traced the history of research on 

granite-related deposits. According to Lehmann, the close relationship between tin deposits and granites has been well known 

since the mid-19th century. Research since then has mainly 

refined the understanding of the chemistry of the associated 

granites, the tectonic environment, and the mechanisms of 

deposition and localization of tin mineralization.

As discussed by Lehmann (1990), Alexander von 

Humboldt (1823) introduced the term “tin granite” as opposed 

to “normal granite.” The first comprehensive treatise on the 

geology of tin and the first scientific theory of magmatichydrothermal ore formation in general was developed by 

Élie de Beaumont (1847). He concluded that tin ore deposits 

are associated with granites and are located preferentially 

in apical portions of granites and their immediate country 

rocks (fig. S6). Daubrée (1841) had already observed that 

tin deposits in the areas of Cornwall (United Kingdom), 

Bretagne (France), and Erzgebirge (Germany) are always 

confined to a zone extending less than 500 m from the granite 

contact. Élie de Beaumont (1847) further concluded that 

tin ore deposits are commonly associated with individual 

granite bodies extending out of larger batholiths, and that the 

Decreasing

temperature

Greisen

Stockscheiders

0 100 200 300 METERS

0 500 1,000 FEET

Vein

Alteration

halo

Granite

Late stage

microgranite

Tin Figure S6

ore occurs in rocks particularly rich in quartz, tourmaline, 

and fluorine-bearing minerals and that formed as a result 

of alteration of the original host rock (that is, greisens). Tin 

granites are anomalous in texture—for example, unidirectional 

solidification textures (USTs, also known as comb quartz 

layers or stockscheiders), graphic intergrowths, and pegmatitic 

dikes and pods (Shannon and others, 1982; White and others, 

1981; Kirkham and Sinclair, 1988); they are also anomalous 

in composition—for example, “specialized, S-type granites” 

enriched in Rb, Cs, Li, Th, U, Nb, Ta, and W and depleted in 

Sr, Eu, Ba, Ti, Co, and Ni (Lehmann, 1990).

Classifications of Granite-Related Tin Deposits

Granite-related tin deposits can be classified according 

to the following general types. This information is modified 

primarily after S.S. Smirnov from Magak’yan (1968), Taylor 

(1979, p. 45–76), and Lehmann (1990).

Stanniferous pegmatites.—“These pegmatites are of the 

quartz-microcline type, with albite, muscovite, [and commonly] 

spodumene, topaz, and tourmaline. Cassiterite is irregularly 

disseminated throughout the pegmatite body and commonly 

in areas where secondary processes have taken place” 

(Taylor, 1979, p. 46–47). Examples of stanniferous pegmatites 

include the Jābal as Silsilah pegmatites in Saudi Arabia 

Figure S6. Schematic vertical 

section across a typical hydrothermal 

mineralized granite cupola showing 

salient features of a shallow graniterelated tin-mineralized system. As a 

tin-bearing granitic magma intrudes 

into existing crustal rock and cools, 

residual metal-rich fluids left over 

from the crystallization of the primary 

igneous minerals are injected into 

voids and fissures within the granite 

and surrounding host rock, forming 

bodies with high concentrations of tin. 

After Cérný and others (2005).
S8 Critical Mineral Resources of the United States—Tin

(Kamilli and Criss, 1996) and the Wodgina and Kamativi 

pegmatites in Australia, which are the largest known examples 

(Daniel J. Kontak, Department of Earth Sciences, Laurentian 

University, Sudbury, Ontario, Canada, written commun., 2014).

Quartz-cassiterite veins, stockworks, and greisens that 

formed in close genetic association with the same granitoid 

intrusions to which stanniferous pegmatites are related.—This 

type of deposit is characterized by intense development of 

greisenization with abundant formation of muscovite- and 

topaz-rich alteration zones that host the cassiterite mineralization; in addition, these alteration zones host wolframite, 

bismuthinite, molybdenite, arsenopyrite, tantalo-columbite, 

and uranium and copper minerals. This type of mineralization is the most widely developed type of granite-related tin 

deposit. Large primary deposits occur rarely, but very rich 

placer deposits formed by weathering and erosion occur 

more frequently. These types of deposits (chiefly the placers) 

account for 60 to 70 percent of the world production of tin 

(Taylor, 1979, p. 46). Examples of these types of deposits 

include the Geevor and Wheal Jane deposits in Cornwall, 

United Kingdom (Dines, 1956; Garnett, 1963; Hosking, 

1964; Rayment and others, 1971); the Herberton deposit in 

Queensland, Australia (Blake and Smith, 1970); the East 

Kemptville deposit in Nova Scotia, Canada (Richardson and 

others, 1982, 1990; Richardson, 1988; Halter and others, 1996, 

1998); and the Chojlla deposit in Bolivia (Lehmann, 1990).

Sulfide-cassiterite deposits.—This type of deposit 

commonly occurs independently of the pegmatite and 

vein-stockwork-greisen deposit types discussed above. 

Sulfide-cassiterite deposits are more commonly known as 

tin porphyries; many have been mined more for their silver 

than for the tin they contain. They are high-to-intermediatetemperature deposits and are related to hypabyssal and 

near-surface granitoid intrusions. Hydrothermal minerals 

commonly present include cassiterite, tourmaline, chlorite, 

pyrrhotite, chalcopyrite, sphalerite, galena, and arsenopyrite. 

The most notable examples are the Llallagua (Turneaure, 

1935; Dietrich and others, 2000) and the Chorolque deposits 

in Bolivia (Sillitoe and others, 1975; Grant and others, 1980) 

and the San Rafael deposit in Peru (Kontak and Clark, 2002); 

other examples may include the Ardlethan deposit in New 

South Wales, Australia (Clarke and others, 1985; Ren and 

others, 1995), the Zeehan deposit in Tasmania, Australia 

(McAndrew, 1965; Taylor, 1979), and the Akenobe deposit 

in Japan (Shimizu and Kato, 1991; Taylor, 1979).

Skarn and carbonate-sulfide replacement deposits in 

which cassiterite is associated with arsenopyrite, scheelite, 

pyrrohotite, marmatite, and chlorite that form in the alteration 

zones (that is, skarns) adjacent to intrusions.—Examples of 

these types of deposits include the Aberfoyle and Rossarden 

deposit (Collins, 1981) and the Renison Bell deposit 

(Patterson and others, 1981; Sutphin and others, 1990, and 

references therein) in Tasmania, Australia; the Cuomolong 

deposit in China (Hou and others, 2007); and the Lost River 

deposit in Alaska (Sainsbury, 1964; Dobson, 1982).

High-temperature hydrothermal deposits of “wood-tin” 

cassiterite in association with specular hematite and topaz 

in fluorine-rich, lithophysal high-silica rhyolite domes, lava 

flows, and associated pyroclastic rocks.—Cassiterite appears 

to have formed largely by precipitation from magmatically 

derived fluids. Redistribution of cassiterite by cooler hydrothermal fluids precipitated wood tin. The Mexican tin-bearing 

rhyolites have mineralogical and chemical characteristics that 

are similar to the topaz rhyolites of the Western United States 

(Burt and others, 1982; Christiansen and others, 1986). 

Small deposits of this type occur in Mexico (Foshag and 

Fries, 1942; Huspeni and others, 1984; Ruiz, 1988), Russia 

(Korostelev and others, 2009), and in Nevada (Fries, 1942) 

and New Mexico (Fries, 1940; Rye and others, 1990) in the 

United States (Foshag and Fries, 1942; Huspeni and others, 

1984; Ruiz, 1988).

Hydrothermal tin-silver deposits connected with 

near-surface stocks of rhyolite with high concentrations of 

silver.—Examples of this type of deposit occur in Cerra Rico 

de Potosí, Bolivia (Sillitoe and others, 1975, 1998), and 

Pirquitas, Argentina (Ross, 1941).

Similarities With Granite-Related Deposits of Other Metals

Although recent literature reviews of ore-deposit models 

have emphasized the differences among molybdenum, tin, and 

tungsten deposits related to highly differentiated granites, it is 

also important to emphasize that there are striking similarities 

among these deposits. Detailed studies by Kamilli and others 

(1993) and Kamilli and Criss (1996) of tungsten-tin greisen and 

tin greisen deposits in Saudi Arabia have documented similarities among wolframite-bearing quartz vein deposits, porphyry 

tungsten-molybdenum deposits, disseminated tin-greisen 

deposits in cupolas, and Climax-type molybdenum deposits. 

(The late stage so-called quartz-sericite-pyrite [QSP] veins 

at the Henderson and Climax Mines in Colorado are steeply 

dipping. They contain quartz, wolframite, topaz, and fluorinebearing mica and garnet—a chemical and mineralogical 

greisen assemblage—even though the rock is fine-grained.) 

Following Burt (1981), it is useful to think of these graniterelated molybdenum, tin, and tungsten deposits as a superclass 

of porphyry-greisen deposits. These deposits are characterized 

by close spatial, temporal, and genetic association with highly 

differentiated peraluminous porphyritic granites and by 

predominantly parallel or subparallel greisen-bordered and 

sheeted vein-granites that either crosscut the intrusion or are 

peripheral to it and that contain different amounts of cassiterite 

and wolframite (Kamilli and Ganster, 1995).

The porphyry-greisen deposits of this superclass are all 

located in tectonic settings that are intracratonic, extensional, 

and late orogenic, and were formed during a changeover 

from compressional to extensional tectonics. These deposits 

can also be closely related to tin- and tungsten-rich porphyry 

deposits associated with collision-related granites in Europe 

and Asia (Laznicka, 2006).
Resources and Production S9

Effect of Cupola Size and Shape on Tin-Greisen 

Mineralization

A principal control of tin-greisen mineralization is the 

degree of magmatic differentiation in the different phases of 

the parent granite’s apical cupolas. Nonetheless, differentiation 

alone does not explain why one intrusion does not produce 

any tin-greisen mineralization of consequence, although it 

is chemically almost identical to other intrusions that are 

mineralized. A possible reason is the differing geometries of 

the intrusions. One intrusion may be a small, plug-like cupola, 

whereas another intrusion may form a more extensive, flattopped mass. The concentration of greisen- and ore-producing 

hydrothermal fluid may be the result of the focusing effect of 

small cupola diameters.

An excellent example of this relation is the Mount Emmons 

molybdenite deposit in Colorado. The richest mineralization is 

at the apex of a 500-meter (m)-diameter cupola located 3,000 m 

above sea level, where the small cross-sectional area of the 

cupola and virtually vertical side contacts with the country rocks 

allowed accumulation of hydrothermal fluids within a small 

volume at the highest elevation of the intrusion. Lower grade 

ore zones occur on an inflection in the gently dipping flank of 

the same intrusion about 2 kilometers (km) to the north at an 

elevation of about 2,400 m. Even lower grade mineralization 

is found about 1 km farther north, where the contact is even 

flatter, at an elevation of about 1,000 m (White and others, 1981).

In the Jābal as Silsilah tin deposit in Saudi Arabia, one 

intrusion of peraluminous granite in a ring structure forms 

a body with an area of 3.5 square kilometers (km2

) at the 

surface. Thus, ascending fluids in this intrusion, although 

trapped under an impermeable aplitic cap, would be spread 

out over an area of at least 3.5 km2. By contrast, the small 

cupolas inside the ring have an outcrop area of only 0.06 km2. 

Given mean tin concentrations of 58 and 145 ppm for these 

two intrusions, respectively, the total amount of tin is actually 

more than an order of magnitude greater in the larger intrusion 

in the ring for any given depth than for an equivalent depth 

in the small cupolas. Using the mean tin content of 58 ppm 

and the 0.06-km2 area of the small cupolas, a magma column 

that is only 245 m high would be necessary to produce the 

measured resources in the small cupolas of 1.2 million metric 

tons grading 0.19 weight percent tin (2,280 metric tons of 

contained tin). This calculation assumes for simplicity that 

100 percent of the tin was scavenged from the melt.

The relation between the volume of magma and the size 

of the resulting deposit also appears to apply to other deposits. 

The Akash granite (which is about 100 km2 in area) in Saudi 

Arabia contains zones of strong greisenization on its western 

flank that are about one order of magnitude larger than the 

greisen zones at Jābal as Silsilah. The Jabal Akash deposit, 

however, is not of potential economic importance because 

the Akash granite is not sufficiently differentiated to have 

given rise to a commercial-grade tin deposit (Kellogg and 

Smith, 1985). The Akash granite is chemically similar to the 

“precursor granites” as described by Tischendorf (1977).

Tectonic Settings of Granite-Related Tin Deposits

The tectonic settings of tin-bearing granites are relatively 

well understood and limited in variation. Tin deposits and their 

associated igneous rocks occur in continental settings worldwide. More than 25 percent partial melt during crustal anatexis 

is probably a precondition for large-scale magma segregation. 

Magma segregation and intrusion into an upper crustal level 

is favored by a brittle lithosphere in crustal extension zones 

(Lehmann, 1990). The tectonic settings of tin granites and 

porphyries, according to Lehmann (1990), are as follows:

1. Post-orogenic magmatism in continental collision belts; 

for example, the Permian-Triassic tin belt in Southeast Asia (Mitchell, 1977; Beckinsale, 1979), and the 

Hercynian tin provinces in Western Europe and Central 

Europe (Mitchell, 1974; Holder and Leveridge, 1986).

2. Internal regions of active continental margins (backarc regions), such as Tertiary tin porphyries in Bolivia, 

Mexico, and eastern Siberia, Russia (Sillitoe, 1976), 

and Cretaceous tin granites in Thailand and Burma 

(Beckinsale, 1979).

3. Intracratonic anorogenic and continental rift zones; for 

example, the Triassic and Miocene tin granites in Peru 

and northern Bolivia (Kontak and others, 1985), the 

Cretaceous tin granites in southern China (Chen, 1989), 

the Cretaceous tin granites in Nigeria (Bowden and Kinnaird, 1984), and the Precambrian Bushveld granites in 

South Africa (Hunter, 1973).

Resources and Production

The locations of selected major tin deposits and districts 

in the world, by deposit type, are shown in figure S7. Additional data on the major tin deposits and districts in the world 

are in table S2 at the back of this chapter.

About 80 percent of the world’s identified tin resources 

occur in unconsolidated secondary or placer deposits in 

riverbeds and valleys or on the sea floor. The largest concentration of both onshore and offshore placers is in the extensive 

tin belt of Southeast Asia, which stretches from China in the 

north, through Thailand, Burma, and Malaysia, to the islands 

of Indonesia in the south. Most of the remainder (about 

20 percent of the world’s identified tin resources) occurs in 

primary veins, stockworks, and greisens in close association 

with silicic, peraluminous granites. The percentages of 

identified tin resources in primary versus placer deposits are 

estimations because many studies do not distinguish between 

the two types of deposits, which are commonly found together 

in the same districts. Furthermore, tin placer deposits are 

almost always found in close proximity to the granites from 

which they originate.

Although most tin is obtained from mining tin ores, 

cassiterite is also found in association with ores of tungsten, 
S10 Critical Mineral Resources of the United States—Tin 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0°

10°20°30°40°50°60°

80°

70°

70°

80°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

1

2

3 4

6

5

7

8

9 10

11 13

1412

1615

17

18

23

19

20

22

21

24

25

26 27

28

29 30

31

32

33

34

35

36

37

4142 43

39

40

38

47–54

46

55

44

57

56 58 60 59

61

62

63

84

64–65

66 68

67

69

70

71727374

75

76

45

7877

79

80

81

82

83

85

86

898887

90

91

92

93

96

97

94 95

98

99

100

EXPLANATION

Hydrothermal and 

placer/alluvial

Deposit type

HydrothermalPlacer/alluvialTailings

Tin Figure S7. Figure S7. World map showing locations of major tin deposits and districts in the world, by deposit type. The numbers are keyed to the identification numbers listed in table S2.
Resources and Production S11

tantalum, and lead. Minor quantities of tin are recovered as 

byproducts of the mining of these metals. Byproduct tin from 

massive sulfide deposits includes production from the Neves 

Corvo Mine in Portugal, which is the leading producer of 

tin in Europe, and the Kidd Creek Mine in Canada, which 

is a long-time producer of tin from stannite (Daniel Kontak, 

Department of Earth Sciences, Laurentian University, 

Sudbury, Ontario, Canada, written commun., 2014).

Compared with commercially viable deposits of copper, 

lead, zinc, nickel, and bauxite, tin deposits are generally small. 

Most tin deposits fall in a grade range of 0.1 to 1 percent tin, 

have a resource tonnage range of from 1 million to 100 million 

metric tons, and have between 10,000 and 100,000 metric tons 

of tin content (fig. S8).

In 2016, world reserves of tin totaled 4,700,000 metric 

tons (table S3; C.S. Anderson, U.S. Geological Survey, written 

comm., December 16, 2016). The distribution of world mine 

tin reserves in 2016, by country, is shown in figure S9.

Total world tin mine production in 2015 was 

289,000 metric tons of contained tin. The distribution of world 

mine production, by country, is shown in figure S10. China 

was the leading producer of tin (accounting for 110,000 metric 

tons, or 38 percent of world output), followed by Indonesia 

(18 percent), Burma (12 percent), Brazil (9 percent) and Bolivia

and Peru (7 percent each) (Anderson, 2017).

Table S3. Tin reserves of the world in 2016, in metric tons 

of contained tin.

[Data are from Anderson (2017)]

Country Reserves

(metric tons)

Australia 370,000

Bolivia 400,000

Brazil 700,000

Burma 110,000

China 1,100,000

Democratic Republic of the Congo 

(Congo [Kinshasa])

110,000

Indonesia 800,000

Malaysia 250,000

Peru 100,000

Russia 350,000

Thailand 170,000

Vietnam 11,000

Other countries 180,000

World total 4,700,000

100,000 t Sn

1,000,000 t Sn

10,000 t Sn

1,000 t Sn

1

0.1

0.01

10 100 1,000 10,000

Grade, in weight percent Sn

Deposit tonnage, in million metric tons

0.1 1 100,000

10

100

Hydrothermal

Hydrothermal and 

 placer/alluvial

Placer/alluvial

Tailings

EXPLANATION

Tin Figure S8 Figure S8. Graph showing grades and resource tonnages of major tin deposits in the 

world, by deposit type or source. Diagnonal lines are isolines of contained tin. Data are 

from table S2 and the references listed therein. Sn, tin; t, metric ton
S12 Critical Mineral Resources of the United States—Tin

Russia

7.5%

Australia

8.0%

Brazil

15%

Indonesia

17%

China

24%

Bolivia

8.6%

Other countries 4.1%

Thailand 3.7%

Burma 2.4%

Peru 2.2%

Malaysia

5.4%

Congo (Kinshasa) 2.4%

Figure S9. Pie chart showing percentage of world tin 

reserves in 2016, by country. Percentages have been rounded 

to two significant figures and do not add to 100. The world 

total reserves were 4.7 million metric tons. Data are from 

Anderson (2017). Congo (Kinshasa) is a short-form name for 

Democratic Republic of the Congo.

Bolivia

20,000 t

Brazil

25,000 t

Burma

34,300 t

China

110,000 t

Indonesia

52,000 t

Peru

19,500 t

Rwanda 2,000 t

Australia 7,000 t

Vietnam 5,400 t

Malaysia 3,800 t

Congo (Kinshasa) 6,400 t

Nigeria 2,500 t

Other countries 1,100 t

Figure S10. Pie chart showing estimated world mine 

production of tin in 2015, by country and amount (in metric 

tons [t]). The estimated world total was 289,000 t. Data are 

from Anderson (2017). Congo (Kinshasa) is a short-form 

name for Democratic Republic of the Congo.

Globally Important Mining Areas

Major tin deposits are confined to a comparatively small 

number of areas. More than one-half of the world’s tin ore 

is mined in the great tin belt of Southeast Asia. Significant 

deposits also occur in South America.

Australia.—Granite-related tin deposits are present 

through much of Australia and include Neoarchean tantalumtin pegmatite fields in western Australia and Paleozoic 

granite-related hydrothermal tin-tungsten deposits in the 

Tasmanides of eastern Australia (Solomon and Groves, 

2000). A variety of deposit types are present in eastern 

Australia, including distal skarn tin deposits, proximal tin 

skarns, greisens, and quartz-cassiterite-wolframite systems. 

Eastern Australia has been a significant tin producer, with 

more than 683,000 metric tons of tin produced to the end of 

1986 (Solomon and Groves, 2000). Tin was first discovered 

in Australia at Mount Bischoff in Tasmania in 1871.

Bolivia.—Bolivia has long been a significant producer 

of tin from primary deposits in the so-called Bolivian tin belt. 

The Bolivian tin belt extends for 1,100 km through the length 

of the Eastern Cordillera of Bolivia and is generally less 

than 100 km wide (Urquidi-Barrau, 1989). It extends from 

southern Peru, where tin is produced from the San Rafael 

Mine, to northern Argentina, where the Pirquitas Mine has 

been the major producer. In addition to tin, deposits in the 

Bolivian tin belt have produced important amounts of tungsten, silver, zinc, lead, antimony, bismuth, and gold. Most of 

the deposits in the northern segment of the belt are generally 

related to plutonic rocks of Late Triassic-Early Jurassic 

and Oligocene age, whereas those in the southern segment 

are related to subvolcanic porphyry stocks of Miocene age 

(Urquidi-Barrau, 1989). The northern tin-tungsten deposits 

occur as fissure-filling veins along fault and fracture systems 

in Paleozoic sedimentary rocks peripheral to granitic plutons 

and within these igneous bodies. The southern deposits are 

associated with high-level porphyry stocks and eruptive 

complexes with several of the largest tin deposits in Bolivia 

(Llallagua/Catavi, Huanuni, San José/Oruro, and Cerro Rico) 

occurring at or near the margins of the large Morococala and 

Los Frailes volcanic fields. It is suspected that additional 

large tin deposits occur buried beneath the volcanic fields 

(Urquidi-Barrau, 1989).

The Bolivian tin deposits are mostly confined to high 

altitudes in the rather inaccessible regions of the central 

Andes. As a result, access at times poses formidable 

infrastructure and logistical challenges. The primary 

hard-rock tin deposits are now relatively low grade, and 

Bolivian mines operate at the highest costs of all the world’s 

primary tin mines (Urquidi-Barrau, 1989). In addition, the 

ores in the southern segment of the Bolivian tin belt have 

relatively high smelting and refining costs owing to their 

complex mineralogy.

Brazil.—In 2015, Brazil was the world’s fourth-ranked 

producer of mined tin (Anderson, 2017). Almost 90 percent 
Resources and Production S13

of the country’s production is obtained from the Pitinga Mine, 

which is located about 300 km northeast of Manaus in the 

Mapuero region of the State of Amazonas (table S2; fig. S7). 

The Pitinga Mine produced about 230,000 metric tons of 

tin between 1982, when the mine opened, and January 2001 

(Lenharo and others, 2003). Tin is produced mainly from 

alluvial deposits in and around two Paleoproterozoic granitic 

complexes, the Àgua Boa and Madeira igneous complexes. 

These complexes are composed of a variety of granitic rocks, 

including fine to coarse grained, equigranular to porphyritic 

rapakivi granite, biotite granite, topaz granite, and albite 

granite (Lenharo and others, 2003). In the Àgua Boa complex, 

primary tin mineralization occurs in cassiterite-topaz-micaquartz greisens as well as disseminated cassiterite in biotite 

and topaz granites. In the Madeira complex, cryolite-zirconcassiterite-pyrochlore-columbite-tantalite-xenotime mineralization occurs disseminated within the roof zone of the albite 

granite. Tin is produced from a saprolite developed on the 

albite granite.

Burma (Myanmar).––The tin deposits in Burma lie along 

the Thailand-Burma border and form the northern end of the 

Southeast Asian tin belt, a north-south elongate zone about 

2,800 km long and 400 km wide, extending from Burma 

and Thailand to Peninsular Malaysia and the Indonesian Tin 

Islands (Schwartz and others, 1995). The Southeast Asian tin 

belt consists principally of Jurassic-Triassic biotite granites 

that host mostly simple tin-tungsten vein deposits and lesser 

stockwork, replacement, pegmatite, and breccia deposits 

(Schwartz and others, 1995). The tin deposits in Burma are 

similar to those in Thailand and Malaysia with both rockhosted and placer deposits related to biotite granites. The tin 

belt in Burma extends from the Byingyi-Mawchi district in the 

north about 1,200 km south to the Mergui Islands. Although 

small placer tin deposits are present in Burma, the mountainous terrane is not favorable for the occurrence of large 

placer deposits such as those found in Thailand, Malaysia, 

and Indonesia (Sainsbury, 1969b).

China.—China has been the leading tin-producing 

country since 2003, when it surpassed Indonesia. The main 

producing area is the Gejiu district in Yunnan Province, which 

is the largest primary tin district in the world; it contains about 

300 million metric tons of tin ore with an average grade of 

1 percent tin (Cheng and others, 2013). Tin mineralization 

occurs in greisen, skarn, strata-bound cassiterite-sulfide 

(mostly oxidized), and vein deposits in extensive hydrothermal 

systems centered on shallow Late Cretaceous granitic cupolas 

(Cheng and others, 2013). Geologic characteristics, metal 

zoning patterns, and new geochronologic data all support that 

the tin-polymetallic deposits are genetically related to the 

nearby granite intrusions.

Principal tin smelter production from China was 

148,000 metric tons in 2012 (Tse, 2013, p. 8.14). According to 

Carlin (2014, p. 77.2), the Guangxi Autonomous Region and 

Yunnan Province accounted for 53 percent of China’s mine 

production of tin in 2011, but production was substantially 

lower in 2012 owing to mine closures caused by cadmium 

contamination and drought-related water shortages.

Indonesia.—Indonesian tin production comes mainly 

from placer deposits both onshore and offshore of the islands 

of Bangka, Belitung (Billiton), and Singkep in the Java Sea. 

Bangka Island is the major source of placer tin and is produced 

with seagoing dredges (Sainsbury, 1969b). The largest hardrock tin deposit is the Klappa Kampit Mine on Belitung Island 

and consists of replacement bodies composed of cassiterite, 

magnetite, and sulfide minerals in shale (Sainsbury, 1969b).

Malaysia.—Historically, Malaysia furnished as much as 

55 percent of the world’s annual tin production (Sainsbury, 

1969b). The tin deposits in Malaysia are part of the Southeast 

Asian tin belt and consist of both hard-rock and placer 

deposits, with placers being the dominant producers. The 

tin deposits extend through Malaysia in two north-trending 

belts that coincide with two mountain ranges that consist 

of Jurassic-Triassic batholiths and stocks of biotite granite 

and related rocks which intrude older sedimentary and 

metamorphic rocks (Sainsbury, 1969b). The two most 

productive tin-mining areas in Malaysia have been the Kinta 

Valley in the State of Perak (Rajah, 1979) and the Selangor 

River in the State of Selangor. Between 1876 and 1950, the 

Kinta Valley alone produced 1.2 million metric tons of tin 

(Sainsbury, 1969b).

Portugal and Spain.—An 800-km-long belt of tin- and 

tungsten-bearing deposits that extends southeastward through 

Portugal and Spain is associated with dominantly porphyritic 

biotite and biotite-muscovite granites (Sainsbury, 1969b). 

The deposits consist of veins, stockworks, and alluvial and 

eluvial placers. As in other tin districts, major tin deposits 

tend to occur within and near the borders of granite intrusions, whereas tungsten deposits tend to occur farther from 

the contacts in surrounding rocks. The Neves-Corvo deposit 

in Portugal—one of the largest and richest massive sulfide 

deposits in the Iberian pyrite belt—is unique because of its 

very high copper and tin grades. The Corvo orebody, which 

accounts for 95 percent of Neves-Corvo tin production of 

about 40 metric tons per year, has reserves of 20 million 

metric tons with grades of 5.20 percent copper, 3.07 percent 

zinc, and 0.30 percent tin (Relvas and others, 2006).

Thailand.—The modern commercial production of tin in 

Thailand began in about 1870, but it was not until the early 

part of the 20th century that the country became a significant 

producer. The tin deposits in Thailand, both hard-rock and 

placer, are a continuation of the deposits in Malaysia and are 

part of the Southeast Asian tin belt. As elsewhere in Southeast 

Asia, placer deposits have been the major tin producers in 

Thailand. The principal tin placers are located on Phuket 

Island and near Ranong to the north (Sainsbury, 1969b). Tin 

production from Phuket Island has mostly been from alluvial 

stream-channel deposits and their submarine extensions. In 

northwestern and western Thailand, hard-rock fissure vein, 

stockwork, and pegmatite tin deposits also contain wolframite 

and sulfide minerals (Sainsbury, 1969b).
S14 Critical Mineral Resources of the United States—Tin

United Kingdom.—Cornwall, England, has a long 

history as a leading producer of tin. Cornish tin deposits have 

been worked extensively for many centuries, dating back 

to at least Roman times; there is, in fact, strong evidence of 

tin working in Cornwall from the early Bronze Age (about 

2100 to 1500 B.C.; Bancroft and Weller, 1993). Although the 

mines of Cornwall produced more than 2 million metric tons 

of tin during their operation, they are now largely exhausted 

(Sainsbury, 1969b). In 1860, more than 300 mines, employing 

100,000 workers, were in operation in Cornwall, but by 1964, 

only 2 mines were still operating (Sainsbury, 1969b). The last 

mine, South Crofty, closed in 1998.

The tin deposits of Cornwall are principally zoned 

greisens and replacement deposits developed in and around 

Permian biotite-muscovite, tourmaline-bearing granites 

intruded into deformed and metamorphosed upper Paleozoic 

marine mudstone-sandstone sequences (Jackson and others, 

1989). The tin deposits contain cassiterite and copper, arsenic, 

iron, tungsten, and zinc sulfide minerals. Because of metal 

zoning in the deposits, some mines produced principally 

copper ores; others produced lead-zinc ores; and others, 

tin, wolframite, and minor sulfide ores (Sainsbury, 1969b). 

Erosion of the primary tin deposits produced tin-bearing placer 

deposits that were the first source of tin in the region (Bancroft 

and Weller, 1993).

Tin Deposits in the United States

Tin deposits are quite rare in the United States and of 

relatively little economic value (Carlin, 2001). They are 

mentioned here because they are potential sources of tin in 

times of limited supply.

In 1948, a byproduct plant at the Climax Mine in 

Colorado was built to treat tailings from the molybdenite 

circuit to recover cassiterite and wolframite. In 1968, 

about 8,200 metric tons (about 18,000,000 pounds) of 

tungsten trioxide (WO3) and about 360 metric tons (about 

800,000 pounds) of cassiterite had been recovered (Wallace 

and others, 1968). The production of tin and tungsten ceased 

in 1982 when the Climax Mine closed (Maurice W. Ganster, 

Climax Mine, Climax Molybdenum Co., written commun., 

1981). The Climax Mine was reopened by Freeport-McMoRan 

Inc. in 2012 as an open pit operation with a new mill, but tin 

and tungsten were not being recovered.

One of the few primary hard-rock tin deposits in the 

United States is the Lost River tin-tungsten-fluorine deposit 

in the Seward Peninsula, Alaska. In the region of the northwestern Seward Peninsula, tin mineralization is related to at 

least nine Late Cretaceous, highly differentiated, fluorine-rich 

granites that intruded massive Ordovician-age limestones 

(Dobson, 1982). Other igneous rocks present in the region 

are diabase dikes, lamprophyre dikes, and alkali basalt. The 

deposits of the Seward Peninsula, which were first reported 

in 1900 by A.H. Brooks, have produced ore containing more 

than 2,200 metric tons of tin and constitute the only known 

domestic deposits of economic grade and size (Sainsbury, 

1964, 1969a, b). In most cases, the tin mineralization is 

evident primarily in placers that surround the exposed cores 

of the intrusions that host bedrock mineralization. The lodes 

were a major source of ore, however. The Lost River deposit 

is also enriched in F, W, Be, Zn, Pb, Cu, and Ag. Mineralization occurs in areas of intense greisenization of the intrusive 

granites and is characterized by the destruction of igneous 

textures and the replacement of feldspars and biotite by quartz 

with topaz, tourmaline, cassiterite, and sulfides, including 

galena, sphalerite, chalcopyrite, arsenopyrite, and stannite. 

Greisenization was apparently contemporaneous with a 

hydrous stage of alteration in the limestone; this alteration 

stage was characterized by fluorite-biotite ± hornblende with 

sulfides and cassiterite. Beryllium mineralization is hosted 

by fluorite-white mica veins. Beryllium-bearing minerals 

are also present and include chrysoberyl (BeAl2O4), euclase 

(BeAlSiO4(OH)), phenakite (Be2SiO4), and minor amounts of 

other beryllium-bearing minerals. Tin was precipitated prior to 

and contemporaneous with the deposition of sulfide minerals; 

the mobilization and redeposition of tin liberated from early 

skarn minerals by later alteration was an important process in 

the formation of the economic tin skarns. Early andradite garnet 

contains up to 6 weight percent SnO2, and idocrase and grossular contain 0.2 to 0.6 weight percent SnO2

 (Dobson, 1982).

Undiscovered Resources

Tin metallogenic provinces are quite well known; thus, 

most undiscovered tin deposits are expected to be located 

spatially close to known deposits or extensions of the same. 

Sainsbury and Reed (1973, p. 637) observed that “The worldwide geologic association of tin with acid granitic rocks and 

rhyolites is so well marked that speculation as to the likelihood 

of finding major tin resources in new geologic environments is 

without redeeming value.” Since then, however, Hudson and 

Reed (1997) have documented more than 100 tin occurrences 

in Alaska in several relatively recently discovered tin-mineralized areas, and the high-grade cassiterite ore in the NevesCorvo massive sulfide deposit of Portugal was a departure 

from conventional tin ore. Alaskan tin prospects occur from 

the central Alaska Range north to the Brooks Range and across 

interior Alaska, including the Yukon-Tanana upland, the Ruby 

geanticline, southwestern Alaska, and the Seward Peninsula 

(fig. S11). In all cases, the deposits are known or inferred to 

be associated with felsic igneous rocks that were generated 

by large-scale melting of continental crust. The widespread 

distribution of tin occurrences and their repeated development 

at different geologic times suggest that a large part of Alaska is 

a tin metallogenic province. Therefore, Alaska is prospective 

for future discoveries and eventual production, in the event 

that economic incentives are presented by market conditions 

(Hudson and Reed, 1997).
Exploration for New Deposits S15

ARCTIC OCEAN

BROOKS RANGE

Seward Peninsula

Norton Sound BERING SEA

Goodnews Bay

Shumagin Islands

Kodiak Island

PACIFIC OCEAN

Anchorage

Fairbanks

Juneau

Yukon Territory

British Columbia

Prince William Sound

Bethel

x x x

x x

x

x

x x

x

x x

x x

x

x

x

x

1 2

3

4

6

7

9

10

14

12

13

11

15

16

5

8

Yukon-Tanana upland Ruby Geanticline

300 KILOMETERS

0

0 100 200

100 200 300 MILES

Tin Figure S11

Yukon River

Kuskokwim River

GULF OF ALASKA

ALASKA RANGE

x

x

N

ALASKA

Principal tin deposit or occurrence

EXPLANATION

 1 Cape Mountain

 2 Potato Mountain

 3 Lost River (Cassiterite Dike, Tin Creek,

 Bessie Maple, and Idaho prospects)

 4 Kougarok

 5 Sleitat

 6 Bismark Creek

 7 Win

 8 Won

 9 Cosna

10 Tofty

11 Sithylemenkat

12 Lime Peak

13 Circle Hot Springs

14 Sheep Creek

15 Coal Creek

16 Groundhog Basin

x Selected other occurrence

Figure S11. Map of Alaska showing the locations of selected tin deposits or occurrences. Modified from Hudson and 

Reed (1997).

Exploration for New Deposits

Historically, prospecting for tin has been carried out by 

the time-honored methods of panning, drilling, trenching, and 

assaying. As rich deposits were exhausted and the costs of 

exploration rose, the search for tin required more sophisticated 

methods. Geophysical and geochemical analyses have been 

employed to cover large areas more rapidly, isolating areas of 

possible tin deposits so that drilling can be more effective and 

less costly (Carlin, 1985).

Exploration for tin deposits is relatively simple, as the 

metal is invariably associated with highly differentiated, 

peraluminous granites or in downstream, downslope placer 

or alluvial deposits. As an example, the discovery of the East 

Kemptville deposit in Nova Scotia was based on a glacial 

till survey, as the deposit area is completely buried under 

several meters of till, and the discovery hole (EK–79–1) was 

centered on the highest tin content in tills (Daniel J. Kontak, 

Department of Earth Sciences, Laurentian University, 

Sudbury, Ontario, Canada, written commun., 2014).

The tin content of residual soils and stream sediments and 

of vegetation has been used successfully as a guide to tin in 

the bedrock. The tin content of granitic rocks in the Malaysian 

alluvial tin fields is on the order of 60 ppm compared to the 

normal granitic background of 3 ppm. Boron, fluorine, lithium, 

rubidium, and thorium, as well as tin, are all enriched in latestage granitic differentiates, and have been cited as local (but 

not universally applicable) ore guides (Rose and others, 1979).

At the Silsilah ring complex in Saudi Arabia, which 

contains peralkaline and peraluminous granites, the tin 

deposits are associated with the strongly greisenized peraluminous granites and pegmatites that have USTs containing 

euhedral quartz crystals, measuring up to 2 m, which have 

their c axes oriented perpendicular to the top of the granite 

cupolas (Kamilli and Criss, 1996). Nevertheless, strong 

albitization is associated with nonmineralized metaluminous 
S16 Critical Mineral Resources of the United States—Tin

granites that are intermediate in composition between the 

peralkaline and peraluminous granites. In summary, the 

strongest field indicators of economic tin deposits associated 

with granites are the presence of mica-bearing (biotite and 

[or] muscovite) pegmatites and especially zones of greisen 

alteration. These associations have been observed in other tin 

provinces worldwide.

The small, uneconomic deposits of cassiterite and wood 

tin that occur in Nevada, New Mexico, and Mexico are related 

to topaz rhyolite flow-dome complexes that are indicators 

of topaz granites that may host economic deposits of tin (or 

tungsten) in greisens or skarns at depth. Examples of these 

deeper formed deposits include those at East Kemptville, 

Nova Scotia, Canada, and at the Anchor Mine in Tasmania, 

Australia (Christiansen and others, 1986).

Trace-element characteristics of biotite granites can 

be useful in identifying individual intrusive complexes that 

are potentially ore bearing. Frequency distributions of tin in 

tin-bearing and mineralized granites are positively skewed and 

bimodal, with a relatively large proportion of values exceeding 

25 ppm. Elemental concentrations and relationships of the 

lithophile elements, especially Rb, Li, K, Ba, Sr, Zr, Nb, Th, 

and Mn, are the most reliable chemical indicators of oreforming processes and tin-bearing potential (Olade, 1980).

Finally, of course, the best places to find new orebodies 

or extensions of existing ones are in areas of known mining 

districts or metallogenic belts. Only four well-defined regions 

account for 80 percent of the cumulative historic tin mine 

output. They are as follows: (a) the Southeast Asian tin belt 

(Burma, Thailand, Malaysia, and Indonesia); (b) the Central 

Andean tin belt (southeastern Peru, Bolivia, and northwestern 

Argentina); (c) the south China tin belt; and (d) the Cornwall 

tin province (United Kingdom; Lehmann, 1990).

Environmental Considerations

Sources and Fate in the Environment

Tin commonly occurs in one of two oxidation states in 

weathering environments: Sn2+ and Sn4+. As summarized by 

Séby and others (2001, and references therein), Sn(IV) is 

relatively insoluble and most commonly occurs in the mineral 

cassiterite, although dissolved Sn(IV) oxyhydroxide complexes 

(for example, SnO(OH)3 and SnO(OH)2) have been observed. 

Sn(II) may form dissolved complexes with hydroxide, chloride, 

sulfate, phosphate, carbonate, and organic species, or, more 

commonly, with solid Sn(II) oxide and sulfide minerals. 

Dissolved Sn(II) species may sorb to the surfaces of clays 

and manganese and iron oxide minerals. Sn(IV) forms the 

basis of organo-tin compounds with the general formula 

RnSnX4–n (n=1 to 4), where R is a hydrocarbon group and 

X is an anionic species, such as chloride, fluoride, oxide, 

hydroxide, carboxylate, or thiolate (Anger, 2004).

Examples of natural concentrations of tin in rocks, 

soils, waters, and air are given in table S4. The tin content 

in soils generally ranges from less than 0.1 to 5 ppm (which 

is equivalent to milligrams per kilogram), with higher tin 

concentrations occurring in more organic soils (KabataPendias and Pendias, 2001). The mean tin concentration at a 

depth of 20 centimeters in soils in the Western United States 

is 0.9 ppm (Shacklette and Boerngen, 1984), but topsoil 

concentrations can be as high as 8.6 ppm (Smith and others, 

2005) in the Eastern United States. Topsoil in Sweden can 

be equally as concentrated in tin, but deeper soils exhibit a 

maximum tin concentration of nearly 3.0 ppm (Eriksson, 

2001). In North Atlantic seawater, tin concentrations range 

from below the detection limit (that is, less than 0.2 nanogram per liter [ng/L]) to about 2.0 ng/L, with the highest 

concentrations occurring at depths of from 50 to 200 m 

(Byrd and Andreae, 1986b). Dissolved and colloidal (that 

is <0.45 micrometer [μm]-sized) concentrations of tin in 

rivers range from 0.1 to 8.4 ng/L, whereas particulate (that 

is, >0.45 μm) concentrations of tin in rivers range from 

0.1 to 109 ng/L (Byrd and Andreae, 1986c). Considered 

together, the average concentration of tin in the suspended 

sediment of world rivers is 4.6 ppm (Viers and others, 2009). 

Tin occurs naturally in the atmosphere as part of mineral 

dust particles, and ranges from 0.3 to 53 picograms per cubic 

meter (pg/m3) in more remote areas of the world (Byrd and 

Andreae, 1986a).

Mining and industrial activities can lead to above-background concentrations of tin in the environment (table S4). 

The common association of tin with organic material, clays, 

and metal oxide minerals increases the likelihood that tin will 

become enriched in soils, but above-background concentrations of tin in soils are not well documented. Tin mobilized 

from rocks tends to concentrate in particulate matter and sediments in aquatic environments. Byrd and Andreae (1986c) 

studied Tasmanian and European rivers and, as noted above, 

found that the concentration of tin in suspended particulates 

was greater than dissolved concentrations of tin (table S4). 

Likewise, unfiltered water samples from industrial European 

countries contain tin concentrations that are orders of 

magnitude greater than dissolved tin concentrations observed 

in more remote regions (table S4; Byrd and Andreae, 1986c; 

Reimann and de Caritat, 1998; Neal and Robson, 2000).

Much of the previous research on tin in aquatic environments focused on organo-tin compounds. Tributyltin (TBT) 

has been a common additive in paint for ships and barges 

to prevent the growth of mollusks, algae, and barnacles 

(Evans and others, 1995). As a result, concentrations of 

butyltin compounds in North American and European harbors 

were observed to range from below the detection limit to 

1,500 ng/L tin (Dirkx and others, 1993, and references therein). 

Because TBT also negatively affects nontarget organisms, 

such as mussels and fish, many countries banned the use of 

TBT-based paints on boats of less than 25 m in length in the 

late 1980s and early 1990s (Evans and others, 1995).
Environmental Considerations S17

Table S4. Tin concentrations in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency. Units of measure: cm, centimeter; ng/L, nanogram per liter; ng/m3, nanogram 

per cubic meter; pg/L, picogram per liter; pg/m3, picogram per cubic meter; ppm, part per million; μg/L, microgram per liter; μm, micrometer]

Environment and (or) location Tin concentration Unit Notes Reference(s)

Rocks

Upper continental crust 5.5 ppm Average Taylor and McLennan (1995)

Bulk continental crust 2.5 ppm Average Taylor and McLennan (1995)

Lower continental crust 1.5 ppm Average Taylor and McLennan (1995)

Soils

Conterminous United States 0.2 to 3.7 ppm Range for 0 to 5 cm depth Smith and others (2005)

Conterminous United States 1 to 1.9 ppm Range for O horizon, if present Smith and others (2005)

Conterminous United States 0.3 to 8.6 ppm Range for A horizon Smith and others (2005)

Conterminous United States 0.2 to 4.2 ppm Range for C horizon Smith and others (2005)

Western United States 0.90 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Eastern United States 0.86 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Topsoil, Sweden 0.4 to 8.6 ppm Range; mean is 1.8 ppm Eriksson (2001)

Subsoil, Sweden 0.4 to 3.0 ppm Range; mean is 1.5 ppm Eriksson (2001)

Proposed DOE benchmark 50 ppm Screening benchmark for 

terrestrial plants

Efroymson and others (1997)

Soil quality guideline 5 ppm Canadian agricultural soil 

guideline

Canadian Council of Ministers 

of the Environment (2007)

Waters

North Atlantic <0.2 to 1.2 ng/L Range; mean is 0.3 ng/L Byrd and Andreae (1986b)

Seawater, Sargasso Sea <0.2 to 1.0 ng/L Range; mean is 0.4 ng/L Byrd and Andreae (1986b)

Seawater, The Bahamas <0.2 to 1.9 ng/L Range; mean is 0.8 ng/L Byrd and Andreae (1986b)

World rivers 4.57 ppm Average suspended sediment Viers and others (2009)

North American rivers 0.5 to 8.4 ng/L Range of dissolved and colloidal 

(<0.45 μm)

Byrd and Andreae (1986c)

North American rivers 0.5 to 109 ng/L Range of particulates (>0.45 μm) Byrd and Andreae (1986c)

South American rivers 0.5 to 4.0 ng/L Range of dissolved and colloidal 

(<0.45 μm)

Byrd and Andreae (1986c)

European rivers 0.6 to 5.5 ng/L Range of dissolved and colloidal 

(<0.45 μm)

Byrd and Andreae (1986c)

European rivers 4.4 to 18 ng/L Range of particulates (>0.45 μm) Byrd and Andreae (1986c)

River water, Tasmania, 

Australia

0.1 to 0.6 ng/L Range of dissolved and colloidal 

(<0.45 μm)

Byrd and Andreae (1986c)

River water, Tasmania, 

Australia

0.1 to 2.0 ng/L Range of particulates (>0.45 μm) Byrd and Andreae (1986c)

River water, Azores, Portugal 1.1 to 1.4 ng/L Range of dissolved and colloidal 

(<0.45 μm)

Byrd and Andreae (1986c)

River water, Azores, Portugal 1.1 to 1.4 ng/L Range of particulates (>0.45 μm) Byrd and Andreae (1986c)

River water, Senegal 2.1 ng/L Dissolved and colloidal (<0.45 μm) Byrd and Andreae (1986c)

River water, eastern 

United Kingdom

100 to 600 ng/L Range of means for 18 rivers in 

rural and industrial areas

Neal and Robson (2000)

Stream water, Germany <0.01 to 1.3 μg/L Unfiltered; median is 0.03 μg/L Compiled by Reimann and 

de Caritat (1998)

Stream water, eastern India 0.55 to 4.12 μg/L Dissolved (<0.2 μm); median is 

1.2 μg/L

Compiled by Reimann and 

de Caritat (1998)
S18 Critical Mineral Resources of the United States—Tin

Table S4. Tin concentrations in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency. Units of measure: cm, centimeter; ng/L, nanogram per liter; ng/m3, nanogram 

per cubic meter; pg/L, picogram per liter; pg/m3, picogram per cubic meter; ppm, part per million; μg/L, microgram per liter; μm, micrometer]

Environment and (or) location Tin concentration Unit Notes Reference(s)

Waters —Continued

Stream water, Romania 5 to 170 ng/L Unfiltered; median 5 ng/L Compiled by Reimann and 

de Caritat (1998)

Lake water, Norway <0.04 to 3.52 μg/L Unfiltered; median is 0.061 μg/L Compiled by Reimann and 

de Caritat (1998)

Rainwater, United States 775 pg/L Volume-weighted mean Byrd and Andreae (1986a)

Groundwater, southern Norway <0.005 to 2.33 μg/L Unfiltered; median is <0.005 μg/L Compiled by Reimann and 

de Caritat (1998)

Proposed EPA benchmark 2,700 μg/L Tier II secondary acute value Suter and Tsao (1996)

Proposed EPA benchmark 73 μg/L Tier II secondary chronic value Suter and Tsao (1996)

Air

Polluted air 1.5 to 800 ng/m3 None Compiled by Reimann and 

de Caritat (1998)

Northern Atlantic Ocean 2.11 ± 1.97 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

Pacific Ocean 1.64 ± 2.14 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

Europe 397 ± 267 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

Cape Grim, Australia 0.26 ± 0.16 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

The Bahamas 52.6 ± 22.1 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

New Zealand 1.55 ± 0.76 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

Tallahassee, Florida, 

United States

134 ± 139 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

Rhode Island, United States 850 ± 71 pg/m3 Mean and standard deviation Byrd and Andreae (1986a)

The primary inputs of tin to the atmosphere are through 

mining and ore processing, combustion of fuel oil and coal, 

and production and use of tin products. Total anthropogenic 

emissions of tin to the atmosphere were estimated to be 

6.14 million grams in 1983 (Nriagu and Pacyna, 1988). Tin 

concentrations in air over urban and industrial regions range 

from 134 to 850 pg/m3 (Byrd and Andreae, 1986a). The 

general range of tin concentrations reported for “polluted 

air” is orders of magnitude higher at 1.5 to 800 nanograms 

per cubic meter (Reimann and de Caritat, 1998, and 

references therein).

Tin continues to be an important component in industrial 

society, so the need for tin continues to increase. Unlike 

many metals, tin recycling is relatively efficient. The recycled 

content (or the proportion of scrap) used in tin production is 

between 10 and 25 percent, and the fraction of tin in discarded 

products that get recycled is greater than 50 percent (Graedel 

and others, 2011, figs. 3 and 4). The factors that contribute to 

the economic and technologic feasibility of tin recycling are 

as follows: (a) tin is used in large quantities, and (b) it occurs 

in alloys in relatively pure form, making it easier to remelt and 

recover (Reck and Graedel, 2012).

Mine Waste Characteristics

Tin deposits are mined by open pit and underground 

methods. The mine waste generated is generally considered 

to be the material that originates and accumulates at the 

mine site that has no current economic value (Lottermoser 

and Ashley, 2006) and includes both solid and liquid waste. 

The character of mine waste varies according to the deposit 

type and the methods used to extract the ore. Most available 

information about tin mine waste relates to lode deposits. 

Mining of the Cerro Rico de Potosí lode deposits in Bolivia 

since 1545 has generated large amounts of mine waste. Two 

mine waste impoundment failures at Bolivia’s Porco Mine in 

1996 released approximately 235,000 cubic meters of tailings 

and fluid into the Río Pilaya (Pilaya River) (Hudson-Edwards 

and others, 2001). Tailings at the Jumna mill in northern 
Environmental Considerations S19

Queensland, Australia, come from various mines in the 

Irvinebank lode tin deposits in the area. The mill closed in 

1989, and the tailings repository on site contains an estimated 

150,000 t of tailings (Garrad and Bultitude, 1999).

The mineralogy of solid mine waste from tin mining is 

similar to that of the mined deposit, but the proportion of tin 

ore minerals is smaller. Tailings at the Porco Mine in Bolivia 

contain small amounts of cassiterite as well as other minerals, 

such as sulfides (arsenopyrite, galena, pyrite, sphalerite, 

and tetrahedrite), quartz, illite, and gypsum; trace metals in 

these tailings include Ag, As, Cd, Cu, Hg, Pb, Sb, Tl, and 

Zn (Hudson-Edwards and others, 2001). The Jumna mill 

tailings in Australia consist of quartz, microcline, plagioclase, 

muscovite, chlorite, kaolinite, sulfides (pyrite, chalcopyrite, 

galena, and arsenopyrite), magnetite, and small amounts of 

hornblende, beryl, fluorite, and expanding clays (Lottermoser 

and Ashley, 2006). Some of these tailings are cemented 

together with secondary hydrous ferric oxides (HFO) known 

as hardpans; the hardpan layers contain significantly higher 

As, Ca, Cd, Co, Cu, Fe, In, Mn, Mo, S, Th, U, Y, and Zn 

contents than the actual tailings because of mobilization and 

subsequent sorption of these elements by the HFO (Lottermoser and Ashley, 2006). Tin-bearing skarns and replacement 

deposits, such as those at Lost River (Alaska) and Dachang 

(China), are hosted in carbonate rocks. Gangue minerals in 

these deposits may include calcite and dolomite, may have 

generally lower sulfide mineral contents, and (or) may include 

the trace metals Ag, As, B, Be, Cs, Cu, F, Li, Pb, Re, Rb, W, 

and Zn (Hammarstrom and others, 1995).

In weathering environments, trace metals in tin deposits 

are expected to remain immobile under oxic, near-neutral 

pH conditions (5 < pH < 8), especially in the presence of 

iron-rich solids (Smith and Huyck, 1999). Under anoxic and 

(or) acidic conditions, however, minerals and associated trace 

metals may dissolve. In those deposits with sulfide minerals 

(for example, pyrite and arsenopyrite), generation of acid 

effluents is possible if sulfide dissolution is not balanced 

by the acid-neutralizing capacity of carbonate minerals 

(for example, calcite and dolomite). Dissolution of sulfide 

minerals releases metals and sulfuric acid, and the acidic pH 

values allow higher concentrations of metals to be dissolved 

in solution—potentially causing the environmental problem 

known as acid mine drainage (AMD). Mining of hydrothermal 

tin lode deposits increases the potential for AMD.

Little information about mine wastes from placer tin 

deposits is available, particularly for the placer deposits in 

Southeast Asia. Because placer deposits form as a result 

of chemical weathering and physical erosion of lode tin 

deposits, many of the potentially deleterious constituents 

originally present in lode deposits, such as sulfide minerals, 

are absent from or reduced in placer deposits (Sainsbury and 

Reed, 1973, p. 644). Heavy and less reactive minerals that 

commonly occur with cassiterite in tin placer deposits include 

columbite-tantalite, wolframite, ilmenite, monazite, zircon, 

and xenotime (Sutphin and others, 1990, p. 10). Monazite and 

zircon may include radioactive elements, such as thorium and 

uranium. Mine tailings in the Jos Plateau tin placer deposit 

in Nigeria contain above-background levels of thorium and 

uranium, and radioactivity dose rates measured in the tailings 

were approximately 100 times those measured in nearby 

control sites (Ibeanu, 2003). Based on a preliminary model 

of tin-gold placer deposits, Bliss and Menzie (1995) showed 

that grades of tin do not correlate with deposit volumes, and 

80 percent of the studied deposits had volumes of between 

8,800 and 91,000,000 cubic meters. These volumes can 

be viewed as the minimum volume of rock that would be 

disturbed during mining of a deposit.

Human Health Concerns

The Agency for Toxic Substances and Disease Registry 

(2005) provides a useful summary of the human toxicology 

of tin and tin compounds. The general public is most likely 

to be exposed to tin through consuming food directly from 

tin-lined cans (though currently many cans are lined with 

epoxy resins to protect against this), breathing tin-bearing 

dust, and eating seafood or using household products that 

contain organo-tin compounds. Occupational exposure to 

tin usually results from inhaling or touching tin-bearing 

dust at industrial and hazardous waste sites. The average 

human diet includes an intake of about 10 milligrams per 

day of tin (Anger, 2004). Exposure to oral doses of tin 

much greater than this amount may lead to stomach ache, 

anemia, and liver and kidney problems. Exposure to some 

organo-tin compounds can interfere with brain and nervous 

system function and, in severe cases, can cause death. Those 

exposed to extended inhalation of tin oxide—an issue mainly 

of concern to people who work in the tin industry—have 

a higher potential to develop stannosis, which is a mild 

disease of the lungs caused by inhalation of tin-bearing 

dust (Anger, 2004). Inorganic tin is poorly absorbed by 

the body, and no evidence exists for the carcinogenicity of 

metallic tin and tin compounds in humans (Agency for Toxic 

Substances and Disease Registry, 2005). Long-term exposure 

to certain organo-tin compounds, however, has been shown 

to cause cancer in animals (Smith and Huyck, 1999, and 

references therein).

Primary and secondary drinking water regulations for 

tin currently do not exist in the United States. Likewise, 

the World Health Organization has not set a recommended 

guideline for inorganic tin in drinking water because of the 

generally low concentrations of tin in aquatic systems and 

its minor toxicity to humans (World Health Organization, 

2004). The Occupational Safety and Health Administration 

has set an exposure limit of 2.0 milligrams per cubic meter 

(mg/m3 ) for inorganic tin dust (Occupational Safety and 

Health Administration, 2013a) and 0.1 mg/m3 for organo-tin 

dust (Occupational Safety and Health Administration, 2013b) 

in workplace air over an 8-hour workday.
S20 Critical Mineral Resources of the United States—Tin

Ecological Health Concerns

Many studies focus on the ecological impacts of tin in 

the environment. Because fish tend to be sensitive to low 

concentrations of dissolved metals, they are commonly useful 

indicators of contamination in aquatic systems. One of several 

useful endpoints used in toxicity tests is the concentration 

that leads to 50 percent mortality (LC50) after exposure to a 

substance for a certain amount of time. Chronic toxicity tests 

(that is, lower doses over longer times) with dissolved tin (as 

tin chloride, or SnCl2) and developing rainbow trout (Salmo 

gairdneri) revealed a mean LC50 value of 0.42 milligram 

per liter (mg/L) after 28 days of exposure (Birge and others, 

1980). Acute toxicity tests (that is, higher doses over shorter 

times) with tin and two marine fish species were attempted 

by Taylor and others (1985), but upon addition of dissolved 

tin chloride to seawater, most of the tin precipitated, so 

the highest concentration of dissolved tin that could be 

maintained was 0.03 mg/L, a concentration that caused no 

mortalities over 4 days. Chronic toxicity tests with juvenile 

rainbow trout (Oncorhynchus mykiss) and bis(tri-n-butyltin) 

oxide (TBTO) (at a concentration of 0.5 or 2.0 micrograms 

per liter [μg/L]) showed that after 21 days, weight gain 

was significantly lower in TBTO-treated specimens relative to control specimens, and internal tin accumulation in 

TBTO-treated specimens increased between 7 and 21 days 

(Triebskorn and others, 1994). Rüdel and others (2007) 

show that over the two decades following Germany’s ban 

on using organic tin compounds on small boats in 1989, 

concentrations of TBT and triphenyltin (TPT) decreased in 

the muscles of bream (Abramis brama) collected from river 

and lake samples. Some regions of the United States have 

adopted secondary acute and chronic screening benchmarks 

of 2,700 μg/L tin and 73 μg/L tin, respectively, for aquatic 

freshwater life, as recommended by the Oak Ridge National 

Laboratory (Suter and Tsao, 1996).

Relatively few studies focus on tin accumulation and 

toxicity in plants. Peterson and others (1976) surveyed the tin 

contents in different plant species in Malaysia and found that: 

(a) different plant species accumulate tin to varying levels 

(1.5 to 32 ppm dry weight) at control sites, and (b) plants 

near a tin smelter accumulated higher tin concentrations 

(8.4 to 2,165 ppm dry weight) than those plants in control 

areas. Experiments with relatively low concentrations of 

tributyltin chloride (TBTCl) and cells from three different 

vascular plant species revealed that tin concentrations 

within plant cells increased with increasing incubation time, 

which has implications for tin accumulation in the food 

chain (Caratozzolo and others, 2007). Some regions of the 

United States have adopted a soil screening benchmark of 

50 ppm tin for terrestrial plants, although the confidence in 

this benchmark is low because of sparse datasets (Efroymson 

and others, 1997). The Canadian agricultural soil quality 

guideline for tin is lower, at 5 ppm (Canadian Council of 

Ministers of the Environment, 2007).

Mine Closure

Most recent and new mining operations include closure 

plans that address issues related to the mine footprint. A 

mine’s footprint includes the waste left on site and in locally 

affected soil and water, as well as ecological impacts, such 

as habitat destruction and loss of biodiversity. Mining of 

both placer and lode tin deposits may generate mine waste 

piles. Wind-blown dust from unprotected mine waste piles 

may enter the atmosphere and surrounding soil and aquatic 

environments. In addition, these waste piles have the 

potential to become unstable. If using mine waste as backfill 

into dry mine workings is not an option, mine waste pile 

stability and dust-generating issues can often be addressed 

through grading and covering piles with vegetation.

Many lode tin deposits have the potential to generate 

AMD. For example, water draining from mine workings 

and tailings at the Ervedosa tin mine in Portugal displays 

a pH range of 3 to 6, and high dissolved metal and arsenic 

concentrations (Gomes and Favas, 2006). Common 

methods for treating AMD include active water treatment 

facilities or passive limestone-lined channels or constructed 

wetlands (Plumlee and Logsdon, 1999). The end result of 

both active and passive approaches is eventual precipitation of dissolved metals. Precipitated metals in passive 

wetland systems tend to be more stable under the prevailing 

anoxic conditions, whereas the metal-rich precipitates that 

result from active treatment facilities form a sludge that, 

if not responsibly disposed of, can cause environmental 

problems. When the potential for AMD exists at a mine site, 

a common preventative measure includes conducting water 

quality surveys before, during, and after mining.

Placer tin deposits tend to be mined by open pit and 

(or) dredging methods. Mining of alluvial placers in modern 

stream and riverbeds is less likely to increase dissolved 

salt loads but is more likely to increase the amount of 

sediment delivered downstream. This, combined with 

potential diversion of rivers and streams, may negatively 

affect downstream ecosystems. Many of the placer deposits 

located in Burma, Indonesia, Malaysia, and Thailand are 

located offshore below sea level. Some of these deposits 

have been mined, whereas others may become more 

economical to mine in the future (United Nations, 2004). 

Much of the processing of tin from offshore deposits takes 

place on land, so many of the same mine waste issues 

mentioned for onshore tin deposits also apply to offshore 

tin deposits. Additionally, most offshore placer tin deposits 

are mined by dredging methods, which have the potential 

to negatively affect benthic, midwater, and pelagic ecosystems. Such marine ecosystem disruption could, in turn, also 

negatively affect the fishing industry. Future developers will 

need to consider all environmental and social issues that 

could result from the mining of offshore placer tin deposits, 

as well as the relevant mitigation measures that address 

such issues.
Acknowledgments S21

Problems and Future Research

Supply Issues

Tin has not been mined or smelted in the United States 

since 1993 and 1989, respectively. Most tin deposits in the 

United States would be uneconomic to mine, even assuming 

higher prices for the metal. The only potentially economic 

deposits at present are the Lost River and Sleitat Mountain 

districts in Alaska (table S2, deposits 96 and 97; and fig. S11, 

deposits 3 and 5). The DOD recommends that material 

mitigation strategies for shortfalls include extra emergency 

purchases, substitutions, export reductions, Federal inventorying, and stockpiling. An additional strategy would be for 

the United States to develop its own (minimal) identified tin 

resources and (or) to explore for and find a major tin deposit. 

According to Sainsbury (1968), any future tin mining in the 

Seward Peninsula, Alaska, would have to mine ores with less 

than 2 percent tin.

Large-tonnage operations might be established on lowergrade disseminated ores, as occur in greisen-type deposits 

(for example, East Kemptville, Nova Scotia, Canada). To be 

profitable, such mines would probably also need to recover 

cassiterite, stannite, wolframite, and base-metal sulfides, as 

well as beryllium minerals and fluorite.

Research Issues

Whereas the mechanism of transport and deposition of 

tin is fairly well understood, the processes by which tin is 

incorporated into the parent magma at the points of magma 

generation and ascent are topics of investigation. As pointed 

out by Sainsbury and Reed (1973, p. 650–651), “alkali rocks 

in many parts of the world contain tin in amounts equal to or 

greater than the so-called tin-granites, yet no commercial tin 

lodes are ever associated with them. In a way, this fact refutes 

the possibility that the tin granites derive their tin from the 

surrounding rocks invaded by the magma or that tin deposits 

are formed by circulating meteoric waters.”

A promising advancement concerning research into the 

origin of tin deposits is the recent development of a reliable 

method for the analysis of tin isotopes in cassiterite (Haustein 

and others, 2010; Yamazaki and others, 2013). The ability 

to fingerprint cassiterite from various deposits or regions 

also shows great potential for identifying the source of tin in 

bronze archeological artifacts, a conundrum that has plagued 

archaeologists for many decades, if not centuries. Isotopic 

variations as large as 0.77‰ (per mil) in 124Sn/120Sn were 

observed in samples from Japan and China using multiple 

collector-inductively coupled plasma-mass spectrometer 

(MC–ICP–MS) techniques. The isotopic fractionation of 

all the sample studies was dependent on isotope mass. 

McNaughton and Rosman (1991) proposed the following three 

possible mechanisms for tin isotopic fractionation: (a) equilibrium fractionation between cassiterite and tin-sulfide minerals; 

(b) kinetic fractionation, which could take place during 

dissolution of tin in a source region and precipitation of tin 

in an ore deposit; and (c) multistage enrichment processes of 

tin from crustal material to ore, in which each step introduces 

kinetic fractionation. Yamazaki and others (2013) concluded 

from the samples that they studied, however, that the presence 

of tin-sulfide minerals does not play an important role in tin 

isotope fractionation; they also concluded that the temperature 

during ore formation was not the main cause of the tin isotope 

fractionation in the samples studied. Although the isotopic 

variation seen by Yamazaki and others (2013) was smaller 

than those observed in cassiterite from Europe by Haustein 

and others (2010), the mechanism of isotopic fractionation 

is still unclear. This is similar to the state of knowledge for 

stable isotopes of carbon, oxygen, silicon, and sulfur in the 

late 1940s and 1950s after the development of a precise mass 

spectrometer by Nier (1947). The technique is reliable and 

constantly being improved; the field of study is promising. 

Nevertheless, many tin deposits occur in regions where there 

is much shale or slate in the country rock, as in Cornwall 

and Bolivia. Stable isotope research using 124Sn/120Sn values 

and variations may go a long way towards improving understanding of the mechanism of tin isotopic fractionation.

Acknowledgments

A co-author of this report, James F. Carlin, Jr., died while 

this manuscript was being written. The first author (RJK) 

especially would like to acknowledge the help and advice 

he received from Jim in understanding tin—and metals in 

general—as commodities that form the basis of the world 

economy through the production of goods, rather than just 

an ore to be mined and shipped to the concentrator. The first 

time that Jim visited me in Tucson, Ariz., he came to my 

house and gave me a comprehensive tour and an entertaining 

lecture on all the materials that went into its construction, 

why and how they were used, and why some of the materials 

should not have been used by the builders. Jim also seemed to 

know everyone in the world associated with the tin industry. 

The table of world tin deposits contained in this paper would 

not have been nearly as complete as it is if Jim had not put 

me in contact with many companies that were developing 

small properties.

Mr. Peter Kettle of the International Tin Research Institute helped me find the addresses of these many small startup 

companies and sent me large amounts of “gray literature” 

wherein so much of what is happening in the mining world is 

hidden. He also provided me with much unpublished data of 

his own that he had compiled.

Finally, thorough and helpful reviews by Diana Kamilli, 

John DeYoung, Daniel Kontak, Klaus Schulz, and Eric Seedorff 

improved this paper greatly. Also, Keith Labay helped create 

the deposit location maps, and Suzanne Nicholson provided 

additional editorial comments after the technical review was 

complete.
S22 Critical Mineral Resources of the United States—Tin

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Table S2
S34 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State,

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or)

comments

Development 

status

Grade Sn 

(wt. %

unless

otherwise 

noted)

(decimal degrees 

WGS 84)

1 Pirquitas Mine Argentina Jujuy –22.72 –66.48 Hydrothermal Lode, subvolcanic 

vein

Underground Inactive since 

1991

n.d. 

2 Ardlethan Australia New South 

Wales

–34.33 146.85 Hydrothermal Lode, breccia Mine closed 

in 1986; 

reprocessing 

tailings in 

2013

Additional 

underground 

mining 

being considered in 

2013

0.420 

3 Doradilla Australia New South 

Wales

–30.33 146.35 Hydrothermal Skarn (F- and 

B-poor)

n.d. Preliminary 

economic 

analysis

0.280 

4 Taronga Australia New South 

Wales

–29.45 151.53 Hydrothermal Lode, stockwork/

sheeted veins

n.d. Prefeasibility 

study

0.150 

5 Maranboy/

Yeuralba

Australia Northern 

Territory

–14.53 132.79 Hydrothermal Fissure veins 

associated 

with quartzmica greisen, 

tourmalite, 

and topazite

Underground; 

operated from 

1913 to 1952

Exploration 2.030 

6 Mt. Wells Australia Northern 

Territory

–13.51 131.72 Hydrothermal 

and placer/

alluvial

Veins, stockworks, 

and alluvial

n.d. n.d. 0.970 

7 Collingwood Australia Queensland –15.77 145.25 Hydrothermal Lode, greisen n.d. Closed in 2008 0.730 

8 Gillian 

(Herberton 

area)

Australia Queensland –17.72 145.05 Hydrothermal Lode, skarn n.d. n.d. 0.810 

9 Mt. Garnet 

(Consolidated 

Tin Mines)

Australia Queensland –17.58 145.17 Hydrothermal Fe-rich skarn Open pit Prefeasibility 

study

0.600 

10 Mt. Garnet 

(MGT 

Resources)

Australia Queensland –17.58 145.17 Hydrothermal Fe-rich skarn Unknown n.d. 0.780 
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

n.d.

3.025 

(2013)

7.81 

(2013)

37.589 

(2013)

0.722 

(2013)

1.137 

(2013)

4.0 

(1984)

2.3 

(1982)

7.3 

(2013)

0.793 

(2013)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

About 

0.15 to 

0.25

n.d.

Schist

Granitic breccia

Marble, mica schist— 

Ordovician

Metamorphosed mudstone, 

siltstone, sandstone

Greywacke, sandstone, 

conglomerate, shale, 

basalt—Neoproterozoic 

to Middle Cambrian; 

sandstone, conglomerate, 

shale and siltstone—

Lower Cretaceous

Burrell Creek Formation: 

greywacke, siltstone, 

shale, biotite-sericite 

quartzite—Early Proterozoic (1862 ±3 Ma)

Greisen

Sedimentary rocks, basalt, 

chert, granite, granodiorite

Chillagoe formation 

(sandstone, siltstone, 

limestone, carbonates, 

chert, basalt)—Silurian

Chillagoe formation 

(sandstone, siltstone, 

limestone, carbonates, 

chert, basalt)—Silurian

n.d.

Biotite-rich, S-type 

granite

Leucoadamellite, 

greisenized 

quartz-feldspar 

porphyritic dikes 

(F- and B-poor)

Adamellite

Adamellite, quartz 

porphyry, and 

adamellite porphyry

Muscovite-biotite 

granite and hornblende granite, 

unfoliated

Granite

Granite

Granite

Granite

n.d.

417±2.5 Ma

Early Devonian

Late Triassic

n.d.

1804 ± 50

Permian

Permian

CarboniferousPermian

CarboniferousPermian

n.d.

n.d.

n.d.

n.d.

n.d.

U-Pb

n.d.

n.d.

n.d.

n.d.

n.d. n.d.

n.d. n.d.

13,000 x 40 to Stratform, near110 x 200 vertical

n.d. n.d.

Main lode— n.d.

3,950 x 2 

to 7.5 x?; 

Stannum 

King 

lode—1,000 

x 3.6 x?

2,400 x 300 x? n.d.

n.d. n.d.

n.d. n.d.

6,000 x 4 to Strata-bound, 

20 x ? tabular; 

three distinct 

orebodies

n.d. Strata-bound, 

tabular

Ross (1941); Velasco 

(1988); Sutphin and 

others (1990)

Ren and others (1995); 

Clarke and others 

(1985); Sutphin and 

others (1990); Australian 

Tin Resources Pty Ltd. 

(2016); Peter Kettle, 

International Tin Research Institute, written 

commun., April 8, 2013

Kwak (1987); Plimer 

(1984)

Offenberg (1982); Sutphin 

and others (1990); Peter 

Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Walpole (1958); Outback 

Metals Ltd. (2014a); 

Peter Kettle, International Tin Research Institute, 

written commun., 

April 8, 2013

Mookhey (1971); Outback 

Metals Ltd. (2014b); 

Peter Kettle, International Tin Research Institute, 

written commun., 

April 8, 2013

Krosch (1985)

Brown and others (1984); 

Sutphin and others 

(1990)

De Lacey (2012); Consolidated Tin Mines Ltd. 

(2013); Michael Hicks, 

Consolidated Tin Mines 

Ltd., written commun., 

July 22, 2013

Kettle (2012); Peter Kettle, 

International Tin Research Institute, written 

commun., April 8, 2013

Table S2 S35
S36 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

11

12

13

14

15

16

Cleveland 

(Aberfoyle 

and 

Rossarden)

Heemskirk 

(Queen Hill) 

(Zeehan tin 

project)

Mount Bischoff

Mount Lindsay

Renison-Bell

Rentails

Australia

Australia

Australia

Australia

Australia

Australia

Tasmania

Tasmania

Tasmania

Tasmania

Tasmania

Tasmania

–41.66

–41.88

–41.41

–41.75

–41.78

–41.78

147.75

145.32

145.52

145.36

145.78

145.78

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Tailings

Greisenized 

calcic skarns 

and carbonate 

replacements 

of calcareous 

sandstone and 

mudstone

Stockwork and 

strata-bound 

carbonate 

replacement 

(three separate 

orebodies)

Lode, carbonate 

replacement; 

alteration of 

quartz-orthoclase porphyry 

dikes

Greisenized 

calcic skarns 

and carbonate 

replacements 

of calcareous 

sandstone and 

mudstone

Lode, carbonate 

replacement

NA

Closed in the 

1970s

UG

Open pit

n.d.

n.d.

Tailings from the 

mill associated with the 

Renison-Bell 

deposit

In exploration 

(2013)

Prefeasibility 

study

n.d.

Bankable feasibility study

n.d.

n.d.

0.680

1.140

1.100

0.200

1.460

0.450
Table S2 S37

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

6.119 

(2013)

6.280 

(2013)

10.54 

(2013)

45 

(2013)

26.0 

(2013)

20.461 

(2013)

0.35

0.6

n.d.

ca. 0.10

n.d.

n.d.

Crimson Creek formation 

(interbedded, volcaniclastic turbiditic wacke, 

siltstone-mudstone, and 

tholeiitic basalt lava 

horizons and intrusive 

sills)—Neoproterozoic to 

Eocambrian; Sandstone, 

basalt, dolorite, argillite, 

chert, limestone, greywacke turbidite— 

Cambrian to Devonian

Quartzite, shale, sandstone, 

dolostone, Crimson 

Creek Formation 

(Interbedded, volcaniclastic turbiditic wacke, 

siltstone-mudstone and 

tholeiitic basalt lava 

horizons and intrusive 

sills)—Neoproterozoic 

and Cambrian

Dolomite, dolomitic 

shale, quartzite—PreCambrian; graywacke, 

mudstone, chert, 

basalts—Cambrian; 

quartz-orthoclase porphyry dikes and sills

Success Creek Group (fluviatile deposits—siliceous 

sandstone; siliceous 

siltstone, mudstone, 

dolomite and stromatolite clast-bearing, oolitic, 

chert breccia) and Crimson Creek formation 

(interbedded, volcaniclastic turbiditic wacke, 

siltstone-mudstone, and 

tholeiitic basalt lava 

horizons and intrusive 

sills)—Neoproterozoic

Dolostone, quartzite, siltstone, shale, sandstone

n.d.

I-type biotite granite; composed 

of adamellite to 

alkali-feldspar 

granite

I-type biotite 

granite; 

composed of 

adamellite to 

alkali-feldspar 

granite

Quartz-orthoclase 

porphyry dikes

I-type biotite 

granite; 

composed of 

adamellite to 

alkali-feldspar 

granite

n.d.

n.d.

350 Ma

Devonian

349±4 Ma

Devonian

Devonian

n.d.

K-Ar, 

Rb-Sr

n.d.

Rb-Sr 

(whole 

rock)

n.d.

n.d.

n.d.

100 x 100 x 

300

Severn—536 

x 223 x 

192; Queen 

Hill—446 x 

134 x 346; 

Montana— 

312 x 89 

x 38

n.d.

950 to 2,150 

long, 

15 to 35 

thick, 4.75 

cross-strike 

width of 

individual 

skarn layers

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Collins (1981); Blissett 

(1959); Mikyla Hickey, 

Venture Minerals Ltd., 

written commun., 

July 22, 2013

Sutphin and others (1990); 

Blight (2012); 

J.A. Anderson, Aberfoyle Ltd., unpub. data, 

1986; Stellar Resources 

Ltd., unpub. data, 2013; 

Peter Blight, Stellar 

Resources Ltd., written 

commun., July 1, 2013

Halley and Walshe (1995); 

Mikyla Hickey, Venture 

Minerals Ltd., written 

commun., July 22, 2013

Radonjic (2012); Mikyla 

Hickey, Venture Minerals Ltd., written commun., July 22, 2013

Sutphin and others (1990); 

R. Morland, Renison 

Associated Tin Mines, 

unpub. data, 1986; 

Mikyla Hickey, Venture 

Minerals Ltd., written 

commun., July 22, 2013

Peter Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013
S38 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

17

18

19

20

21

22

23

24

25

26

27

28

29

Greenbushes Australia

Colquiri tailings Bolivia

Central Bolivia

(Llallagua, 

(San Jose/

Oruro, 

Huanuni)

Bolivian placers Bolivia

Catavi tailings Bolivia

Southern Bolivia

(Cerro Rico, 

Chorolque)

Northern Bolivia

(Chojlla, 

Colquiri)

Mapuero tin Brazil

district (Pitinga Mine is 

the largest 

producer, accounting for 

about 90% 

of Brazil’s 

production)

Serra Branca Brazil

Brascan Group Brazil

(Rondonia 

tin province)

Mocambo Brazil

Rhodia-Espeng Brazil

Brumadinho Brazil

Group

Western 

Australia

La Paz 

Department

Oruro 

Department

Potosi 

Department

Potosi 

Department

Potosi 

Department

San Jose 

Department

Amazonas 

State

Goias State

NA

Pará State

Pará State

Rondonia 

and Goias 

States

–33.83

–17.38

–18.00

–18.00

–18.43

–19.00

–16.00

–0.75

–13.62

–9.00

–6.83

–6.03

–9.55

115.98

–67.12

–67.00

–65.00

–66.57

–66.00

–67.00

–60.12

–48.07

–63.00

–51.95

–53.72

–62.42

Hydrothermal

Tailings

Hydrothermal

Placer/alluvial

Tailings

Hydrothermal

Hydrothermal

Placer/alluvial

Hydrothermal 

and placer/

alluvial

Hydrothermal 

and placer/

alluvial

Hydrothermal 

and placer/

alluvial

Placer/alluvial

Hydrothermal 

and placer/

alluvial

Lode, pegmatite, 

placer

Sn-polymetallic 

vein

Lode, subvolcanic

Placer, alluvial

NA

Lode, subvolcanic

Lode, vein

Placer

Placer, lode, 

massive 

greisen

Placer, lode, 

massive 

greisen,

Placer, lode, 

massive 

greisen

Placer

Placer and lode

Surface placer

Tailings from the 

mill associated with the 

Colquiri tin 

mine

n.d.

n.d.

Catavi is a mill 

that served 

numerous 

mines in the 

Potosi district

n.d.

n.d.

Onshore dredge

n.d.

n.d.

n.d.

Gravel pump

n.d.

n.d.

Uncertain

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.1175 (lode); 

0.0268 

(placer)

0.600

0.550

0.021

0.350

0.550

0.780

n.d.

1.250

0.425

0.061

0.034

0.093
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

39.0 (lode); 

8.812 

(placer) 

(1985)

11.0 

(2013)

7.116 

(1982 and 

1984)

57.709 

(1982)

40.0 

(2013)

7.1176 

(1982)

8.667 

(1982)

0.500 

contained 

Sn (1987)

0.551 

(1982)

73.833 

(1982)

13.000 

(1982)

26.647 

(1982)

65.428 

(1982)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Pegmatite, laterized gravel 

and sand

Shale, sandstone, slightly 

metamorphosed— 

Devonian

Slate, metasediment, 

volcanics

Sediments

Slightly metamorphosed 

shale—Devonian

n.d.

Granodiorite, quartzite, 

shale

“Apogranite,” rhyolite 

and tuffs

Metasediments

Granite, Rapakivi granite, 

paleovalley sediments

Metasediments

n.d.

Granite, Rapakivi granite, 

paleovalley fill

Pegmatite

Only dikes seen; 

significance 

unknown

n.d.

n.d.

n.d.

n.d.

n.d.

“Apogranite,” 

rhyolite and tuffs

Greisenized granite

Greisenized granite

Greisenized granite

Granite

Granite

2.53 Ga

Miocene

Tertiary

n.d.

n.d.

Tertiary

Tertiary

“Apogranite” 

1.9 to 

1.5 Ga; 

rhyolite and 

tuffs—midProterozoic

Precambrian

Precambrian

Precambrian

Precambrian?

n.d.

n.d.

Stoping

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

3,000 x 100 x 

500

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Tabular

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Blockley (1980); Sutphin 

and others (1990)

Campbell (1947); Peter 

Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Fox (1971); Sillitoe and 

others (1975); Sutphin 

and others (1990)

Fox (1971); Sillitoe and 

others (1975); Sutphin 

and others (1990)

Peter Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Fox (1971); Sillitoe and 

others (1975); Sutphin 

and others (1990)

Fox (1971); Sillitoe and 

others (1975); Sutphin 

and others (1990)

Mareno (1986); Thorman 

and Drew (1988); Sutphin and others (1990)

Sutphin and others (1990)

Mareno (1986); Sutphin 

and others (1990)

Sutphin and others (1990)

Sutphin and others (1990)

Mareno (1986); Sutphin 

and others (1990)

Table S2 S39
S40 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

30

31

32

33

34

35

36

37

38

39

40

41

42

43

Best Group Brazil

C–75 Garimpo Brazil

Mawchi Mine Burma (also 

referred 

to as 

Myanmar)

Tenasserim Burma (also 

Valley, referred 

Southeast to as 

Asian tin Myanmar)

belt

Mount Pleasant Canada

East Kemptville Canada

Dachang China

Cuomolong China

Datongchang China

Diantan China

Gejiu district China

Lailishan China

Laopingshan China

Xiaolonghe China

Rondonia 

and Pará 

States

Rondonia 

State

Kayah 

Province

Taninthayi 

region

New Brunswick 

Province

Nova Scotia 

Province

Guangxi 

Province

Sichuan 

Province

Yunnan 

Province

Yunnan 

Province

Yunnan 

Province

Yunnan 

Province

Yunnan 

Province

Yunnan 

Province

–9.45

–9.90

18.82

12.00

45.43

44.10

24.83

30.50

25.53

25.65

23.33

24.92

25.00

25.45

–65.13

–63.43

97.17

99.00

–66.82

–65.68

107.83

99.37

98.73

98.43

103.33

98.27

98.37

98.43

Placer/alluvial

Hydrothermal 

and placer/

alluvial

Hydrothermal

Placer/alluvial

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Placer Gravel pump 

Placer, eluvial, n.d.

lode, greisen 

and vein

Lode, vein n.d.

Placer (alluvial) n.d.

Lode, subvolcanic n.d.

vein/porphyry

Lode, greisen n.d.

Lode, carbonate n.d.

replacement

Skarn n.d.

Vein, strata-bound, n.d.

skarn

Skarn n.d.

Lode, carbonate Surface-underreplacement, ground

some vein-type 

skarn

Vein, strata-bound n.d.

Greisen n.d.

Greisen n.d.

n.d.

n.d.

n.d.

n.d.

Prefeasibility 

study

Preliminary 

economic 

analysis

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.023

n.d.

0.070

n.d.

0.370

0.165

0.5 to 1.0

0.22 to 1.38

0.016 to 0.26

n.d.

1.000

0.63 to 1.58

n.d.

0.18 to 0.42
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

24.386 

(1982)

0.100 

contained 

Sn (1987)

0.7 

(1983) 

n.d.

15.2 

(2013)

56.0 

(1984)

50–100 

(1985)

11.5 

(2007)

<10.0 

(2007)

n.d.

100 

(1985)

42.6 

(2007)

n.d.

26,200 

(2007)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Paleovalley sediments

Granite, gneiss, basement 

rocks

Clastic and carbonate 

metasediments, granite

Eluvial and alluvial 

sediments, granite 

(tourmaline-rich)

Granite and quartz-feldspar 

porphyry

Granite and metasediments

Shale, marl, limestones, 

granite

Triassic sandstone

Dolomite, limestone, 

arenaceous mudstone, 

sandstone

Dolomitic limestone, 

arenaceous mudstone, 

sandstone

Limestone and dolomite, 

minor argillite, tuff and 

basalt, gneiss and schist, 

biotite granite

Metamorphosed feldspathic 

sandstone, argillaceous 

siltstone, dolomite

Sandy slate

Sandy slate

Greisen

Granite

Granite

Granite

Granite

Granite

Granite

Granite

Biotite granite

Biotite granite

Granite

Alkali-feldspar 

granite

Biotite granite

Biotite granite

Precambrian n.d.

1,150 t 925 Ma n.d.

Late n.d.

Mesozoic– 

Early 

Tertiary

Late n.d.

Mesozoic– 

Early 

Tertiary

340 to 330 Ma K-Ar

295±5 Ma Ar-Ar

97 Ma K-Ar

n.d. n.d.

Mesozoic n.d.

Mesozoic n.d.

Porphyritic K-Ar

biotite 

granite, 115 

to 84 Ma; 

Equigranular biotite 

granite, 80 

to 64 Ma

n.d. n.d.

Mesozoic n.d.

Mesozoic n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Length: to 442; 

thickness: 

1.40 to 8.58

n.d.

n.d.

n.d.

Length: 40 to 

600; width: 

0.89 to 23.8

n.d.

Length: 30 

to 300; 

thickness: 

1 to 20

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Lenticular, 

strata-bound

Vein, stratabound

Lens, stratabound

n.d.

n.d.

Vein, stratabound

Irregular

Bettencourt and others 

(1981); Sutphin and 

others (1990)

Thorman and Drew (1988); 

Sutphin and others 

(1990)

U Khin Zaw and Daw Khin 

Myo Thet (1983); Sutphin and others (1990)

Sutphin and others (1990)

Kooiman and others 

(1986); Sutphin and 

others (1990)

Richardson and others 

(1982); Sutphin and 

others (1990)

Premoli (1986); Sutphin 

and others (1990)

Hou and others (2007) 

Hou and others (2007) 

Hou and others (2007) 

Kwak (1987); Sutphin 

and others (1990)

Hou and others (2007) 

Hou and others (2007) 

Hou and others (2007) 

Table S2 S41
S42 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

44

45

46

47

48

49

50

Cinovec 

(Zinnwald)

Kivu mines

Abu Dabbab

Antonsthal 

district in 

the Westerzgebirge 

(Western Ore 

Mountains)

AuersbergWildenthal 

district in 

the Westerzgebirge 

(Western Ore 

Mountains)

Breitbrunn 

district in 

the Westerzgebirge 

(Western Ore 

Mountains)

ElbenstockSosa district 

in the Westerzgebirge 

(Western Ore 

Mountains)

Czech 

Republic

Democratic 

Republic 

of the 

Congo 

(Congo 

[Kinshasa]) 

(formerly 

Zaire)

Egypt

Germany

Germany

Germany

Germany

Teplice 

District

North Kivu 

Province

Al Bahr al 

Ahmar 

Governorate

Saxony

Saxony

Saxony

Saxony

50.73

–6.50

25.34

50.51

50.46

50.48

50.49

13.77

27.50

34.54

12.74

12.64

12.77

12.63

Hydrothermal

Placer/alluvial

Hydrothermal 

and placer/

alluvial

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Greisen

Placer

Skarn and greisen 

veins (primarily

a tantalum 

deposit); 

associated 

tin placers

Skarn ± schist 

replacement

Greisen veins

Skarn ± schist 

replacement

Greisen veins

n.d.

Gravel pump

Open pit planned

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Feasibility 

study

Exploration

Exploration

Exploration

Exploration

0.370

n.d.

0.09 

(0.13%, 

according 

to Roskill 

Information 

Services, 

Ltd., 2015)

0.190

0.300

0.250

0.260
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

28.1 

(2013)

0.265 

contained 

Sn (1986)

44.5 

(2013)

14.7

10.8

22.3

2.3

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Granite

Alluvium and eluvium, 

granite, metasediments

Quartz-bioite and quartz 

chlorite schists with 

intercalated beds of 

tuff and agglomerate, 

diabase—Neoproterozoic

Cambrian-Ordovician 

sediments, mainly pelitic, 

but with some limestone 

beds (now marble/ 

calc-silicate and skarn)

Cambrian-Ordovician 

sediments, mainly pelitic, 

but with some limestone 

beds (now marble/ 

calc-silicate and skarn)

Cambrian-Ordovician 

sediments, mainly pelitic, 

but with some limestone 

beds (now marble/ 

calc-silicate and skarn)

Cambrian-Ordovician sediments, mainly pelitic, but 

with some limestone 

beds (now marble/ 

calc-silicate and skarn)

Alkali-feldspar granite, rhyolite

Granite

Muscovite aplite 

leucogranite

Granite

Granite

Granite

Granite

330 to 295 Ma

Precambrian

Neoproterozoic

Hercynian

Hercynian

Hercynian

Hercynian

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

400 x 200 x 

480

2,500 x 1,300 

x variable 

thickness

500 x 500 x 

unknown 

depth

2,000 x 1,400 

x variable 

thickness

200 x 5,000 x 

unknown 

depth

Flat-lying, subhorizontal 

veins form 

regular 

domal structures that 

mimic the 

upper granite 

cupola 

contact

n.d.

Ellipsoidal and 

tabular

Veins

Veins

Strata-bound

Veins

Schwarz-Schampera and 

Herzig (2002); Porter 

GeoConsultancy Pty 

Ltd. (1995); Peter 

Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Angermeier and others 

(1974); Sutphin and 

others (1990)

Amin (1947); Sabet and 

others (1971); Sarin and 

others (1976); Roskill 

Information Services, 

Ltd. (2015); Peter Kettle,

International Tin Research Institute, written 

commun., April 8, 2013

Antony Truelove, Indo 

Gold Ltd., written commun., September 8, 2013

Antony Truelove, Indo 

Gold Ltd., written commun., August 8, 2013

Antony Truelove, Indo 

Gold Ltd., written commun., August 8, 2013

Antony Truelove, Indo 

Gold Ltd., written commun., August 8, 2013

Table S2 S43
S44 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

51

52

53

54

55

56

57

58

59

60

61

62

GeyerSouthwest

Gottesberg

HammerleinTellerhauser 

district in 

the Westerzgebirge 

(Western Ore 

Mountains)

Muhlleithen 

district in 

the Westerzgebirge 

(Western Ore 

Mountains)

Sauschwemme, 

Westerzgebirge 

(Western Ore 

Mountains)

P.T. Koba Tin

P.T. Tambang 

Tamah

Kelapa Kampit

P.T. Tambang 

Belitun

Pt Timah 

Singkep 

Mine

Akenobe

Suzuyama

Germany

Germany

Germany

Germany

Germany

Indonesia

Indonesia

Indonesia

Indonesia

Indonesia

Japan

Japan

Saxony

Saxony

Saxony

Saxony

Saxony

Bangka 

island

Bangka 

island

Belitung 

island

Belitung 

island

Singkep 

Island

Hyogo 

Prefecture, 

Honshu 

Island

Kyushu 

Island

50.62

50.42

50.45

50.41

50.44

–2.58

–2.00

–2.75

–2.67

–2.00

35.22

31.50

12.90

12.48

12.87

12.45

12.66

106.40

106.00

108.08

107.83

114.00

134.68

130.43

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Placer/alluvial

Placer/alluvial

Placer/alluvial

Tailings

Placer/alluvial

Placer/alluvial

Hydrothermal

Hydrothermal

Skarn

Greisen stockwork 

and breccia 

pipes

Skarn, greisen 

veins, replacements in schist 

below skarn

Breccia pipe

Alluvial

Placer

Placer

Lode, vein

Placer

Placer

Lode, polymetallic 

veins

Lode, pegmatite 

and quartz 

veins

Sublevel stoping 

or more 

selective 

mining being 

considered

Sublevel, open 

stoping 

planned

n.d.

n.d.

n.d.

Gravel pump, 

dredge

Gravel pump, 

onshore 

dredge

n.d.

Gravel pump, 

onshore 

dredge

Gravel pump, 

onshore 

dredge

n.d.

n.d.

Prefeasibility

Prefeasibility

Exploration

Exploration

Exploration

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.370

0.270

0.570

0.300

0.040

n.d.

0.220

1.670

0.013

n.d.

n.d.

1.250
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

11.6 

(2013)

42.1 

(2013)

25.3 

(2013)

0.75 

(“Soviet” 

estimates)

6.5 

(“Soviet” 

estimates)

0.0467 

contained 

Sn (1982)

2.259 

(1982)

1.390 

(1982)

23.381 

(1982)

n.d.

0.04 

contained 

Sn (1980)

0.448 

(1982)

0.150

0.150

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Cambrian-Ordovician 

schists, dolomite and 

calcite-dolomite 

sediments

Granite and granite 

porphyry

Cambrian-Ordovician 

sediments, mainly pelitic, 

but with some limestone 

beds (now marble/ 

calc-silicate and skarn)

Cambrian-Ordovician 

sediments, mainly pelitic, 

but with some limestone 

beds (now marble/calcsilicate and skarn)

Alluvium/stream sediment

Formed in current and 

ancient channels

Alluvial and eluvial 

sediments

Quartzite, sandstone, 

slate, granitoids

n.d.

Bedrock is alternating beds 

of sandstone and shale 

intruded by stocks 

that range from gabbro 

to granite

Basalt, basaltic tuff, 

phyllite

Granite, pegmatite, shale, 

and sandstone

Granite

Granite

Granite

Granite

n.d.

n.d.

n.d.

Granite

n.d.

Biotite granite

n.d.

Granite and 

pegmatite

Hercynian

Hercynian

Hercynian

Hercynian

Alpine orogeny(?) to 

Holocene

n.d.

Post-Triassic

Lower Triassic

n.d.

n.d.

Tertiary?

Tertiary

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

2,600 x 400 

x ?

800 x 400 x 

900

about 5,600 x 

0.5 x 2 to 3 

(seams)

30 x 30 x ≥ 

200

≥1,000 x 30

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Pipe

Strata-bound

Pipe

Alluvium

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Lehmann (1990); Kettle 

(2012 )

Beyschlag and others (1914);

Lindgren (1933); 

Lehmann (1990); 

Kettle (2012 )

Truelove, 2012; Antony 

Truelove, Indo Gold 

Ltd., written commun., 

September 8, 2013

Antony Truelove, Indo 

Gold Ltd., written commun., September 8, 2013

Antony Truelove, Indo 

Gold Ltd., written commun., September 8, 2013

Prijono (1986); Sutphin 

and others (1990)

Sutphin and others (1990)

Omer-Cooper and others 

(1974); Sutphin and 

others (1990)

Prijono (1986); Sutphin 

and others (1990)

Sainsbury (1969b)

Sato and Akiyama (1980); 

Sutphin and others 

(1990)

Saito and others (1960); 

Sutphin and others 

(1990)

Table S2 S45
S46 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

63 Syrymbet Kazakhstan North 

Kazakhstan 

Province

53.48 67.98 Hydrothermal Vein stockwork, 

skarn, endogreisen

n.d. Feasibility 

study

0.840

64 Malaysian 

dredges

Malaysia NA 4.00 103.00 Placer/alluvial Placer Onshore dredge n.d. n.d.

65 Malaysian 

gravel pump 

mines

Malaysia NA 4.00 103.00 Placer/alluvial Placer Gravel pump n.d. n.d.

66 Sungai 

Lembing

Malaysia Pahang State 3.87 103.08 Hydrothermal Lode, vein n.d. n.d. 0.710

67 Kinta tin field Malaysia Perak State 4.28 101.17 Mostly placer/

alluvial and 

minor hydrothermal lode

Placer and lode Gravel pump, 

dredge

Almost 

completely 

played out

n.d.

68 Kuala Langat 

Mining 

Sdn Bhd

Malaysia Selangor 

State

2.82 101.62 Placer/alluvial Placer Dredge Proposed 0.190 kg/m3

69 Oortsog Ovoo Mongolia Dundgovi 

Province

45.57 108.10 Hydrothermal Skarn Open pit Initial metallurgical study0.007

70 Narsiin 

Khundlen 

(Amerilangui Ujin)

Mongolia Khentii (or 

Hentiy) 

Province

48.74 112.53 Hydrothermal Greisen veins Open pit and 

underground

Feasibility 

study completed, but 

construction 

not yet 

started as 

of 2010 

0.450

71 Achmmach Morocco Meknes 33.32 –5.45 Hydrothermal Veins and disseminated 

in alteration 

envelopes

n.d. In development 

as of 2013

0.85 

(combined 

measured, 

indicated, 

 and inferred)

72 Bou El Jaj Morocco Meknes 33.28 –5.48 Hydrothermal Veins and disseminated 

in alteration 

envelopes

n.d. In development 

as of 2014

NA

73 Achmmach Morocco MeknesTafilalet 

region

33.32 –5.45 Hydrothermal Veins and disseminated 

in alteration 

envelopes

Long-hole 

stoping

Drilling and 

prefeasibility

(expected 

startup in 

2016)

0.850
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

12.0 

(2013)

0.509 

contained 

Sn (1985)

0.900 

contained 

Sn (1985)

1.073 

(1982)

n.d.

n.d.

3.700

7.8 

(2013)

15.3 

(combined 

measured, 

indicated, 

and inferred)

NA

15.3 

(2013)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.003

n.d.

n.d.

n.d.

n.d.

Neoproterozoic sediments

Alluvium, granite, sediments

Alluvium, granite, 

limestone, sandstone, 

metasediments

n.d.

Mostly calcareous rocks 

and some shale, phyllite, 

and schist

Mainly sedimentary terrane, 

middle Paleozoic 

n.d.

Devonian volcanics and 

Jurassic granites

Interbedded sandstone and 

siltstone (315 to 325 Ma)

Interbedded sandstone and 

siltstone (315 to 325 Ma)

Interbedded sandstone and 

siltstone (315 to 325 Ma)

Leucocratic granite

Granite

Granite

Granite

Granite

n.d.

n.d.

Granite

Granite at depth 

presumed

Granite at depth 

presumed

Granite assumed

Middle n.d.

Devonian

Mesozoic n.d.

Mesozoic n.d.

Carboniferous n.d.

Probably Mapping 

Triassic to and 

Jurassic stratigraphy

n.d. n.d.

n.d. n.d.

Jurassic n.d.

Presumed Planned 

Hercynian longhole 

stoping

Presumed n.d.

Hercynian

Late Hercynian n.d.

n.d.

n.d.

n.d.

n.d.

48 km

6.6 km

n.d.

Five prospects 

with ore 

bodies 60 to 

500 m long 

and 0.8 to 

2.5 m thick

about 600 x 

1,600 

n.d.

1,600 x 600 x 

400

Stockwork 

veins

n.d.

n.d.

n.d.

Multiple deposits

n.d.

n.d.

n.d.

NNE-WSWstriking 

vertical 

mineralized 

structures in 

a fold and 

thrust belt

NNE-WSWstriking 

vertical 

mineralized 

structures in 

a fold and 

thrust belt

NNE-WSW 

vertical 

mineralized 

strucures

Velichkin and others 

(2000); Rubinstein and 

Barsky (2002); Syrymbet Mining Co. (2016); 

Peter Kettle, International Tin Research Institute, 

written commun., 

April 8, 2013

Govett and Robinson 

(1980); Sutphin and 

others (1990)

Govett and Robinson 

(1980); Sutphin and 

others (1990)

Sutphin and others (1990)

Rajah (1979)

Bliss and Menzie (1995); 

U.S. Geological Survey 

(2016)

n.d.

Amerilangui Ujin LLC 

(2010); Dandar (2009); 

Jargalsaikhan (1996); 

Peter Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Blight (2012); Pierre 

Chaponniere, Kasbah 

Resources Ltd., written 

commun., July 11, 2013

Blight (2012); Pierre 

Chaponniere, Kasbah 

Resources Ltd., written 

commun., July 11, 2013

Bramwell (2012); Wayne 

Bramwell, Kasbah 

Resources Ltd., written 

commun., July 11, 2013

Table S2 S47
S48 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

74 Bou El Jaz Morocco MeknesTafilalet 

region

33.28 –5.48 Hydrothermal Veins and disseminated 

in alteration 

envelopes

n.d. Drilling Unpublished 

as of 

July 2013

75 Uis tin mine Namibia Southern 

Damaraland

–21.22 14.87 Hydrothermal Lode, altered 

pegmatite

n.d. n.d. 0.260

76 Jos Plateau deposits (Bauchi, Benue, 

Kano, and 

Zaria States)

Nigeria Bauchi, 

Benue, 

Kano, 

and Zaria 

States

9.75 8.87 Placer/alluvial Placer n.d. n.d. n.d.

77 San Rafael Peru Puno Department–14.20 –70.47 Hydrothermal Lode (veins and 

breccias)

Open stope n.d. 10.100

78 San Rafael 

tailings

Peru Puno Department–14.20 –70.47 Tailings NA Tailings from the 

mill associated with the 

San Rafael 

Mine

n.d. 1.300

79 Neves-Corvo 

District. Four 

principal 

individual 

deposits: 

Graca, 

Corvo, 

Neves, 

Zambujal 

and 

Lombador

Portugal Baixo 

Alentejo 

Province

37.60 –7.97 Hydrothermal Lode, volcanichosted massive 

sulfide

Underground Active Graca: 0.17; 

Corvo: 0.30; 

Neves: 0.17; 

Zambujal and 

Lombador: 

0.07 

(0.25 overall, 

with grades 

up to 60)

80 Southern mining

regions of 

the former 

Soviet Union

Russia Chita, Khabarovsk53.00 124.00 Hydrothermal Lode n.d. n.d. n.d.

81 Pyrkakay Russia Chukotskiy 

Avtonomnyy 

Okrug

65.28 175.78 Hydrothermal Porphyry tin Proposed open 

pit

Prefeasibility 0.250

82 Pravourmiyskoe Russia Khabarovsk 

Kray

50.43 134.25 Hydrothermal Tin greisen n.d. n.d. 1.000
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

Unpublished 

as of 

July 2013

16.7 

(1977)

0.111 

contained 

Sn (1969)

0.223 

(1980)

7.7 

(2013)

Graca: 

3.45; 

Corvo: 

20.47; 

Neves: 

28.72; 

Zambujal and 

Lombador: 

32.34 (300 

total)

0.150 

contained 

Sn (1989)

93.0 

(2013)

10.7 

(2013)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Interbedded sandstone and 

siltstone (315 to 325 Ma)

Quartzite schist, pegmatite

Alluvium and eluvium, 

granites, basalt

Metapelites and metapsammites—Ordovician to 

Devonian; psammites 

and carbonates—late 

Paleozoic; red beds and 

intercalated volcanics—

mid-Permian–Triassic;

psammites and carbonates—Cretaceous; 

felsic ignimbrites and red 

beds—Miocene–Pliocene

n.d.

Volcanics, granite

Quartz porphyry granite, 

volcanic quartz porphyry 

tuff

n.d.

Felsic volcanic rocks

Granite assumed

Pegmatite

Granite

Polyphase intrusion of granite, 

leucogranite, and 

granodiorite

n.d.

Granite

Granite

n.d.

Granite and leucogranite (75 to 

85 Ma, K-Ar; 

95 to 83 Ma, 

Rb-Sr)

Late Hercynian n.d.

Unknown n.d.

Jurassic n.d.

22.7±0.7 to Ar-Ar on 

21.9±0.2 hydroMa thermal 

adularia and 

muscovite

n.d. n.d.

Carboniferous n.d.

97 to 91 Ma n.d.

Late Cenozoic n.d.

sedimentary 

rocks

Late Early n.d.

Cretaceous

n.d.

n.d.

n.d.

3,000 x 2 to 50 

x 1,300

n.d.

Graca: 700 x 

500 x 80; 

Corvo: 600 

x 1,100 x 

92; Neves: 

1,200 x 700 

x 55; Zambujal: 550 

x 600 x 53; 

Lombador: 

600 x 1,350 

x 100

n.d.

n.d.

6,750 x 375 x 

2,250

n.d.

n.d.

n.d.

n.d.

n.d.

Massive sulfide; 

fissure; 

stockwork

n.d.

n.d.

n.d.

Blight (2012); Wayne 

Bramwell, Kasbah 

Resources Ltd., written 

commun., July 11, 2013

Richards (1986); Sutphin 

and others (1990)

MacLeod and others 

(1971); Sutphin and 

others (1990)

Mlynarczyk and others 

(2003); Wagner and 

others (2009)

Peter Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Mining Journal (1988); 

Sutphin and others 

(1990); Relvas and 

others, 2006

Smirnov (1977); Sutphin 

and others (1990)

Peter Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Nokleberg and others 

(2005); Peter Kettle, 

International Tin Research Institute, written 

commun., April 8, 2013

Table S2 S49
S50 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

83

84

85

86

87

88

89

90

91

92

93

Northern mining regions 

of the former 

Soviet Union

Lendakhskaya 

Kuils River

Van Rooi’s 

Vley

Rooiberg tin 

mines

Union tin mine

Zaaiplaats tin 

mine

Oropesa

Thai dredging 

operations, 

Southeast 

Asian tin 

belt

Thai gravel 

pump operations, Southeast Asian 

tin belt

Thai hard-rock 

and open pit 

operations, 

Southeast 

Asian tin 

belt

Russia

Russia

South Africa

South Africa

South Africa

South Africa

South Africa

Spain

Thailand

Thailand

Thailand

Magadan, 

Oblast 

and Sakha 

Republic

Western 

Siberia

Cape 

Province

Cape 

Province

Transvaal

Transvaal

Transvaal

Cordoba 

Province

NA

NA

NA

69.00

59.33

–33.93

–28.50

–24.75

–24.48

–24.05

38.00

9.00

8.00

13.87

140.00

92.85

18.73

20.88

27.75

28.50

28.75

–5.48

99.00

99.00

99.12

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Placer/alluvial

Placer/alluvial

Hydrothermal

Lode, vein n.d.

Cassiterite greisen n.d.

Lode, greisen n.d.

veins

Lode, vein n.d.

stockwork and 

replacement 

pipes

Lode Closed in

1993

Lode, vein n.d.

Lode, veins, pipes n.d.

and disseminated ores

Veins, dissem- Open pit

inated ore 

associated with 

shear zones 

and breccias. 

Type unclear. 

Post-orogenic 

system of the 

deep sub volcanic type?

Placer Dredge

Placer Gravel pump

Lode n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Preliminary 

economic 

assessment 

completed 

in 2014

n.d.

n.d.

n.d.

n.d.

0.830

2.440

0.210

n.d.

n.d.

0.165

0.470

0.024

0.015

0.136
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

0.150 

contained 

Sn (1989)

3.1 

(2013)

0.705 

(1973)

2.5 

(1984)

n.d.

n.d.

5.0 (1986)

5.787 

(2013)

2,583.743 

(1982)

1,630.500 

(1984)

54.220 

(1982)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.1

n.d.

n.d.

n.d.

Sandstone and siltstone, 

granite

Granite

Cape granite suite

Paragneiss

Arkosic quartzite, 

shaley quartzite

Indurated tuff, felsite

Leucogranite to mildly 

alkali granite, granophyre, microgranite

Devonian to upper Carboniferous quartzite, slate, 

conglomerate of shale, 

arenite and quartzite 

pebbles and cobbles, 

sandstone, greywacke, 

shale

Alluvial and eluvial sediments, granite

Detrital sediments, granite

Granite

Granite

Granite

Granite

n.d.

n.d.

Tuff?

Granite

Foliated granite(?)

Granite

Granite

Granite

143 to 148 Ma

880 to 864 Ma

553±8 Ma

940±50 Ma

ca. 1,960 Ma

2,050 Ma

ca. 1.96 Ga

Unknown 

(postCarboniferous)

Cretaceous

Tertiary and 

Cretaceous

Cretaceous

K-Ar

SHRIMP 

U–Pb 

zircon 

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

1,300 x 185 x 

200

n.d.

n.d.

n.d.

n.d.

Granite intrusion contacts

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Smirnov (1977); Sutphin 

and others (1990)

Khiltova and Pleskach 

(1997); Vernikovsky 

and others (2007); Peter 

Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Hill and Brunker (1981); 

Sutphin and others 

(1990)

van Backström (1950); 

Sutphin and others 

(1990)

Rozendal and others 

(1986); Sutphin and 

others (1990)

Pringle (1986); Sutphin 

and others (1990)

Zaaiplaats Tin Mining 

Co. (1983); Sutphin 

and others (1990)

Baker (2012)

Sutphin and others (1990)

Nutalaya and others (1979); 

Sutphin and others 

(1990)

Sutphin and others (1990)

Table S2 S51
S52 Critical Mineral Resources of the United States—Tin

Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Deposit 

ID

Deposit 

or mine Country

State, 

Territory, or 

Province

 Latitude Longitude

Deposit class Ore type

Operation type 

and (or) 

comments

Development 

status

Grade Sn 

(wt. % 

unless 

otherwise 

noted)

(decimal degrees 

WGS 84)

94

95

96

97

98

99

100

Geevor

Redmoor

South Crofty

Wheal Jane

Lost River 

district

Sleitat 

Mountain

Kamativi tin 

mines

United 

Kingdom

United 

Kingdom

United 

Kingdom

United 

Kingdom

United States

United States

Zimbabwe

Cornwall

Cornwall

Cornwall

Cornwall

Alaska

Alaska

Matabeleland 

North 

Province

50.15

50.50

50.23

50.23

65.48

60.05

–17.50

–5.68

–4.31

–5.28

–5.12

–167.13

–157.08

27.12

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Hydrothermal

Lode, vein

Stockwork and 

sheeted veins

Lode, vein

Lode,vein

Lode, vein, 

greisen skarn

Sheeted greisen 

veins

Lode, pegmatite

n.d.

Shrinkage 

methods—

overhand 

shrinkage

Shrinkage 

methods—

overhand 

shrinkage

n.d.

n.d.

n.d.

n.d.

n.d.

Scoping study

n.d.

n.d.

n.d.

n.d.

n.d.

0.570

0.210

0.440

0.890

1.300

0.300

0.114
Table S2. Location, grade, tonnage, and other data for selected tin deposits of the world with greater than 1,000 metric tons of 

contained tin.—Continued

[WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern Hemisphere; negative values for longitude 

indicate that the deposit is in the Western Hemisphere. Elements: B, boron; F, fluorine; Fe, iron; Sn, tin. Dating method: Ar-Ar, argon-argon; K-Ar, potassiumargon; Rb-Sr, rubidium-strontium; U-Pb, uranium-lead. Abbreviations: ca., circa; Ga, giga-annum, or billion years before present; ID, identification; kg/m3

, 

kilogram per cubic meter; m, meter; Ma, mega-annum, or million years before present; NA, not applicable; n.d., no data; NNE-WSW, north-northeast—

west-southwest; SHRIMP, sensitive high-resolution ion microprobe; wt. %, weight percent]

Resource 

tonnage 

(million 

metric tons) 

(Year 

determined)

Cutoff 

grade 

(wt. %)

Host rock Mineralizing 

rock-type

Mineralization 

age

Dating 

method

Deposit 

dimensions 

(in meters 

unless 

otherwise 

noted)

Deposit shape/ 

mineralization

style

Primary references

0.696 

(1986)

9.1 

(2013)

1.331 

(2013)

2.380 

(1987)

0.0024 

contained 

Sn (1963)

25.9 

(2013)

100.00 

(1962)

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Slate, granite

Killas slate—Devonian

Granite, slate

Slate, quartz porphyry 

dikes

Rhyolite porphyry dikes, 

limestone, granite

Mid-Cretaceous siltstones 

of the Kuskokwim Group

Pegmatite, schist

Granite

Granite

Granite

Quartz porphyry 

dikes

Granite

Biotite-muscovite 

granite

Pegmatite

270±5 Ma, 

215±5 Ma

Middle Permian

270±5 Ma, 

215±5 Ma

270±10 Ma, 

215±5 Ma

80.2 Ma

56.8±2.8 Ma

Precambrian

K-Ar, 

Rb-Sr, 

U-Pb

n.d.

K-Ar, 

Rb-Sr, 

U-Pb

n.d.

K-Ar on 

biotitie

K-Ar on 

muscovite

n.d.

n.d.

500 x 80 x 400 

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Lenticular

n.d.

n.d.

n.d.

n.d.

Dines (1956); Sutphin and 

others (1990)

Dines (1956); Peter Kettle, 

International Tin Research Institute, written 

commun., April 8, 2013

Dines (1956); Sutphin and 

others (1990); Peter 

Kettle, International 

Tin Research Institute, 

written commun., 

April 8, 2013

Dines (1956); Sutphin and 

others (1990)

Dobson (1982); Sainsbury 

(1968); Sutphin and 

others (1990)

Burleigh (1991); Hudson 

and Reed (1997)

Bellasis and van der Heyde 

(1962); Rijks and van 

der Veen (1972); Sutphin 

and others (1990)

Table S2 S53

Titanium

By Laurel G. Woodruff, George M. Bedinger, and Nadine M. Piatak

Chapter T of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802–T

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter T

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................ T1

Introduction.................................................................................................................................................. T1

Titanium Commodities ....................................................................................................................... T1

Settings for Titanium Oxide Mineral Occurrences ....................................................................... T2

Uses and Applications ...................................................................................................................... T2

Strategic and Critical Resource Issues ......................................................................................... T4

Geology......................................................................................................................................................... T4

Geochemistry...................................................................................................................................... T4

Mineralogy.......................................................................................................................................... T4

Deposit Types ..................................................................................................................................... T6

Titanium Deposits Related to Igneous Rock ......................................................................... T6

Titanium Deposits Related to Metamorphic Rocks ............................................................. T8

Titanium Deposits Related to Coastal Shorelines and Dunes, and 

Older Equivalent Rocks .............................................................................................. T9

Resources and Production ...................................................................................................................... T11

Identified Resources ....................................................................................................................... T11

Undiscovered Resources ............................................................................................................... T12

Exploration for New Deposits ................................................................................................................. T15

Environmental Considerations ................................................................................................................ T15

Sources and Fate in the Environment .......................................................................................... T15

Mine Waste Characteristics .......................................................................................................... T16

Human Health Concerns ................................................................................................................. T16

Ecological Health Concerns ........................................................................................................... T16

Mine Closure..................................................................................................................................... T17

Problems and Future Research .............................................................................................................. T18

References Cited....................................................................................................................................... T18

Suggested citation:

Woodruff, L.G., Bedinger, G.M., and Piatak, N.M., 2017, Titanium, chap. T of Schulz, K.J., DeYoung, J.H., Jr., Seal, R.R., II, 

and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental geology and 

prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. T1– T23, http://dx.doi.org/10.3133/

pp1802T.
Titanium

By Laurel G. Woodruff, George M. Bedinger, and Nadine M. Piatak

Abstract

Titanium is a mineral commodity that is essential to the 

smooth functioning of modern industrial economies. Most of 

the titanium produced is refined into titanium dioxide, which 

has a high refractive index and is thus able to impart a durable 

white color to paint, paper, plastic, rubber, and wallboard. 

Because of their high strength-to-weight ratio and corrosion 

resistance, titanium metal and titanium metal alloys are used 

in the aerospace industry as well as for welding rod coatings, 

biological implants, and consumer goods.

Ilmenite and rutile are currently the principal titaniumbearing ore minerals, although other minerals, including 

anatase, perovskite, and titanomagnetite, could have economic 

importance in the future. Ilmenite is currently being mined 

from two large magmatic deposits hosted in rocks of 

Proterozoic-age anorthosite plutonic suites. Most rutile and 

nearly one-half of the ilmenite produced are from heavymineral alluvial, fluvial, and eolian deposits. Titanium-bearing 

minerals occur in diverse geologic settings, but many of 

the known deposits are currently subeconomic for titanium 

because of complications related to the mineralogy or because 

of the presence of trace contaminants that can compromise the 

pigment production process.

Global production of titanium minerals is currently 

dominated by Australia, Canada, Norway, and South Africa; 

additional amounts are produced in Brazil, India, Madagascar, 

Mozambique, Sierra Leone, and Sri Lanka. The United States 

accounts for about 4 percent of the total world production 

of titanium minerals and is heavily dependent on imports of 

titanium mineral concentrates to meet its domestic needs.

Titanium occurs only in silicate or oxide minerals and 

never in sulfide minerals. Environmental considerations for 

titanium mining are related to waste rock disposal and the 

impact of trace constituents on water quality. Because titanium 

is generally inert in the environment, human health risks 

from titanium and titanium mining are minimal; however, the 

processes required to extract titanium from titanium feedstock 

can produce industrial waste.

Introduction

Titanium is an abundant element that has become an 

important industrial commodity only within the past 150 years. 

Ilmenite (FeTiO3 ) and rutile (which is the most common form 

of titanium dioxide [TiO2 ]) are the two principal minerals that 

are sources of titanium. Pure titanium metal was first isolated 

in 1910, but the metal was not commercially available until the 

development of the Kroll process by William Kroll in 1938. 

In this process, titanium is produced by reduction of titanium 

tetrachloride with magnesium. Titanium is now considered 

essential for modern industrial technology production as well 

as to meet other needs of modern affluent societies. Many of 

the uses for titanium in modern industry, including in pigments 

for paints and paper and in manufacturing by the aerospace 

industry, are described by Gambogi (2012) and Murphy and 

Frick (2006). Although the United States is dependent on 

imports of titanium mineral concentrates and titanium sponge, 

the United States continues to be a net exporter of TiO2

pigment and titanium wrought-metal products.

Titanium Commodities

The major titanium commodities and their industrial 

uses are described in Murphy and Frick (2006). Titanium 

commodity sources include both titanium-bearing oxide 

minerals extracted from natural sources and synthetic titaniumbearing materials processed from these minerals. Some 

93 percent of the world’s titanium resources are accounted for 

by the mineral ilmenite; the remaining percentage is mainly 

from rutile and leucoxene. Leucoxene is not a true mineral 

but is, rather, a finely crystalline aggregate of rutile, anatase, 

or brookite with greater than 70 percent TiO2 that is created 

by the weathering of ilmenite. Rutile, anatase, and brookite 

are TiO2 polymorphs; that is, minerals that share the same 

chemical composition of TiO2

 but have different crystal forms.

Ilmenite, rutile, and leucoxene are processed mainly into 

titanium-rich slag or synthetic rutile and then further processed 

for use in pigments and as titanium metal. Titanium slag, which 
T2 Critical Mineral Resources of the United States—Titanium

is derived by smelting ilmenite, contains about 75 to 86 percent 

TiO2; with further processing by high-pressure acid leaching, 

titanium slag can be upgraded to about 94.5 percent TiO2

(Guéguin and Cardarelli, 2007). Synthetic rutile is produced 

by removing iron from secondary ilmenite feed, which 

increases its TiO2 content to about 90 to 95 percent TiO2 from 

about 50 to 55 percent TiO2 (Mackey, 1994). Synthetic rutile 

can be used as a substitute for natural rutile.

Two main industrial processes—a sulfate process and 

a chloride process—are currently used to convert the raw 

materials of ilmenite and rutile into pure white TiO2 for use in 

pigment or into titanium metal (Mackey, 1994). The sulfate 

process is the older of the two processes; it was perfected in 

1916 by two Norwegian chemists (Korneliussen and others, 

2000a). In the sulfate process, finely ground ilmenite or 

high-TiO2 slag is digested with concentrated sulfuric acid. In 

the chloride process, high titanium-content feedstock, such 

as natural and synthetic rutile, high-TiO2 slag, or high-TiO2

ilmenite sands, is converted into titanium chlorides at high 

temperature. As of 1992, the sulfate and chloride processes 

produced about equal amounts of product; however, because 

it produces a superior TiO2 pigment product and fewer waste 

products, the chloride process has largely supplanted the 

sulfate process. This shift from the sulfate process to the 

chloride process for TiO2 pigment production has increased 

demand for a high-grade titanium feedstock, such as rutile, 

upgraded titanium-rich slag, and synthetic rutile.

Settings for Titanium Oxide Mineral Occurrences

The diverse types of economic and potentially economic 

titanium deposits originally categorized by rock type and 

environment by Force (1991) have been updated in table T1 

and are described in the geology section of this chapter. Major 

deposit types listed in table T1 include (a) magmatic deposits 

in igneous and related rocks with ilmenite as the primary ore 

mineral; (b) metamorphic deposits with rutile as the primary 

ore mineral; (c) rutile in hydrothermal porphyry deposits, 

and (d) heavy-mineral concentrations in unconsolidated and 

lithified sand deposits that may include rutile, ilmenite, and 

alteration products, such as leucoxene. Most deposit types 

(for example, those in the I–8 and S–4 classes) can be further 

altered by weathering, at times upgrading the titanium content 

of the ilmenite, if present.

Ilmenite accounts for about 92 percent of the world’s 

consumption of titanium minerals. The world’s two active 

magmatic titanium mines—the Tellnes Mine in Norway 

and the Lac Tio Mine in the Province of Quebec, Canada— 

together produce an estimated 32 percent of the world’s 

titanium (Gross, 1996; Schiellerup and others, 2003). The 

remaining 68 percent comes from high-grade ilmenite and 

rutile heavy-mineral deposits in unconsolidated heavy-mineral 

shoreline and dune deposits, fluvial placers (surface mineral 

deposits formed by mechanical concentration of minerals 

from weathered rock), and older lithified equivalents.

Magmatic iron-titanium-oxide-rich deposits are hosted 

by a complex suite of mafic rocks that includes anorthosite, 

gabbro, and norite. The principal ore minerals are ilmenite and 

hemo-ilmenite, which is ilmenite that has extensive hematite 

(Fe2O3) exsolution lamellae, so that hematite occurs as small 

parallel plates within larger ilmenite crystals. Titaniferous 

magnetite, which is a complex oxide made up primarily 

of magnetite (Fe3O4) with iron-titanium-oxide exsolution 

lamellae, can also occur in these rocks in high concentrations, 

but it is of less economic importance because of its low 

titanium content.

Heavy-mineral deposits contain concentrations of 

iron-titanium-oxide minerals that are denser than quartz 

and so lag behind during erosion and weathering as lighter 

minerals are more rapidly transported away or broken down. 

A typical mineral assemblage in heavy-mineral deposits can 

include rutile and ilmenite (and minerals produced from 

the weathering of ilmenite, such as anatase, leucoxene, and 

pseudorutile), as well as other high-density, erosion-resistant 

minerals, including chromite, garnet, kyanite, monazite, 

staurolite, tourmaline, and zircon. Heavy-mineral deposits 

develop when relatively resistant iron-titanium-oxide minerals 

and other heavy minerals, such as monazite and zircon, are 

eroded from the parent rocks, transported and sorted, and 

finally deposited in fluvial, alluvial, or eolian settings. Many 

productive heavy-mineral provinces, such as the East Coast of 

the United States, the east and west coasts of South Africa, and 

the east and west coasts of Australia, are located on passive 

continental margins that are usually backed by elevated and 

often highly weathered high-grade metamorphic or mafic 

igneous hinterlands (Force, 1991).

Uses and Applications

About 93 percent of the world’s titanium is refined into 

TiO2 for pigment; consequently, trends in the TiO2 pigment 

industry exert a large influence on the development of 

titanium resources (Towner and others, 1988; Murphy and 

Frick, 2006). Titanium dioxide has properties of whiteness, 

opacity, and chemical inertness that make it especially 

suitable for use as pigment to impart a durable white color to 

paints, paper, plastic, sunscreen, toothpaste, and wallboard 

(Miller, 1957; Stamper, 1970). The leading uses of TiO2

pigment in the United States in 2012 were paint (59 percent), 

plastics (28 percent), paper (9 percent), and other (4 percent) 

(Bedinger, 2013). Titanium is different than most other 

metallic elements in that it is mined primarily to satisfy 

demands for a chemical product (TiO2 for pigment) rather than 

for the metal itself (Murphy and Frick, 2006). The high cost 

of extracting titanium metal from ore curtails broader use of 

titanium metal and alloys.

About 2.5 percent of the world’s titanium is used for 

the production of corrosion-resistant titanium metal (Murphy 

and Frick, 2006). Titanium metal alloys are desirable for 

their high strength-to-weight ratio and corrosion resistance, 
Introduction T3

Table T1. Classification of selected titanium mineral deposits based on their geologic setting and the processes through which they 

were formed.

[Modified from Force (1991). The relative economic importance of each deposit type, which is subjective, is indicated as follows: 1, of great current importance; 2, of probable great importance in the near future; 3, of possible importance; 4, of moderate importance; and 5, of minor importance. Locality abbreviations: Ark., Arkansas; Colo., Colorado; Fla., Florida; Md., Maryland; Minn., Minnesota; N.Y., New York; Va., Virginia] 

Deposit 

class Deposit type1 Typical mineralogy

Level of 

economic

importance

Example deposits

Magmatic (igneous and igneous related)

I–1 Massif anorthosite Hemo-ilmenite 1 Lac Tio, Quebec; Tellnes, Norway

I–2 Anorthosite-gabbro Ilmenite, titanomagnetite 2 Sanford Lake district, N.Y.

I–3 Layered mafic intrusion Ilmenite, titanomagnetite 2 Magpie, Quebec; Panzihuia, China

I–4 Troctolite/ultramafic Ilmenite, titanomagnetite 3 Longnose, Minn.

I–5 Albititic/metasomatized Rutile 3 Kragerø, Norway; Roseland, Va. (in part)

I–6 Alkalic/metasomatized Perovskite, brookite, rutile 3 Iron Hill, Colo.; Magnet Cove, Ark.

I–7 Weathered alkalic rocks Anatase 2 Catalão, Salitre, and Tapira, Brazil

I–8 Weathered anorthositic rocks Ilmenite, rutile 4 Roseland, Va. (in part)

Metamorphic

M–1 Eclogite-hosted Rutile 2 Engebøfjellet, Norway; Piampaludo, Italy

M–2 Amphibolite-hosted Rutile 3 Daixian, China

M–3 Greenschist-hosted Rutile 5 Dinning, Md.

M–4 Contact metasomatized anorthosite Ilmenite, rutile 4 Roseland, Va. (in part)

M–5 Metasomatized aluminum-rich schist Rutile 5 Evergreen, Colo.

Hydrothermal (igneous and [or] metamorphic)

I/M Hydrothermal porphyry ore deposits Rutile 3 Bingham, Utah; El Teniente, Chile

Sedimentary-related

S–1 Fluvial Rutile, ilmenite 1 Mogbwemo and Sherbo River,

Sierra Leone

S–2 Beach (strandline) and (or) 

coastal dune

Ilmenite, altered ilmenite, 

rutile, leucoxene

1 North Stradbroke Island, eastern Australia;

Richards Bay, South Africa; 

Trail Ridge, Fla. (in part)

S–3 Lithified paleo-placers (fluvial

and [or] beach)

Ilmenite, altered ilmenite, 

rutile, leucoxene

2 Bothaville, South Africa

S – 4 Weathered sedimentary deposits Altered ilmenite, leucoxene 1 Trail Ridge, Fla. (in part)

1

Individual deposit types are further classified and described more fully in the text.

and they are used widely in the manufacturing of aircraft. In 

general, production of titanium mill products precedes aircraft 

deliveries by about 1 year. In 2010, mill product shipments 

increased by 62 percent because of increased aircraft build 

rates and higher inventory levels throughout the titanium 

supply chain (Gambogi, 2012). The aerospace industry 

(67 percent) was the leading end user of mill products.

Titanium has other important end uses. For example, 

titanium is used in welding-rod coatings, where it protects 

the surfaces of a weld from unwanted reaction with the 

atmosphere during welding, as well as in the manufacture 

of mineral fibers; it is used in ceramic components for 

electronic equipment, and in carbides and other titanium 

chemicals. Because of titanium’s resistance to corrosion 

in saltwater, it is used in shipbuilding, deep-sea petroleum 

exploration and production, and geothermal power-generating 

facilities. A significant quantity of titanium in the form of 

ferrotitanium, scrap, and sponge (a porous form of titanium 

created during the first stage of processing) is consumed in 

the steel and nonferrous metal alloy industries. In the steel 

industry, titanium is used for deoxidation, grain size control, 

and controlling and stabilizing carbon and nitrogen content. 

Titanium-intensive steels include interstitial-free, stainless, 

and high-strength low-alloy steels. Titanium has the property 

of osseointegration, which means that it can be structurally 

accepted by bone, and because titanium is nonreactive, it is 

increasingly used to make medical devices, such as hip and 

knee joints, bone screws and plates, and dental implants.
T4 Critical Mineral Resources of the United States—Titanium

Strategic and Critical Resource Issues

Titanium is considered a critical and strategic mineral 

because of the unique properties of both titanium metal 

(and its alloys) and TiO2 pigment. There are no completely 

satisfactory substitutes for titanium, especially titanium metal 

(Towner and others, 1988). Titanium metal’s combination 

of corrosion resistance, excellent weight-to-strength ratio, 

and very high melting point is not found with other metals. 

Substitutes for TiO2 pigment, such as zinc oxide, lithopone 

(a mixture of barium sulfate and zinc sulfide), and calcium 

carbonate, generally result in an inferior product and 

are less environmentally safe. Although lead has similar 

whitening properties, titanium dioxide pigment replaced 

lead as a whitener because of public health issues related to 

lead toxicity.

Consumption of titanium mineral concentrates in the 

United States exceeds domestic production (Gambogi, 

2012). From 2008 to 2012, U.S. consumption of titanium 

concentrates averaged about 1,400,000 metric tons. The net 

import reliance for the United States in that same time period, 

as a percentage of apparent consumption, was between 

65 and 78 percent. The primary sources of U.S. imports of 

titanium mineral concentrates, which include ilmenite, rutile, 

synthetic rutile, and titanium slag, were Australia, Canada, 

Mozambique, and South Africa.

The titanium feedstock industry is dominated by a 

relatively few companies at both ends of the supply chain; 

consequently, the actions of a few industry leaders can have 

implications for the entire industry (Murphy and Frick, 2006). 

Titanium resources are located in relatively few countries, 

which means that many industrial nations are dependent on 

titanium imports (Towner and others, 1988). In addition, the 

influence of China is an important strategic issue for the titanium industry, as increased economic growth in that nation 

will likely increase its consumption of titanium minerals, 

pigments, and metal in the long term (Gambogi, 2012). 

Increasing demand in China for titanium will be paced by 

increasing demand in the United States as the U.S. economy 

recovers from the recession that began in 2008 and as 

aerospace, defense, and industrial uses grow. For example, 

new lighter and fuel-efficient aircraft and ground vehicles 

will require significantly more titanium in their structures to 

reduce weight and increase durability. Global consumption 

of titanium by the defense industry alone was expected to 

increase by slightly more than 30 percent between 2010 

and 2015 (Gambogi, 2012). Improving global economic 

conditions resulted in a 7 percent increase in domestic TiO2

pigment production from 2009 to 2010 (Gambogi, 2012). 

Thus, the dynamics of the global titanium industry are a 

complex interdependence of supply and demand, industry 

activities, and environmental regulations, all of which 

could be highly influenced by the discovery of new uses for 

titanium and by technological advances in the extraction 

and production of titanium metal and TiO2 pigment.

Geology

Geochemistry

Titanium is the ninth most abundant element in Earth’s 

crust, with an average TiO2 abundance of 0.7 weight percent 

(Rudnick and Fountain, 1995). Titanium can be found in 

nearly all rocks and sediments, often at contents greater than 

1 weight percent, and thus, unlike many other elements of 

economic interest, it is considered to be a major element 

rather than a trace element. Titanium is a transition element 

with atomic number 22 and an atomic mass of 47.867 atomic 

mass units. Titanium’s chemistry has similarities to the 

chemistries of silicon and zirconium. The main oxidation state 

is Ti4+, although Ti2+ and Ti3+ are known to exist. Titanium is 

a lithophile element with a strong affinity for oxygen, and it 

typically forms oxide minerals. It is not found as a pure metal 

in nature, and processing is required to produce a low-density, 

high-strength metal. Titanium metal has excellent corrosion 

resistance, and it is as strong as steel but 45 percent lighter.

Mineralogy

The most economically important titanium oxides are 

ilmenite (FeTiO3 ) and the titanium dioxide (TiO2 ) polymorphs 

rutile, anatase, and brookite. Only some titanium-bearing 

oxide minerals currently have economic value. Less common, 

subeconomic minerals include pseudobrookite (Fe2TiO5 ), 

perovskite (CaTiO3), geikielite ((Mg,Fe)TiO3 ), pyrophanite 

(MnTiO3 ), and titanite or sphene (CaTiSiO5 ). The titanium 

content in titanomagnetite ranges between about 2 percent 

and about 20 percent, and titaniferous magnetite as a group 

of iron-titanium-oxide ore minerals contains between about 

1 percent and about 15 percent titanium (fig. T1; Rose, 1969).

The mineralogy of iron-titanium oxides in magmatic 

deposits is dependent on several factors, including the initial 

magma composition, the magma oxygen fugacity, and the 

cooling history. At magmatic temperatures (>700 degrees 

Celsius [°C]), iron-titanium-rich liquids are in a single, 

homogenous phase after which, during cooling, the liquids 

will separate into two or more oxide minerals. In a hightemperature titanium-rich magma, a small amount of hematite 

(Fe2O3 ) will separate out of the cooling oxide phase to form 

thin lamellae within ilmenite crystals. Deposits dominated 

by the composite mineral hemo-ilmenite (fig. T2A) are the 

largest and most economically important magmatic deposits. 

In more iron-rich magmas under varying magmatic conditions, a titanium-rich phase (usually ilmenite or ulvöspinel 

[Fe2TiO4 ]) will separate from magnetite (Fe3O4 ) to form 

(a) titanomagnetite, which is a complex mixture of irontitanium oxides with thin exsolution lamellae of ilmenite or 

ulvöspinel in magnetite, or (b) granular, interlocking discrete 

grains of ilmenite, ulvöspinel, and magnetite (fig. T2B). 

Titaniferous magnetite is a general mining term applied 

to low-temperature aggregations of hematite, ilmenite, 
Geology T5

100

Titanium content, in weight percent

80

60

40

20

0

Rutile, anatase, brookite

Pseudorutile

Ilmenite

Titanium Fig T1

Titanite (sphene)

Ulvöspinel

Pseudobrookite

Perovskite

Leucoxene

Titanomagnetite

and

titaniferous

magnetite

Figure T1. Chart showing common titanium-bearing 

oxide minerals (black text) and common titanium-bearing 

silicate minerals (red text) with their approximate titanium 

content (in weight percent). The blue shaded boxes 

represent the range of titanium content for the mineral 

species in the boxes.

magnetite, titanomagnetite, and other intermediate members 

of the iron-titanium-oxide solid solution series (Rose, 1969). 

With current technologies, deposits dominated by titaniferous 

magnetite are generally too high in titanium to mine for iron 

ore yet too low in titanium to mine for titanium; however, if 

metallurgical problems with extraction and separation could 

be resolved, then deposits containing very large resources of 

titaniferous magnetite could be economically viable for both 

elements (Rose, 1969). Substitution of aluminum, calcium, 

chromium, copper, magnesium, manganese, silicon, vanadium, 

and zinc in the ilmenite crystal structure can lower the quality 

of ilmenite concentrate, as these elements decrease the TiO2

grade and may affect pigment color (Chernet, 1999). Ilmenite 

also is a common detrital mineral, being somewhat resistant to 

erosion processes. Weathering of ilmenite typically involves 

progressive leaching of iron from its mineral structure, thereby 

enriching the altered ilmenite in titanium and potentially 

forming pseudorutile (Fe2Ti3O9), leucoxene, and rutile 

(fig. T1). Because of this upgrading in titanium content, 

altered ilmenite is a preferred ore target.

A

B

Figure T2. Reflected-light photographs showing A, lamellae of 

hematite (light-pink) in ilmenite (dark pink) in hemo-ilmenite from 

the magmatic Lac Tio hemo-ilmenite deposit, and B, titaniferous 

magnetite (Ti-Mte) and ilmenite (Ilm) from the magmatic La Blache

iron-titanium-oxide deposit. Scale bar is 0.5 millimeters in length. 

Photographs courtesy of Serge Perreault, SOQUEM, Inc.

Rutile is the most common of the TiO2 polymorphs. It has 

a high index of refraction and thus a high degree of opacity, 

which accounts for rutile’s desirability for use in pigments. 

Rutile occurs as an accessory mineral in many igneous rocks 

and in all grades of metamorphic rocks. With increasing 

metamorphic grade, the breakdown of ilmenite and silicates 

will release titanium, which then crystallizes into rutile. Rutile 

is highly stable across most geologic environments, which 

explains its presence in a wide range of rock types and as a 

common detrital mineral. Rutile can be a major host mineral 

for niobium, tantalum, tungsten, uranium, zirconium, and 

ferric iron. The presence of trace elements is in part dependent 

on the availability of trace constituents during rutile formation. 

Anatase and brookite are low-temperature alteration products 

of ilmenite or titanite.
T6 Critical Mineral Resources of the United States—Titanium

Deposit Types

Table T1 shows the classification of titanium mineral 

deposit types. These classifications can have considerable 

overlap. For example, magmatic concentrations of iron- and 

titanium-bearing silicates and oxides can be metamorphosed 

to form rutile, weathered from parent rocks to subsequently 

reside in place within weathered zones, or transported and 

sorted by wind or water to form heavy-mineral deposits. 

A common thread among all the deposit types described in 

table T1 is their ultimate origin from rocks somewhat enriched 

in titanium by geologic processes.

The worldwide distribution of titanium deposits is shown 

in figure T3. The economic value and mining potential of a 

titanium deposit is highly dependent on its mineralogy rather 

than its overall titanium content. Although there are very large 

global resources of ilmenite and rutile combined, there is a 

worldwide lack of high-grade, high-quality ilmenite deposits 

as well as of high-quality rutile deposits (Korneliussen and 

others, 2000a). Mineral assemblage, grain size, morphology, 

texture, and type and quantity of included trace elements all 

contribute to the economic potential of a deposit.

Qualitative differences in the approximate TiO2

 content 

of different deposit types are shown in figure T4. In sedimentary deposits, titanium-bearing minerals have been released 

from their parent rock; thus, low-grade deposits of this type 

may be mined if the titanium minerals are of sufficiently 

high quality, either through alteration of ilmenite or by the 

presence of rutile. Brooks (1976, p. 149), in postulating 

that the ultimate “cut-off grade” for each metal is its crustal 

abundance, acknowledges former U.S. Geological Survey 

Director Vincent McKelvey for pointing out that the titanium 

content of ilmenite beach placers mined in Florida is below 

the crustal abundance of titanium.

The presence or absence of discrete crystals of ilmenite 

is a key factor that determines the economic value of most 

magmatic oxide deposits. In magmatic hemo-ilmenite 

deposits, such as the deposit at Tellnes (Norway), ore has to be 

mined and crushed and the ilmenite separated from the waste 

rock; therefore, ore grades in these types of deposits must be 

high (Korneliussen and others, 2000a). Deposits with unconventional titanium ore minerals, such as Iron Hill (Colorado; 

also called Powderhorn in the literature), which host perovskite, 

are currently uneconomic; likewise, the anatase deposits in 

Brazil (the Catalão, Salitre, and Tapira deposits) are not being 

exploited for titanium at this time.

Titanium Deposits Related to Igneous Rock

Mafic intrusions of various age and rock types can 

contain significant iron-titanium-oxide deposits. Deposits 

dominated by ilmenite and hemo-ilmenite are hosted by 

discordant and layered bodies in Proterozoic-age massif 

anorthosite plutonic complexes. These unique rock suites 

are restricted in time, from about 0.5 to 2.5 Ga (giga-annum, 

or billion years ago); the majority that contain significant 

titanium resources were emplaced from about 1.0 to 1.8 Ga 

(Ashwal, 2010). The plutonic suites are dominated by different 

and often unusual mafic rock types. These rocks are called, 

for example, anorthosite, charnockite, jotunite, gabbro, norite, 

or troctolite. The different names distinguish among rocks 

with various proportions and compositions of plagioclase 

and pyroxene. Anorthosite plutonic suites of the Grenville 

Province in eastern Canada and parts of the Northeastern 

United States and the Rogaland anorthosite province in 

Norway contain numerous iron-titanium-oxide deposits (Rose, 

1969; Krause and others, 1985; Corriveau and others, 2007). 

Hemo-ilmenite (with little or no magnetite) is currently being 

mined at Lac Tio in Quebec Province, Canada, and is hosted 

by the Allard Lake anorthosite in the Havre-Saint-Pierre 

anorthosite plutonic suite (Gross, 1996). Hemo-ilmenite is 

also being mined at Tellnes, which is a hemo-ilmenite-bearing 

norite lens in the Åna-Sira anorthosite pluton in Norway 

(Wilmart and others, 1989). A deposit model for titanium 

deposits hosted by Proterozoic-age massif anorthosite plutonic 

suites was completed by Woodruff and others (2013).

Deposits dominated by titaniferous magnetite with 

minor amounts of ilmenite are hosted in layered and massive 

intrusions of gabbro, leucogabbro (light-colored gabbro), 

and norite, which can include, but are not restricted to, 

Proterozoic-age massif anorthosites (Gross, 1996). These 

types of deposits have somewhat different characteristics 

than the magmatic deposits typified by Lac Tio or Tellnes, 

including differences in host rock form, dominant ore 

mineralogy, and proposed magmatic history. Examples include 

numerous deposits in layered and discordant anorthositenorite-gabbro intrusions in the Grenville Province, such as 

Magpie Mountain and Lac la Blache (Canada) and Sanford 

Lake district in the Adirondacks (New York), and large 

layered mafic intrusions of various ages, such as the Bushveld 

Complex (South Africa), which is currently being exploited 

for vanadium from titaniferous magnetite. Titanomagnetite 

and ilmenite are mined for iron, titanium, and vanadium from 

several deposits in the Permian-age Panzhihua layered mafic 

intrusion in China (Zhou and others, 2005).

Other magmatic occurrences are somewhat enigmatic. 

For example, in the Duluth Complex of Minnesota, small 

bodies of ilmenite and titanomagnetite hosted by ultramafic 

rocks that cross-cut gabbroic rocks are a significant (but 

presently subeconomic) titanium resource. The origin of these 

deposits remains under debate (Hauck and others, 1997). 

In Finland, the relatively small Koivusaarenneva gabbro 

intrusion contains layers of ilmenite (with minor amounts 

of magnetite) that are interpreted as iron-titanium-oxides 

deposited in a flow-through magmatic system that has created 

a large resource in a small-volume intrusion (Kärkkäinen and 

Bornhorst, 2003).

Formation of iron-titanium-oxide deposits in mafic 

intrusions requires favorable combinations of composition, 

oxygen fugacity, and crystallization history, which determine 

if a deposit is dominated by ilmenite or titaniferous magnetite 

(Woodruff and others, 2013). Massive hemo-ilmenite 
Geology T7

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

Titanium figure 3. [D, District; S, State]

Sedimentary/heavy mineralIgneousMetamorphic

Large titanium deposits

EXPLANATION

Sedimentary/heavy mineralIgneousMetamorphic

Small titanium deposits

Agnes Water

Fraser Island

Stockton

Tomago

Jurien, Cooljarloo,Gingin, and Waroona Ginkgo

Cox's Bazaar-Moiskal Island

Campo Alegre de Lourdes

Lac du Pin

Rouge

Suncor

Syncrude

Heishan

Behai D

Xun Jiang

Wazhaung D

Daixian

Otanmaki

Pluma Hidalgo

Westport

Storgangen Kodal

Bothaville

Hillendale-Fairbreeze

Transeki (Wave Crest) Namakwa Sands

Gravellote

Lukoyanovskoye

Taraskova

Tsentral'noye

Tuganskoye

Yugo-Vostochnaya Gremyakha

Bol'shoy Seyim

Shubino Village

Cerro Blanco

Svoranzkoye Lakehurst and Manchester Bingham Canyon

Longnose

Magnet

Cove

Piney River (Roseland), Old Hickory, Brink, Halifax, and Aurora

Yadkin Valley

Amelia, Brunswick, Folkston, and Cumberland Island

San Gabriel Mountains

Laramie Range 

Tucker 

Ødegård

Engebøfjellet

Bridge Hill Ridge

Moreton Island

}

}

}

} }

Cyclone, Typhoon, and Atacama

Oak Grove, Kerr-McGee,

and Natchez Trace

}

Evergreen

North

Stradbroke

Island

Barrambie

Capel (2)

Eneabba (2)

JacinthAmbrosia

Kulwin

Woornack, Rownack

& Pirro

Catalao

Salitre

Vitoria D

Lac la Blache Magpie Mountain

Lac Tio

St. Urbain D

Bad Vermillion

Panzhihua

Andhra Pradesh S (8)

Kerala S (3) 

Odisha S (2)

Tamil Nadu S (3)

Ranobe

QMM

Moma

Kwale

Grand Côte

Gbangbama D

Barrytown

Tellnes

Pulmoddai

Richards Bay

Afrikanda

Chineyskoye

Yaregskoye

Irshansk D

Volnogorsk D

Green Cove Springs

Iron Hill (Powderhorn)

Sanford Lake D

Trail Ridge/Highland (Maxville)

Patrocinio

Tapira

Mataraca

Piampaludo

described more fully in World map showing global distribution of titanium deposits, by size and type of depost. Deposit types are from table T1 and are 

ize is denoted by the relative size of each 

into-Madagascar Government joint venture to extract ilmenite from sands 

Figure T3.the text. D, district; S, State; numbers in parentheses indicate the number of deposits included in a single symbol. Deposit s

owner and others (1988) and Force (1991).

circle and descriptive text. QMM stands for QIT Madagascar Minerals, which is a Rio Taround Fort Dauphin. Deposit information was compiled primarily from T
T8 Critical Mineral Resources of the United States—Titanium

Figure T4. Bar chart comparing approximate titanium 

2

) content in titanium deposits of different 

deposit types (modified from Korneliussen and others, 

2000a). The sources of ore given here as examples are 

average crustal abundance (Rudnick and Fountain, 1995); 

average rutile and ilmenite sands (sedimentary) (Force, 

1991); rutile in eclogite from the Engebøfjellet deposit in 

Norway (metamorphic) (Korneliussen and others, 2000a); 

perovskite in pyroxenite at the Iron Hill (Powderhorn) 

deposit in Colorado (igneous) (Van Gosen, 2009); hemoilmenite in anorthosite at the Tellnes deposit in Norway 

(igneous) (Korneliussen and others, 2000a); and anatase 

in weathered rocks from the Catalão, the Salitre, and the 

Tapira deposits in Brazil (Force, 1991).

dioxide (TiO

Sedimentary Metamorphic Igneous Weathered

20

22

18

16

14

12

10

8

6

4

2

0

Average crustal abundance 

 (for comparison)

Hemo-ilmenite in anorthosite (Tellnes)

Rutile in eclogite (Engebøfjellet)

Perovskite in pyroxenite (Iron Hill)

Anatase in weathered rock

 (Catalão, Salitre,Tapira)

Deposit type

TiO2 content, in weight percent

Rutile sands

Ilmenite sands

EXPLANATION

Titanium Fig 4.

Source of titanium

deposits, such as those at Lac Tio and Tellnes, are thought to 

have formed by early magmatic crystallization of ilmenite 

(Duchesne, 1999; Charlier and others, 2010). Deposits 

dominated by titaniferous magnetite in layered intrusions may 

represent late-stage concentrations of iron and titanium in 

magmatic liquids following extensive silicate crystallization.

When rocks are affected by metasomatism (a process that 

changes the chemical composition of rocks by the introduction 

of external fluids), titanium minerals in the host rock may be 

altered. For example, at the Bamble-Arendal iron-titanium 

province of Norway, Proterozoic-age mafic rocks were 

affected by local and regional alkali-metasomatism during 

which original ilmenite was converted to rutile by the removal 

of iron (Korneliussen and others, 2000a). In this district, rutile 

was mined at Kagerø from a rutile-rich vein (Force, 1991). At 

Roseland, Va., rutile that occurs in metasomatized shear zones 

along complex contacts of anorthosite/ferrogabbro intrusions 

and dikes with adjacent country rocks was mined from 

1900 to 1949 (Herz and Force, 1987).

Concentrations of titanium minerals also occur in carbonatites (unusual rocks of magmatic or metasomatic origin 

dominated by carbonate minerals) and alkali igneous rocks. 

The Iron Hill (Powderhorn) carbonatite complex in Colorado 

is a series of successively intruded alkali rocks, including 

such rock types as carbonatite, nepheline, pyroxenite, and 

syenite (Temple and Grogan, 1965). Mainly because it 

contains abundant amounts of perovskite (a calcium-titanium 

oxide), the pyroxenite ranges from 1.9 to 9.6 weight percent 

TiO2 (with an average of about 5.3 weight percent TiO2) and 

significant amounts of rare-earth elements, thorium, niobium, 

and vanadium (Van Gosen, 2009).

In another carbonatite-hosted setting, extensive titanium 

(plus vanadium and niobium) mineralization is related to 

intrusion of the Magnet Cove carbonatite-alkalic ring dike 

complex in Arkansas, and accompanying metasomatic events 

(Fryklund and Holbrook, 1950; Flohr, 1994). At the Magnet 

Cove deposit, rutile occurs in feldspar-carbonate veins. Other 

deposits at Magnet Cove that formed by contact metasomatism 

include the Christy titanium-niobium-vanadium deposit and 

the Hardy-Walsh prospect. Both deposits contain brookite 

in novaculite (chert) where country rock was intruded by 

and reacted with titanium-, vanadium-, and niobium-bearing 

fluids from adjacent alkali-rich rocks (Erickson and Blade, 

1963; Flohr, 1994). Brookite at the Christy deposit can 

contain exceptionally high amounts of niobium and vanadium 

(Flohr, 1994).

Highly weathered rocks, such as bauxites (white-gray 

rock dominated by aluminum-rich oxides and hydroxides) and 

laterites (reddish rock dominated by iron- and aluminum-rich 

oxides and clays), can contain significant titanium resources. 

The titanium deposits at Catalão, Salitre, and Tapira (Brazil) 

formed by weathering of perovskite, magnetite, and titanite in 

titaniferous carbonatite, resulting in development of anatase in 

an about 200-meter (m)-thick laterite cap (Gomes and others, 

1990; Force, 1991). Significant amounts of rutile and ilmenite 

occur in saprolite (weathered rock) over mineralized contacts 

of the ilmenite-bearing Roseland anorthosite pluton (Virginia), 

with minor transport and accumulation of both minerals in 

streams that cut through the saprolite (Fish, 1962).

Titanium Deposits Related to Metamorphic Rocks

Rutile is the most economically valuable mineral for 

titanium, but rutile sand ores are limited in volume, and, given 

that there are inadequate supplies of high-grade (>55 percent 

TiO2 ) ilmenite sands, much of the recent search for high-grade 

ore has focused on rutile potential from metamorphic rocks 

(Korneliussen and others, 2000a). The main exploration 

focus is on high-grade metamorphic rocks, such as eclogite 

(a rock composed primarily of garnet and pyroxene, with 

lesser amounts of rutile, and typically formed under highpressure conditions). Based on surveys of eclogite in Norway, 

Korneliussen and others (2000b) state that metamorphism 

of titanium-rich protoliths (the unmetamorphosed rock from 
Geology T9

which a metamorphic rock forms) is a requirement to generate 

sufficient rutile to be of economic interest. Complications 

for exploitation of eclogite include highly variable rutile 

concentrations and grain sizes, retrograde alteration of 

high-grade metamorphic rutile back to ilmenite or titanite, 

and the possible presence of impurities, such as calcium, 

magnesium, iron, and aluminum in trace silicates in rutile 

concentrates (Korneliussen and others, 2000b). For example, 

in northern Italy, the Piampaludo eclogite has high TiO2

content (about 6 weight percent), but during later metamorphism, some rutile was replaced by secondary ilmenite and 

titanite (Liou and others, 1998), making this resource less 

economically attractive. The only currently active rutile 

mines in metamorphic rocks are the Daixian rutile deposits 

in east-central China, which consist of lenses and layers of 

rutile-bearing high-grade metamorphic rocks that average 

about 2 weight percent TiO2 (Shi and others, 2012).

Intense alteration by hydrothermal fluids reacting with 

large volumes of rock can result in profound changes in 

the original bulk rock composition, creating aluminum-rich 

rocks depleted in most other elements. Metamorphism of 

these hydrothermally altered rocks can create a rock rich in 

the minerals kyanite (Al2SiO5) and quartz and containing 

trace amounts of rutile (Marsh and Sheridan, 1976; Owens 

and Pasek, 2007). In addition, metamorphism of deeply 

weathered rocks, such as laterites and bauxites, can also 

result in rutile-bearing aluminum-rich rocks. Because these 

aluminum-rich rocks are stripped of many other elements, they 

potentially create a low-grade (about 1 percent or less rutile) 

and commonly low-tonnage but high-quality TiO2 resource that 

is generally free of other contaminants (Force, 1991). Rutile in 

such deposits can be a valuable byproduct of primary industrial 

commodities, such as alumina, andalusite, kaolinite, or kyanite.

Rutile also commonly occurs in mineralized zones and 

alteration halos of many hydrothermal and metamorphosed 

mineral deposits (Czamanske and others, 1981; Force, 1991; 

Clark and Williams-Jones, 2004). Large bodies of igneous 

rock that contain disseminated chalcopyrite and other sulfides 

are known as porphyry copper deposits. In the alteration zones 

around some porphyry copper deposits, such as Bingham 

Canyon (United States) or El Teniente (Chile), rutile can form 

from the breakdown of titanium-bearing silicates and oxides 

by the introduction of high-temperature hydrothermal fluids 

into the surrounding rock (Czamanske and others, 1981; 

Rabbia and others, 2009).

Titanium Deposits Related to Coastal Shorelines 

and Dunes, and Older Equivalent Rocks

Much of the high-grade ilmenite and rutile used today 

is extracted from unconsolidated heavy minerals (minerals 

that typically have a specific gravity higher than 2.85 grams 

per cubic centimeter) in shoreline beach placers and concentrations in sand dunes. Heavy-mineral sand deposits develop 

when relatively resistant titanium-bearing minerals and 

other heavy minerals are eroded from primary parent rocks, 

then transported and sorted, and finally deposited in fluvial, 

beach, or eolian settings. In addition to rutile and ilmenite 

(and minerals produced from weathering of ilmenite, such as 

anatase, leucoxene, and pseudorutile), other heavy minerals 

may enhance or detract from a deposit’s value. For example, 

zircon is a valuable commodity for the ceramics industry, 

whereas monazite (a phosphate mineral that can contain 

rare-earth elements and thorium) can make a deposit more 

problematic because of potential radioactivity issues with 

waste disposal.

Fluvial deposits, as used here, consist of heavy minerals 

deposited by streams and rivers above sea level, including 

deltaic deposition in rivers and lakes. Fluvial heavy-mineral 

deposits in old drainage courses in the Ghangbama district in 

Sierra Leone (for example, Mogbwemo and Sherbo River) 

are dominated by rutile, and contain lesser amounts of zircon 

and ilmenite; the deposits have been eroded from deeply 

weathered garnet amphibolite and leucocratic garnet granulite 

highlands (Force, 1991). Many fluvial deposits are small and 

geologically young and have a heavy-mineral assemblage 

that closely resembles that of the unaltered source rock 

(Force, 1991).

The abundance and distribution of heavy minerals in 

active and older (mostly Pleistocene) beach placer and eolian 

deposits are related to the area’s geology, physiography, 

and coastal dynamics. Many productive heavy-mineral sand 

provinces have the following characteristics: (a) deposits are 

located on passive continental margins that almost always are 

backed by elevated (typically highly weathered) high-grade 

metamorphic or mafic igneous hinterlands that supply detritus 

to the sands; (b) shoreline regions are dominated by sea-level 

changes and relatively low rates of sedimentation; (c) sands 

have been repeatedly reworked during marine transgressions 

and (or) by high-energy waves and littoral drifting; and 

(d) many sand bodies form topographic highs that represent 

former barrier island and (or) eolian dunes (Force, 1991; 

Hamilton, 1995; Roy, 1999; Pirkle and others, 2007; Hou 

and others, 2017).

Iron-titanium oxides in older, lithified heavy-mineral sand 

deposits may have undergone extensive post-depositional alteration, which diminishes the proportion of nonheavy minerals 

and enhances the conversion of ilmenite to leucoxene, causing 

an upgrade of the titanium content (Force, 1991). Ilmenite-rich 

layers of the Tertiary Cohansey Formation in New Jersey were 

mined at the Lakehurst deposit from 1962 until 1978 (Puffer 

and Cousminer, 1982). The Bothaville deposit in Late Paleozoic 

sandstones of the Ecca Group in the Karoo basin, South Africa, 

is a lithified paleo-beach deposit that contains (in order of 

abundance) ilmenite, rutile, and anatase (Hamilton, 1995). 

Heavy-mineral placers in shorelines of the Eocene-Pliocene 

Eucla basin (Hou and others, 2011) and the Pliocene Murray 

basin (Roy and Whitehouse, 2003) along the southern coast 

of Australia are major heavy-mineral sand provinces.
T10 Critical Mineral Resources of the United States—Titanium

A grade-and-tonnage model for shoreline placer titanium 

was developed by Attanasi and DeYoung (1992). Titanium 

grade and tonnage data for these deposits were combined 

with grade and tonnage data for titanium deposits related 

to both igneous and metamorphic rocks to create figure T5. 

The data for figure T5 are taken from disparate sources and 

represent a wide variety of deposit types and settings. Many 

of the deposits depicted in figure T5 are not economic. The 

high-grade and low-tonnage igneous deposits are typically 

characterized by the presence of abundant ilmenite and hemoilmenite (for example, Coulombe East and West, Bignel, and 

Lac Brûlé, all in the Grenville Province, Canada). Low-grade 

and high-tonnage igneous deposits are dominantly titanomagnetite (for example, Magpie Mountain and Panzhihua). 

Anatase is the titanium mineral at Salitre and Tapira, and 

perovskite is the titanium mineral at Iron Hill (Powderhorn), 

Colorado. Manavalakurichi is a heavy-mineral deposit in the 

State of Tamil Nadu, India.

0.1 1 10 100 1,000 10,000

0

10

20

30

40

Deposit tonnage, in million metric tons

Grade, in weight percent TiO2

Lac Tio

Salitre and Tapira

Panzhihua

Tellnes

Magpie Mountain

EXPLANATION

Titanium Fig 5

Igneous

Metamorphic

Sedimentary

Figure T5. Plot of titanium dioxide grade and tonnage for selected igneous, metamorphic, 

and sedimentary deposits in the world, by deposit type. Data were compiled from Towner 

and others (1988), Attanasi and DeYoung (1992), Corriveau and others (2007), and the 

Fennoscandian Ore Deposit Database (FODD), which is maintained by the Geological Survey 

of Finland (2013). Deposits extracted from the FODD have titanium, titanium and vanadium, 

or iron and titanium listed as their main metals. Some of the major deposits discussed in 

the text are labeled.
Resources and Production T11

Resources and Production

Identified Resources

Figure T6 shows world ilmenite and rutile mine production and reserves for 2012 (based on the data available), 

which are reported as titanium mineral concentrates by the 

U.S. Geological Survey (Bedinger, 2013). A major difficulty 

in quantifying titanium resources is the range of standards that 

individual countries and companies apply to the categorization 

of resources (Murphy and Frick, 2006). Resource calculations 

also depend on complex mineralogical factors, such as ore 

mineralogy, grain size, and trace element concentrations, 

which are all critical to a determination of whether a deposit 

can be mined economically or not. Because of the widespread 

distribution of titanium minerals in many geologic environments, development of new technologies to extract titanium 

from nonconventional ore minerals and deposit types has great 

potential to open up large, currently untapped resources.

Australia 15% Brazil 1%

Canada 11%

China 11%

India 9%

Madagascar 4%

Mozambique 6%

Norway 6%

Other countries 1%

Sri Lanka 1%

Ukraine 5%

United States 5%

Vietnam 8%

Australia 15%

Brazil 7%

Canada 5%

China 31%

India 13%

Madagascar 6%

Mozambique 2%

Norway 6%

South Africa 10%

Ukraine 1%

United States <1%

Vietnam <1%

Other countries 4%

Australia 58%

Brazil 1% India 3%

Mozambique 1%

Sierra Leone 12%

Ukraine 7%

Other countries 2%

Brazil 3%

Mozambique 1%

Other

countries 1% Ukraine 6%

A. World ilmenite production, 2012

(6.2 million metric tons)

B. World rutile production, 2012

(830,000 metric tons)

C. World reserves of ilmenite, 2012

(650 million metric tons)

D. World reserves of rutile, 2012

(42 million metric tons)

Titanium Fig 6.

Australia 43%

Sierra Leone 9%

India

17% 

South

Africa

20%

South

Africa

16%

South

Africa

17%

Figure T6. Pie charts showing percentage of estimated 2012 world mine production of A, ilmenite, and B, rutile, and 

percentage of global reserves of C, ilmenite, and D, rutile, by country. Total world mine production and reserves are also 

shown. Ilmenite data from the United States include rutile to avoid disclosing company proprietary data. Ilmenite reserves 

for Sri Lanka are not available. Data are from Bedinger (2013). Percentages are rounded and may not add to 100. The term 

reserves as used here is defined as that part of the reserve base that would be economically extracted or produced at the 

time of determination (U.S. Bureau of Mines and U.S. Geological Survey, 1980).
T12 Critical Mineral Resources of the United States—Titanium

The United States has large resources of ilmenite and 

perovskite, as well as the potential for rutile resources. 

Table T2 lists the titanium mineral resources (including 

mineral reserves) of the United States. The East Coast of 

the United States is part of the eastern North America 

titanium province, as described by Stanaway (1996). 

Although deposits in the region include titanium oxides 

related to anorthosite and gabbro in New York (Sanford 

Lake district) and Virginia (Piney River [Roseland] 

anorthosite complex), the bulk of the resource is contained 

in heavy-mineral concentrations in beach, bar, dune, and 

stream sands along the Atlantic Coastal Plain and in large 

parts of the Gulf Coastal Plain (fig. T3; Pirkle and others, 

2007). Nonetheless, the estimated amount of more than 

100 million metric tons of contained TiO2 resources in the 

United States given in table T2 represents less than 10 percent 

of the world’s titanium resources. The low grade of some 

U.S. resources, the possible presence of unfavorable trace 

constituents in ore, challenges with mineral separation 

and beneficiation of such nonconventional ore minerals as 

titaniferous magnetite and perovskite, anticipated high costs 

of permitting and production, and competition from foreign 

sources continue to curtail domestic production.

Total world resources of anatase, ilmenite, and rutile are 

estimated to be more than 2 billion metric tons (Bedinger, 

2013). Compiled world resource estimates given in table T3 

total more than this amount but that is because they include 

hypothetical and speculative resources in addition to identified 

resources. Rose (1969), Towner and others (1988), and Force 

(1991) provide detailed geologic and location information for 

a number of deposits included in the estimates in table T3. 

Important points from table T3 are (a) the predominance of 

ilmenite resources versus rutile resources, and (b) the relative 

proportions of heavy-mineral sand resources compared with 

igneous and metamorphic sources. Discounting the speculative 

estimate of greater than 2,000 million metric tons of ilmenite 

resources in Canada, which is largely based on estimates 

from Gross (1996), heavy-mineral sand deposits represent the 

largest resource type for worldwide titanium supplies.

Undiscovered Resources

Undiscovered (hypothetical and speculative) resources 

of titanium are considered to be very large but difficult to 

quantify. Ilmenite is a common mineral in mafic and ultramafic rocks, and rutile is a highly stable mineral in metamorphic rocks. Many of the resources identified in table T3 

include hypothetical and speculative titanium resources that 

await technological advances or favorable economics before 

they are likely to be exploited. For example, Rose (1969) 

states that high-grade magmatic ilmenite and titaniferous 

magnetite resources in Canada are measurable in hundreds of 

millions of tons, whereas low-grade deposits are potentially 

available in billions of tons. The Alberta oil sands also 

represent a very large speculative source of titanium (Towner 

and others, 1988). Although heavy minerals do not occur in 

large concentrations in these sands, the large quantity of sand 

processed for hydrocarbons creates a potentially significant 

titanium resource. The hot-water extraction process that 

recovers bitumen from the oil sands concentrates heavy 

minerals in treatment tailings, which contain on average 

about 11.5 percent TiO2

 (Liu and others, 2006). Attempts to 

isolate a high-grade titanium concentrate from these tailings 

have met with only limited success, however (Kaminsky and 

others, 2008).

About 500 million metric tons of black sands on the 

North Island of New Zealand contain titanomagnetite 

(which averages about 23 percent of the raw sand) with 

a TiO2 content of 7 to 8 weight percent (Christie and 

Braithwaite, 1998). The North Island black sands are 

currently mined for iron with vanadium as a byproduct, but 

they are not yet considered economic for titanium (Christie 

and others, 2000). Magmatic iron-titanium-vanadium deposits 

in layered anorthosite-gabbro rocks that contain abundant 

titaniferous magnetite but only minor amounts of ilmenite 

represent another very large potential source of titanium. 

Among numerous examples are the Magpie Mountain 

deposits (Canada) (which hosts an estimated 1.5 billion 

metric tons of titaniferous magnetite ore containing 

45.37 weight percent iron and 10.8 weight percent titanium; 

Gross, 1996) and the large Suwalki district (Poland), which 

contains significant resources in terms of tonnage that are 

not economic at this time (Charlier and others, 2009). As 

with perovskite at Iron Hill (Powderhorn), however, titanium 

mineralogy rather than the titanium content determines the 

economic value of the deposit.

Other speculative resources include very large but very 

low-grade rutile deposits in porphyry copper systems, which 

Force (1991) suggests have great potential because of the 

purity and ubiquitous nature of rutile in altered rock. In addition, shallow underwater offshore placers paralleling known 

exposed coastal placers are sparsely developed and largely 

unexplored but in some areas are known to contain commercially exploitable heavy minerals. Offshore placers represent 

large speculative titanium resources (Grosz and others, 1986).
Resources and Production T13

Table T2. Titanium resources of the United States.

[Values are from Force and Lynd (1984) with some updates from Force and Creely (2000), except that the Virginia heavy-mineral sands estimate 

is from Newton and Romeo (2006). Deposit classes are taken from table T1; —, none reported, or negligible; ?, identification uncertain]

State District, deposit,

or deposit type

Deposit

class

Contained titanium dioxide, by primary ore mineral 

(thousand metric tons)

Rutile Ilmenite Perovskite

Alabama Cretaceous sands S–3 — 100 —

Arizona Porphyry copper ore I/M 4,000 100

Arkansas Magnet Cove I–7 200 — —

California San Gabriel Mountains I–2 — 4,800 —

California Ione placer S–1 — 4,000 —

California White Mountains M–5 300 — —

Colorado Iron Hill (Powderhorn) I–7 — — 20,000

Colorado Evergreen M–5 200 — —

Florida Old beach sands S–2/S–4 1,100 9,700 —

Georgia Heavy-mineral sands S–2 500 2,400 —

Maryland Dinning M–3? 700 — —

Minnesota Duluth Complex I–2 — 10,000 —

Mississippi Ship Island S–2 — 100 —

New Jersey Lakehurst S–3 — 10,100 —

New Mexico Cretaceous sands S–3 — 700 —

New York Sanford Lake district I–2 — 8,600 —

New York Port Leyden S–1 — 6,300 —

North Carolina Yadkin Valley M–2? — 200 —

North Carolina Heavy-mineral sands S–2 — 400 —

Oklahoma Wichita Mountains S–1 — — —

South Carolina Heavy-mineral sands S–2 200 1,400 —

Tennessee Cretaceous sands S–3 1,300 8,400 —

Utah Porphyry copper ore I/M 4,000 — —

Utah Escalante S–3 — 200 —

Virginia Heavy-mineral sands S–2 150 2,200 —

Virginia Roseland I–5/I–8 1,000 5,500 —

Virginia Willis Mountain kyanite M–5 300 — —

Wyoming Laramie Range I–2 — 2,700 —

Wyoming Cretaceous sands S–3 — 500 —

Total1 14,000 77,900 20,000

1

The U.S. resource total for titanium dioxide contained in the three primary ore minerals (rutile, ilmenite, and perovskite) is about 112 million 

metric tons of titanium dioxide. Totals have been rounded.
T14 Critical Mineral Resources of the United States—Titanium

Table T3. Titanium mineral resources of the world (excluding the United States) for ilmenite (including titanomagnetite and 

leucoxene) and rutile (including anatase and brookite).

[Numbers are estimates. Deposit classes are taken from table T1; S, sedimentary, primarily heavy-mineral sands; I, igneous, primarily ilmenite and (or) titanomagnetite in mafic rocks; M, metamorphic, primarily rutile in eclogite; —, none reported or negligible; ?, amount uncertain; 

~, approximately. Values for heavy-mineral sand resources (deposit class S), with the exceptions of New Zealand and Russia, are from Elsner 

(2010); deposit class S values for New Zealand from Christie and Braithwaite (1998) are for ilmenite on the South Island and do not include 

titanium in the 500-million-metric-ton titanomagnetite iron sand resources of the North Island; deposit class I values for Brazil, Italy, and Mexico 

are from Force (1991); deposit class I values for Canada are generalized from Rose (1969); deposit class I values for China are calculated from 

China Mining Association (2006); deposit class M values for China are from Shi and others (2012); deposit classes I and M values for Norway 

are from Korneliussen and others (2000a); deposit classes I and S values for Russia are from Bykhovskii and Tigunov (2011)]

Country Deposit type Deposit

class

Contained titanium dioxide, 

by primary ore mineral 

(million metric tons)

Rutile Ilmenite/

titanomagnetite

Australia Heavy-mineral sands S 39.9 246

Brazil Heavy-mineral sands S 0.1 7

Brazil Anatase in weathered rock I 150 —

Cameroon Heavy-mineral sands S 2.9 —

Canada Heavy minerals in oil sands S — 183?

Canada Gabbro/anorthosite I — >2,000

China Heavy-mineral sands S 0.7 22

China Titanomagnetite in layered mafic intrusions I — ~52

China Rutile in anthophyllite M 6 —

Egypt Heavy-mineral sands S 0.5 —

India Heavy-mineral sands S >18.0 >348

Italy Rutile in eclogite M 9 —

Kazakhstan Heavy-mineral sands S 0.3 ~3

Kenya Heavy-mineral sands S 3 61

Madagascar Heavy-mineral sands S — 60

Malawi Heavy-mineral sands S >13 190

Malaysia Heavy-mineral sands S — ~10

Mexico Metasomatized anorthositic rocks I — 3

Mozambique Heavy-mineral sands S 3.2 237

Namibia Heavy-mineral sands S 0.6 36

New Zealand Heavy-mineral sands S — 5

Norway Gabbro/anorthosite I — 238

Norway Rutile in metamorphic/metasomatic rocks M 61 —

Russia Titanomagnetite/ilmenite/titanite I — 121

Russia Heavy-mineral sands S 2.3 9

Senegal Heavy-mineral sands S 0.6 22

Sierra Leone Rutile/ilmenite in fluvial deposits S 10.2 5

South Africa Heavy-mineral sands S 4.4 82

Sri Lanka Heavy-mineral sands S 0.5 ~12

Ukraine Heavy-mineral sands S 0.5 ~14

Vietnam Heavy-mineral sands S — 8

Total1 ~325 ~3,975

1

The world resource total for rutile and ilmenite combined is about 4.3 billion metric tons, which includes hypothetical and speculative resources 

in addition to identified resources.
Environmental Considerations T15

Exploration for New Deposits

Heavy-mineral analysis is an effective exploration guide 

for titanium-bearing heavy-mineral concentrations (Hou 

and others, 2017). In addition to high levels of titanium, 

geochemical signatures include high levels of zirconium, rareearth elements, thorium, and uranium (Grosz and Schruben, 

1994). For example, data indicating anomalously high levels 

of hafnium, rare-earth elements, titanium, and zirconium for 

streambed sediments collected for the U.S. Department of 

Energy’s National Uranium Resource Evaluation program 

helped locate major heavy-mineral deposits in an area between 

the Piedmont and the Atlantic Coastal Plain in Virginia and 

North Carolina (Grosz and Schruben, 1994). Exploration for 

fluvial placer deposits is best focused on source rock lithology, 

as many deposits are proximal and reflect the mineralogy of 

their source rocks (Force, 1991). Thus, a rutile source is more 

promising than an ilmenite source. More recent exploration for 

heavy-mineral deposits has relied on increasingly sophisticated 

radiometric methods and on stratigraphic and sedimentological 

modeling (Hou and others, 2017). Multichannel gamma-ray 

spectrometers coupled with geologic and land-use and landcover data have allowed recognition of anomalies that could 

be tied to thorium radiation from monazite (Pirkle and others, 

2007). In the Eucla basin in southern Australia, high-resolution 

satellite nighttime thermal imagery and detailed elevation data 

were combined with geologic data to identify beach-barrier 

facies and favorable heavy-mineral assemblages (Hou and 

others 2011, 2017). In most cases, modern beaches or interpreted older beaches must be cored to determine the presence 

of heavy minerals, and samples have to be characterized by 

mineralogy, grain size, and titanium content (Pirkle and others, 

2007). The large induced polarization response of ilmenite 

permits exploration for offshore marine heavy-mineral 

deposits using shipborne geophysical methods (Wynn, 1988). 

Innovative and modern exploration marine exploration 

methods, such as induced polarization, could cover a large 

swath of the continental shelf (Wynn, 1988).

Exploration for magmatic iron-titanium-oxide deposits 

is not only focused on grade and tonnage, but also on finding 

ilmenite suitable for pigment processing (for example, 

chromium- and magnesium-poor ilmenite). Suitable 

mineralogy and chemistry depend on relationships among 

magmatic evolution, major- and trace-element chemistry, 

and low temperature re-equilibration. Geophysical methods 

have been used for exploration and to extend known 

deposits. Prospecting for workable and chemically suitable 

ilmenite resources relies in part on aeromagnetic signatures 

of iron-titanium-oxides that record ore microtextures that 

can be related back to ore chemistry (McEnroe and others, 

2000). Aeromagnetic maps over favorable terranes may 

have a spectacular range of positive and negative anomalies 

because of highly contrasting rocks and possible deposit types 

(McEnroe and others, 2001). Concentrations of nonmagnetic 

hemo-ilmenite commonly have distinctive negative magnetic 

anomalies or mixed patterns of positive and negative magnetic 

anomalies. Strong positive anomalies related to magnetite 

content may point toward large titanomagnetite concentrations. Because of contrasting physical properties compared 

to silicates, iron-titanium-deposits commonly have higher 

gravity anomalies than surrounding anorthositic, granitic, or 

gneissic country rock (Gross, 1996).

Exploration for rutile in metamorphic rocks, specifically eclogite, depends on recognition of favorable geologic 

circumstances that would create high-grade and high-quality 

rutile concentrations. These favorable circumstances include 

(a) high-grade metamorphism of large volumes of titaniumbearing gabbroic protoliths; (b) favorable mineralogical 

factors, such as grain size, mineral intergrowths, modal 

variations, and degree of retrograde mineral alteration; 

(c) favorable locations with regard to deposit development, 

shipping, and waste treatment; and (d) potential byproducts 

(for example, garnet) that could enhance the economics of a 

deposit (Korneliussen and others, 2000b).

Environmental Considerations

Sources and Fate in the Environment

The behavior of titanium in the environment is 

dominated by its low solubility and the limited reactivity 

of titanium minerals. The main sources of anthropogenic 

titanium in the environment are from the combustion of 

fossil fuels and, to a lesser extent, incineration of titaniumcontaining products (International Programme on Chemical 

Safety, 1982). The mean concentration of titanium in soils 

worldwide is 0.33 weight percent titanium (Kabata-Pendias 

and Mukherjee, 2007); the average and range in titanium 

concentrations for soils in the conterminous United States 

are 0.29 weight percent and 0.007 to 2.0 weight percent, 

respectively (Shacklette and Boerngen, 1984). Titanium in 

stream waters worldwide ranges from 12 to 926 micrograms 

per liter (μg/L) (Reimann and de Caritat (1998). River waters 

in North America contain between 2 and 107 μg/L titanium, 

and drinking waters in the United States contain from 0.5 

to 15 μg/L titanium (International Programme on Chemical 

Safety, 1982). The average concentration of titanium in 

suspended sediments in rivers worldwide is 0.44 weight 

percent (Viers and others, 2009). In the atmosphere, titanium 

ranges from 0.5 to 2.5 nanograms per cubic meter (ng/m3) 

for remote areas and from 15 to 25 ng/m3 for urban areas 

(Kabata-Pendias and Mukherjee, 2007). Atmospheric 

concentrations of titanium in urban areas of the United States 

are typically below 100 ng/m3, but values of up to 1,000 ng/m3

have been reported in some industrial regions (International 

Programme on Chemical Safety, 1982).

A significant amount of the titanium manufactured as a 

metal is recycled. For TiO2 used for pigment, however, there is 

no role for recycling because the compound is so profoundly 
T16 Critical Mineral Resources of the United States—Titanium

incorporated into the end product that it cannot be recovered 

with today’s technology (Murphy and Frick, 2006). Because 

the bulk of the world’s titanium is used for pigment, recycling 

currently plays little or no role in titanium consumption or 

pricing trends.

Mine Waste Characteristics

Magmatic ilmenite and sedimentary-related titanium 

deposits produce very different volumes and types of mine 

waste. The open pit methods used to extract ore from massive 

magmatic ilmenite deposits produce large volumes of waste 

rock and tailings. The Storgangen Mine (Norway), which 

operated from 1917 to 1965, produced more than 10 million 

metric tons of ore (Schiellerup and others, 2003) and more 

than 8 million metric tons of tailings (Ettner, 1999). Annual 

production (1999) from the Tellnes open pit mine (Norway) 

is about 2 million metric tons of ore and 1.6 million metric 

tons of waste rock (Norges Geologiske Undersøkelse, 2012). 

Tailings from the Tellnes Mine are composed predominantly 

of feldspar, ilmenite, and pyroxene with a size range of 

10 to 200 micrometers (μm) (Olsgard and Hasle, 1993). 

Mining of the Lac Tio deposit (Quebec, Canada) began in the 

early 1950s. Waste-rock piles at Lac Tio are estimated to be 

more than 72 million metric tons and to cover approximately 

100 hectares of land, and they are between 20 and 80 m 

in height (Pepin and others, 2008). Waste rock at Lac Tio 

is dominantly plagioclase, followed by lesser amounts of 

hemo-ilmenite and trace sulfides (Pepin and others, 2008). 

Pyrite is the main sulfide (1 to 2 weight percent) in waste rock 

with only trace amounts (<1 weight percent) of chalcopyrite 

(Plante and others, 2011). Laboratory tests on the Lac Tio 

Mine waste to approximate weathering produced neutral to 

slightly alkaline leachate (pH 6.5 to 9.0), which is consistent 

with the near-neutral natural drainage from the waste piles 

(Plante and others, 2011). Field monitoring of experimental 

mine-waste cells produced drainage with a pH of between 

4 and 8, which mimicked the pH of rain water (Pepin and 

others, 2008). Thus, any acid generated by sulfides in mine 

waste from the Lac Tio Mine is neutralized by silicates; the net 

neutralization potentials range from – 6.0 to 3.1 kilograms (kg) 

of calcium carbonate (CaCO3) per metric ton of ore, all within 

the uncertainty zone of ±20 kg of CaCO3 per metric ton of ore 

for the test method (Plante and others, 2011).

Mining and milling methods for heavy-mineral sand 

deposits involve physical separation of a bulk concentrate and 

quartz-rich tailings by mechanical means, typically magnetic 

or electrostatic methods or density separation. During further 

concentration, the mineral-bearing sand fraction is separated 

from finer textured slimes (clays, silts, very fine sand), 

which is mixed with a flocculent to aid settling, and then 

pumped back as a slurry into a reclamation pit (Daniels and 

others, 2003).

The average grain size of heavy minerals and the 

percentage of slimes in a deposit are variable, which results in 

different amounts and types of waste material. For example, 

two mineral-sand blocks in India (Kuttam and Sattankulam) 

contain up to 10 percent heavy minerals and 15 percent 

slimes (Murty and others, 2007). The sands in these blocks 

are very homogeneous—85 percent of the sands range from 

<2 millimeters to 63 μm in size. Heavy-mineral sands on the 

northeastern coast of Sri Lanka (Pulmoddai sands) are very 

high grade; 71 percent of the beach sands are smaller than 

355 μm, and more than 99 percent of the titanium content is 

in this size fraction (Premaratne and Rowson, 2003). The two 

blocks in India cover approximately 120 square kilometers 

(km2) and contain an estimated 400 million metric tons of raw 

sand and 30 million metric tons of ilmenite (Murty and others, 

2007). Pulmoddai sands cover an area of only 3.2 km2 with an 

even thickness of 6 m (Premaratne and Rowson, 2003); thus, 

the amount and type of waste generated by the two deposits 

in India would be very different than the waste generated 

at Pulmoddai.

Human Health Concerns

Human health concerns related to titanium are minimal. 

Titanium can enter the human body by ingestion or inhalation, 

but most ingested titanium is eliminated unabsorbed. An 

exception to this is some industrial inhalation exposures to 

high amounts of dust during titanium processing (KabataPendias and Mukherjee, 2007). Mining processes conducted 

on ilmenite ore or heavy-mineral sands may create high 

amounts of airborne dust, and control measures are needed to 

reduce workers’ exposure. The U.S. Environmental Protection 

Agency (EPA) has not found any human or animal studies 

that are useful for developing toxicity values for titanium 

(U.S. Environmental Protection Agency, 2010).

The processing of heavy-mineral sands at mine sites may 

expose workers to minor amounts of radiation. The radiation 

levels in heavy-mineral concentrates are usually too low 

to be classified as radioactive; however, when radioactive 

minerals, such as monazite or xenotime, are separated at 

the dry processing stage, precautions may need to be taken 

to protect employees from low levels of ionizing radiation 

(Abuodha, 2002).

The EPA and the World Health Organization (WHO) 

have not established drinking water guidelines for titanium 

nor has the EPA established soil screening levels for industrial 

or residential land use for remediation purposes. In fact, 

TiO2 replaced lead carbonate and lead oxide as the whiteners 

in paints in the United States because of its low toxicity 

to humans.

Ecological Health Concerns

The EPA also has not established water-quality criteria 

for the protection of aquatic life for titanium because titanium 

minerals are generally insoluble and not bioaccessible. 
Environmental Considerations T17

Recent findings do suggest that TiO2 nanoparticles, commonly 

defined as particles smaller than 100 nanometers in at least 

two dimensions, may be toxic to some aquatic and terrestrial 

organisms (Federici and others, 2007; Wang and others, 

2008; Sharma, 2009). Titanium dioxide nanoparticles are 

used in a range of products, including sunscreen, cosmetics, 

paint, and building materials, and have been shown to behave 

differently in the environment than larger particles of the same 

composition (Sharma, 2009). Applications for nanoparticles 

are increasing and, consequently, this form of titanium may be 

of increasing environmental concern in the future.

Titanium deposits and related waste products are usually 

sulfide-poor and generally generate near-neutral pH drainage 

after interacting with rain, surface water, or groundwater. 

Therefore, acid mine drainage, which is a common potential 

problem for many types of ore deposits, is not a significant 

concern at titanium mines. Although titanium may not be an 

environmental concern, other metals found in trace sulfides 

may be deleterious to the environment. Some metals, such as 

cobalt, nickel, and zinc, will remain in solution even at neutral 

pH. Drainage water with near-neutral pH and some quantity of 

dissolved metals is termed “contaminated neutral drainage.” 

Nickel values greater than 0.5 milligram per liter (mg/L) have 

been reported from waters draining waste piles at the Lac Tio 

Mine, at times exceeding U.S. and Canadian water-quality 

guidelines for the protection of aquatic life (Plante and others, 

2010). Initially, nickel was adsorbed by residual ilmenite and 

plagioclase in fresh waste rock, but as mine waste weathered, 

adsorption sites filled and a lowered adsorptive capacity 

resulted in increased nickel in drainage waters, creating 

contaminated neutral drainage (Plante and others, 2010). This 

interpretation is consistent with field observations on mine 

waste experimental cells at Lac Tio that produced drainage 

with higher concentrations of nickel (0.1 to 0.4 mg/L) from 

approximately 25-year-old mine waste compared with 

drainage concentrations of nickel (<0.05 mg/L) from freshly 

blasted waste rock (Pepin and others, 2008). These studies 

suggest that ecosystem health concerns from high nickel 

concentrations in mine drainage waters may increase as mine 

wastes weather.

Approximately 750,000 cubic meters of water drains 

annually from a tailings impoundment for the closed 

Storgangen Mine (Norway) (Ettner, 2007). Runoff from the 

former mine that subsequently reaches the North Sea at a 

distance of about 5 kilometers downstream has a neutral pH 

and nickel concentrations of up to 6 mg/L (Ettner, 2007). An 

anaerobic passive treatment system was installed in 2003 to 

remove nickel, and initially it had variable efficacy because 

of the immaturity of the system and extreme rain storms; since 

2007, however, the system has successfully removed nickel 

from the tailings runoff (Ettner, 2007).

Mine tailings from Tellnes are almost inert and are 

nontoxic to marine life (Olsgard and Hasle, 1993). Disposal 

of tailings underwater can lead to increased water turbidity, 

however, and, depending on the sedimentation rate, to the 

smothering of organisms living in bottom sediment. At 

Tellnes, about 2 million metric tons per year of tailings were 

deposited into a local fjord throughout the 1980s, which 

resulted in the reduction in number and diversity of species. 

After underwater tailings deposition ceased in 1994, bottom 

fauna in the fjord began to recover. Although there was a 

reduction in species richness compared with that of unaffected 

background locations for several years after tailings deposition 

ceased, within 4 years, all the main macrobenthic phyla 

were represented (Olsgard and Hasle, 1993). Some of the 

waste rock at the Lac Tio Mine is adjacent to a water body 

called Lac Petit Pas. A study by the operating company of the 

Lac Tio Mine showed that there were elevated concentrations 

of some metals in the first several centimeters of sediment 

within the lake but concluded that the sediments act as a 

retention medium (Rio Tinto, 2009).

Mine Closure

The post-mining landscape at titanium mines depends 

on the type of deposit exploited and the mining methods 

employed. Pit lakes remaining from open pit magmatic 

titanium deposits could contain low concentrations of trace 

elements typically found in ores, such as cobalt, chromium, 

nickel, vanadium, and zinc. In the Sanford Lake district of 

New York, the pit lake at the McIntyre Mine, which was 

closed in 1982, has a neutral pH of 8 and low to nondetectable 

concentrations of trace elements (Woodruff and others, 2013; 

S. Hollmeyer, undergraduate student, University of Vermont, 

unpub. data, December 2003).

Titanium minerals from heavy-mineral sand deposits 

are most commonly mined using surface methods, including 

wet dredging and dry mining techniques. Pits are backfilled 

with tailings and slimes. Commonly, topsoil initially removed 

at a site is stored during active mining and later mixed with 

tailings and slimes to facilitate the return of vegetation. 

Daniels and others (2003) examined soil profiles within 

reclamation pits at the Old Hickory Mine site (Virginia). 

Soils were a mixture of dewatered slimes and tailings with 

added soil amendments, such as biosolids, fertilizer, and lime. 

Amended surface soils from reclaimed pits had near neutral 

pH values and up to 1.5 percent organic matter. Reclaimed 

mine soils from the Green Cove Springs Mine in Florida are 

similar, with organic-rich topsoil-amended surface layers 

and quartz-rich tailings in the subsurface (Daniels and 

others, 1992).

Surface mining of mineral sands has the potential to 

disturb prime agricultural farmlands along the eastern coast 

of the United States (Daniels and others, 1992). Orndorff and 

others (2011) and Zimmerman and others (2008), however, 

suggest that disturbed lands around such sites could be 

successfully returned to local regional agricultural levels if 

best agricultural and science-based practices, such as replacement of topsoil, pH adjustments with lime, and addition of 

fertilizers and biosolids, are applied.
T18 Critical Mineral Resources of the United States—Titanium

Problems and Future Research

Titanium is an abundant element; it is found at concentrations higher than 20 percent in many minerals (see fig. T1). 

It is currently not possible to extract a high-grade titanium 

product from titanite and perovskite, however, and largescale recovery of titanium from low-titanium oxides, such 

as titanomagnetite, is mainly in the research stage. A role 

for future research for titanium resources revolves around 

the ability to develop cost-effective extraction methods for 

titanium minerals other than anatase, brookite, ilmenite, 

and rutile. Depletion of rutile resources and the shift from 

the sulfate process to the chloride process, which requires 

higher-grade TiO2 feedstock compared with the older sulfate 

process, increases the need to develop low-cost, energyefficient hydrometallurgical processes to, in effect, upgrade 

the more abundant ilmenite (Zhang and others, 2011). If these 

issues of mineralogy could be resolved, there are many large 

titanium resources—some of which are in advanced stages of 

exploration—that could become economic.

Rutile is in high demand, so the search for additional 

resources is likely to focus on exploration for large, high-grade 

domestic rutile resources. Increased examination of the myriad 

geologic environments where rutile occurs (see, for example, 

Meinhold, 2010) might make it possible to discover previously unrecognized or underutilized rutile sources, such as 

porphyry systems, metamorphic terranes, or highly weathered 

rock. Rutile as a magmatic mineral is described for some 

massif anorthosite ilmenite deposits in the Saint-Urbain and 

Lac Allard anorthosite plutonic suites (Quebec). Morisset and 

others (2010) attribute this occurrence to an exceptionally 

high alumina content of parent magmas and the subsequent 

oxide crystallization path. Rutile also occurs along the 

metasomatized margins of the Roseland anorthosite (Virginia) 

and in the Pluma Hidalgo anorthosite district (Mexico), 

where swarms of anorthosite dikes and sills cut country rock 

(Force, 1991). In these two deposits, anorthosite has a more 

alkali composition than is typical for most other anorthositehosted magmatic deposits, which Force (1991) links to the 

relative abundance of rutile. Thus, research into a possible 

relation between host anorthosite chemistry and occurrences 

of abundant rutile with ilmenite in magmatic deposits could 

result in delineation and identification of favorable environments for higher grade rutile-bearing magmatic deposits.

Recent research in anorthosite plutonic terranes that 

host ilmenite-dominated magmatic systems, such as Lac Tio 

and Tellnes, has led to recognition of the role that magmatic 

processes, such as contamination, mineral fractionation, 

and flow dynamics, play in the development of world-class 

titanium orebodies (Wilmart and others, 1989; Charlier and 

others, 2010). An updated deposit model for the formation of 

these types of titanium deposits (Woodruff and others, 2013) 

could guide exploration for this type of mineralization as either

a direct source or as protoliths for weathered ore deposits.

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Vanadium

By Karen D. Kelley, Clinton T. Scott, Désirée E. Polyak, and Bryn E. Kimball

Chapter U of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802– U

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter U

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................U1

Introduction..................................................................................................................................................U2

Uses and Applications ......................................................................................................................U2

Demand and Availability of Supply .................................................................................................U9

Geology.......................................................................................................................................................U10

Geochemistry....................................................................................................................................U10

Mineralogy........................................................................................................................................U10

Deposit Types ...................................................................................................................................U13

Vanadiferous Titanomagnetite Deposits .............................................................................U13

Sandstone-Hosted Vanadium Deposits ..............................................................................U14

Shale-Hosted Vanadium Deposits .......................................................................................U16

Vanadate Deposits ..................................................................................................................U17

Other Magmatic-Hydrothermal Vanadium Resources .....................................................U17

Fossil Fuels...............................................................................................................................U18

Resources and Production ......................................................................................................................U18

Identified Resources (United States and World) ........................................................................U18

Undiscovered Resources ...............................................................................................................U19

Exploration for New Deposits .................................................................................................................U20

Exploring for Vanadiferous Titanomagnetite Deposits ..............................................................U20

Exploring for Sandstone-Hosted Vanadium Deposits ................................................................U21

Exploring for Shale-Hosted Vanadium Deposits .........................................................................U21

Environmental Considerations ................................................................................................................U22

Sources and Fate in the Environment ..........................................................................................U22

Mine Waste Characteristics ..........................................................................................................U22

Human Health Concerns .................................................................................................................U24

Ecological Health Concerns ...........................................................................................................U25

Mine Closure.....................................................................................................................................U25

Problems and Future Research ..............................................................................................................U26

Acknowledgments....................................................................................................................................U26

References Cited.......................................................................................................................................U27

Suggested citation:

Kelley, K.D., Scott, C.T., Polyak, D.E., and Kimball, B.E., 2017, Vanadium, chap. U of Schulz, K.J., DeYoung, J.H., Jr.,

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. U1– U36, http://dx.doi.org/

10.3133/pp1802U.
Vanadium

By Karen D. Kelley, Clinton T. Scott, Désirée E. Polyak, and Bryn E. Kimball

Abstract

Vanadium is used primarily in the production of steel 

alloys; as a catalyst for the chemical industry; in the making of 

ceramics, glasses, and pigments; and in vanadium redox-flow 

batteries (VRBs) for large-scale storage of electricity. World 

vanadium resources in 2012 were estimated to be 63 million 

metric tons, which include about 14 million metric tons of 

reserves. The majority of the vanadium produced in 2012 

was from China, Russia, and South Africa.

Vanadium is extracted from several different types of 

mineral deposits and from fossil fuels. These deposits include 

vanadiferous titanomagnetite (VTM) deposits, sandstonehosted vanadium (with or without uranium) deposits (SSV 

deposits), and vanadium-rich black shales. VTM deposits are 

the principal source of vanadium and consist of magmatic 

accumulations of ilmenite and magnetite containing 

0.2 to 1 weight percent vanadium pentoxide (V2O5). SSV 

deposits are another important source; these deposits have 

average ore grades that range from 0.1 to greater than 1 weight 

percent V2O5. The United States has been and is currently the 

main producer of vanadium from SSV deposits, particularly 

those on the Colorado Plateau. Vanadium-rich black shales 

occur in marine successions that were deposited in epeiric 

(inland) seas and on continental margins. Concentrations in 

these shales regularly exceed 0.18 weight percent V2

O5 and 

can be as high as 1.7 weight percent V2O5. Small amounts 

of vanadium have been produced from the Alum Shale in 

Sweden and from ferrophosphorus slag generated during the 

reduction of phosphate to elemental phosphorus in ore from 

shales of the Phosphoria Formation in Idaho and Wyoming. 

Because vanadium enrichment occurs in beds that are typically only a few meters thick, most of the vanadiferous black 

shales are not currently economic, although they may become 

an important resource in the future. Significant amounts of 

vanadium are recovered as byproducts of petroleum refining, 

and processing of coal, tar sands, and oil shales may be 

important future sources.

Vanadium occurs in one of four oxidation states in nature: 

+2, +3, +4, and +5. The V3+ ion has an octahedral radius that 

is almost identical to that of Fe3+ and Al3+ and, therefore, it 

substitutes in ferromagnesian minerals. During weathering, 

much of the vanadium may partition into newly formed clay 

minerals, and it either remains in the +3 valence state or 

oxidizes to the +4 valence state, both of which are relatively 

insoluble. If erosion is insignificant but chemical leaching is 

intense, the residual material may be enriched in vanadium, 

as are some bauxites and laterites. During the weathering 

of igneous, residual, or sedimentary rocks, some vanadium 

oxidizes to the +5 valence state, especially in the intensive 

oxidizing conditions that are characteristic of arid climates.

The average contents of vanadium in the environment 

are as follows: soils (10 to 500 parts per million [ppm]); 

streams and rivers (0.2 to 2.9 parts per billion [ppb]); and 

coastal seawater (0.3 to 2.8 ppb). Concentrations of vanadium 

in soils (548 to 7,160 ppm) collected near vanadium mines in 

China, the Czech Republic, and South Africa are many times 

greater than natural concentrations in soils. Additionally, if 

deposits contain sulfide minerals such as chalcocite, pyrite, 

and sphalerite, high levels of acidity may be present if sulfide 

dissolution is not balanced by the presence of acid-neutralizing 

carbonate minerals. Some of the vanadium-bearing deposit 

types, particularly some SSV and black shale deposits, contain 

appreciable amounts of carbonate minerals, which lowers the 

acid-generation potential.

Vanadium is a micronutrient with a postulated requirement for humans of less than 10 micrograms per day, which 

can be met through dietary intake. Primary and secondary 

drinking water regulations for vanadium are not currently in 

place in the United States. Vanadium toxicity is thought to 

result from an intake of more than 10 to 20 milligrams per 

day. Vanadium is essential for some biological processes 

and organisms. For example, some nitrogen-fixing bacteria 

require vanadium for producing enzymes necessary to convert 

nitrogen from the atmosphere into ammonia, which is a more 

biologically accessible form of nitrogen.
U2 Critical Mineral Resources of the United States—Vanadium

Introduction

Vanadium (V) is a strategic metal that is used principally 

in the production of metal alloys, such as high-strength steel 

and alloys for use in the aerospace industry. Secondary uses 

are as catalysts for the chemical industry, and in ceramics, 

glasses, and pigments. In its native state, vanadium is a hard, 

silvery gray, ductile, and malleable transition metal. Vanadium 

consumption trends are heavily influenced by trends in 

steel production.

The emerging need for large-scale electricity storage 

makes vanadium redox-flow batteries (VRBs) a major 

potential future use of vanadium. Because of their largescale storage capacity, development of VRBs could prompt 

increases in the use of wind, solar, and other renewable, intermittent power sources. Lithium-vanadium-phosphate batteries 

produce high voltages and high energy-to-weight ratios, which 

make them ideal for use in electric cars. Vanadium use in 

lithium batteries is expected to increase to 1,700 metric tons 

in 2017 from 200 metric tons in 2012 (Perles, 2013).

Vanadium is the 22d most abundant element in Earth’s 

crust, and it is an essential constituent of many minerals. A 

total of 156 minerals contain vanadium as a major (>10 weight 

percent) constituent. Several diverse mineral deposit types 

contain vanadium-bearing minerals, and vanadium is a 

common component of petroleum and other fossil fuels. 

Vanadium deposits are globally distributed (fig. U1) and 

comprise four principal deposit types: vanadiferous titanomagnetite (VTM), sandstone-hosted vanadium (SSV), 

shale-hosted vanadium, and vanadate deposits (table U1). 

Additionally, significant amounts of vanadium are available 

for commercial use as a byproduct of petroleum refining, and 

processing of coal, tar sands, and oil shales may be important 

future sources. World vanadium resources in 2012 were 

estimated to be 63 million metric tons of vanadium. Reserves 

were estimated to be 14 million metric tons. The majority of 

vanadium production in 2012 was from China (37 percent), 

South Africa (35 percent), and Russia (25 percent) (fig. U2; 

Polyak, 2013).

Uses and Applications

The vanadium market closely follows that of the steel 

industry, which in turn follows economic trends. Metallurgical 

applications in steel continued to dominate United States 

vanadium usage in 2011 (fig. U3), accounting for 93 percent 

of reported consumption (Polyak, 2013). Vanadium is used 

in steel to impart strength, toughness, and wear resistance. 

The formation of vanadium-rich carbides and nitrides imparts 

the strength to steel; the addition of only a few kilograms of 

vanadium per ton of steel increases the strength of the steel by 

as much as 25 percent. Apart from its strengthening characteristic, vanadium also inhibits corrosion and oxidation.

There are many sources of vanadium, and it is used in a 

number of common products (fig. U4). Commercial products 

developed through the processing of vanadium ores are mainly 

ferrovanadium (FeV), which is an iron-vanadium alloy, and 

vanadium pentoxide (V2

O5 ). Most vanadium is added to steel 

as ferrovanadium. Ferrovanadium is available in compositions containing 45 to 50 percent vanadium and 80 percent 

vanadium. The 45- to 50-percent grade is produced from slag 

and other vanadium-bearing residues; the 80-percent grade is 

produced by the reduction of V2O5.

The high-strength, low-alloy (HSLA) steels containing 

vanadium are widely used for the construction of auto parts, 

buildings, bridges, cranes, pipelines, rail cars, ships, and 

truck bodies, including armor plating for military vehicles 

(Polyak, 2012). Such HSLA steels are increasingly being 

used in the oil and gas industry to meet demand for pipelines 

with higher strength and higher low-temperature toughness 

(Roskill Information Services, Ltd., 2010, p. 150). Vanadium 

is used in tool steels in various combinations with chromium, 

niobium (columbium), manganese, molybdenum, titanium, 

and tungsten. Only a limited degree of substitution is possible 

among these metals, however. Replacement of vanadium with 

other mineral commodities requires significant technical adjustments to the steel production process to ensure that product 

specifications and quality are not compromised. For example, 

use of vanadium generally requires less energy consumption 

during production than does niobium to give equivalent steel 

properties. Therefore, substitution for vanadium is normally not 

considered for short-term changes in market conditions because 

of the considerable effort involved in implementing the change.

Vanadium is irreplaceable for its role in aerospace 

applications because vanadium-titanium alloys have the 

best strength-to-weight ratio of any engineered material 

yet discovered. Vanadium, when combined with titanium, 

produces a stronger and more stable alloy, and when combined 

with aluminum produces a material suitable for jet engines 

and high-speed airframes. No acceptable substitutes exist for 

vanadium in aerospace titanium alloys.

Nonmetallurgical applications of vanadium include 

catalysts, ceramics, electronics, and vanadium chemicals. 

For catalytic uses, platinum and nickel can replace vanadium 

compounds in some chemical processes. Vanadium dioxide 

is used in the production of glass coatings that block 

infrared radiation.

Vanadium is becoming more widely used in green 

technology applications, especially in battery technology. 

One battery technology that continues to show promise in 

stabilizing energy distribution in renewable systems is the 

VRB, which consists of an assembly of power cells in which 

two vanadium-based electrolytes are separated by a proton 

exchange membrane. The main advantages of the VRBs are 

(a) their nearly unlimited capacity, which is made possible 

simply by using sequentially larger storage tanks; (b) their 

ability to be left completely discharged for long periods of time 

with no detrimental effects; (c) the ease of recharging them 

by replacing the electrolyte if no power source is available to 

charge it; and (d) their ability to withstand permanent damage 

if the electrolytes are accidentally mixed (Polyak, 2012).
Introduction U3

180° 135° 90° 45° 0° 45° 90° 135° 180°

60°50°40°30°20°10°

0°

10°20°30°40°50°

80°

70°

70°

80°

60°

EXPLANATION

Vanadium deposits—Larger  symbol represents deposit  with resources of >1 million  metric tons of V O2 5

Sandstone-hosted vanadiumShale-hosted vanadiumVanadateVanadiferous titanomagnetite

5

2

6

7 8

4 3

25

23

26

24

29

30

31

41 

40 42

51 45 52 39 46

9

48

43

44

49

47

50

32

58

60

59

61

27

28

36

37

35

38

55

53 54

56 33

34

80 81

57

1

17 18 14

19 20

21 15 16

12

13

22

63

62

77 7675 71

69

79

68

73

72

74 78

70 66

65

67

64

10

11

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection;

Vanadium, Fig 1.

 World Geodetic System 1984 datum

Figure U1. World map showing locations of major vanadium deposits of the world, by deposit type. Numbers refer to the identifier number in table U1, which contains additional information about the vanadium resources in these deposits.
U4 Critical Mineral Resources of the United States—Vanadium

Table U1. Location, grade, tonnage, and other data for selected vanadium deposits of the world.—Continued

[ID, identifier, shown in figs. U1 and U5; WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern 

Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; %, percent; V2O5, vanadium oxide; VTM, vanadiferous titanomagnetite; n.d., no data available; NA, not available; t, metric ton; SSV, sandstone-hosted vanadium]

ID Deposit Country

District, 

region,

or State/

Province

Deposit 

type

Latitude Longitude Resource

tonnage 

(million 

metric 

tons)

Grade

(% V2

O5)

V2

O5

content 

(million 

metric

tons)

Production, 

resources, 

or reserves

Source

(decimal degrees, of data

WGS 84)

1 Lueca Mine Angola NA Vanadate–7.92 13.65 n.d. n.d. n.d. n.d. Fischer

(1975a)

2 Windimurra Australia Mount 

Magnet

VTM –28.36 118.84 242.6 0.48 1.16 Resources Britt and others (2014)

3 Mount Peake Australia Northern 

Territory

VTM –21.66 133.66 160 0.3 0.48 Resources TNG Ltd. 

(2015)

4 Julia Creek Australia Queensland Shalehosted

–20.65 141.73 411 0.44 1.1 Resources Lewis and others (2010)

5 Balla Balla Australia Western 

Australia

VTM –21.14 118.06 456 0.64 2.91 Resources Britt and others (2014)

6 Speewah Australia Western 

Australia

VTM –16.47 128.20 4,712 0.3 14.1 Resources Britt and others (2014)

7 Gabanintha Australia Western 

Australia

VTM –27.00 118.60 125.8 0.7 0.88 Resources Britt and others (2014)

8 Barrambie Australia Western 

Australia

VTM –27.42 119.11 47.2 0.63 0.3 Resources Britt and others (2014)

9 Svorantskoe Azerbaijan Chelyabinskaya 

Oblast

VTM 39.00 46.00 n.d. n.d. n.d. n.d. Fischer 

(1975b)

10 Maracas Brazil Bahia VTM –13.41 –40.43 24.6 1.11 0.27 Resources Fischer 

(1975b)

11 Tapajos Brazil Pará VTM –3.61 –54.49 n.d. n.d. n.d. n.d. Fischer 

(1975b)

12 Banks Island Canada British

Columbia

VTM 53.47 –130.13 3 0.6 0.018 Resources Fischer 

(1975b)

13 Porcher 

Island

Canada British

Columbia

VTM 54.00 –130.25 n.d. n.d. n.d. n.d. Fischer 

(1975b)

14 Lac Doré Canada Matagami, 

Quebec

VTM 49.76 –77.62 100 0.49 0.49 Resources Fischer 

(1975b)

15 Lake Michikamau

Canada Newfoundland and 

Labrador

VTM 54.12 –63.98 n.d. n.d. n.d. n.d. Fischer 

(1975b)

16 Ring of Fire Canada Ontario VTM 52.80 –86.50 n.d. n.d. n.d. n.d. Fischer 

(1975b)

17 Lac Tio (Allard Lake)Canada Quebec VTM 50.56 63.41 350 0.3 1.05 Resources Fischer 

(1975b)

18 Sept Îles Canada Quebec VTM 50.34 –66.51 n.d. n.d. n.d. n.d. Fischer 

(1975b)

19 Magpie 

Mountain

Canada Quebec VTM 51.38 –64.07 n.d. n.d. n.d. n.d. Fischer 

(1975b)

20 Lac St. Jean Canada Quebec VTM 49.33 –71.50 n.d. n.d. n.d. n.d. Fischer 

(1975b)
Introduction U5

Table U1. Location, grade, tonnage, and other data for selected vanadium deposits of the world.—Continued

[ID, identifier, shown in figs. U1 and U5; WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern 

Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; %, percent; V O , vanadium oxide; VTM, vanadiferous titano- 2 5

magnetite; n.d., no data available; NA, not available; t, metric ton; SSV, sandstone-hosted vanadium]

ID Deposit Country

District, 

region, 

or State/ 

Province

Deposit 

type

Latitude Longitude Resource 

tonnage 

(million 

metric 

tons)

Grade 

(% V O ) 2 5

V O 2 5

content 

(million 

metric 

tons)

Production, 

resources, 

or reserves

Source 

(decimal degrees, of data

WGS 84)

21

22

23

24

25

26

27

28

29

30

31

32

33

34

35

LaBlacheHervieuxShmoo 

Lakes

Nick

Luanping

Maocaoping, 

Baiguoyuan

Nalaqing, 

Gongshan, 

Damakan, 

Lanjian

Shanyang

Otanmaki

Mustavaara

SinghbhumMayurbhanj

Ganjang

Shaltora

Rivash

Green giant

Berg Aukas, 

Abenad, 

Tsumeb

StjernoySeiland

Canada

Canada

China

China

China

China

Finland

Finland

India

India

India

Iran

Madagascar

Namibia

Norway

Quebec

Yukon

Hubei

Province

Hubei

Province

Panzhihua 

region, 

Sichuan 

Province

Shaanxi 

Province

Kajaani, 

Kainuu 

Region

Oulu, North 

Ostrobothnia 

Region

Bihar and 

Odisha

KarbiAnglong 

District

West Bengal

Razavi 

Khorasan 

Province

Tulear 

Region

Otavi Mountainland

Finnmark 

County

VTM

Shalehosted

VTM

Shalehosted

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

Shalehosted 

(meta)

Vanadate

VTM

50.06

68.65

41.50

31.27

26.65

33.87

64.13

65.78

22.00

26.16

23.42

35.64

–24.69

–19.51

70.50

–69.63

–135.25

117.50

111.08

102.00

109.94

27.10

27.99

86.00

93.33

86.92

58.33

44.75

18.25

23.00

101.7

n.d.

n.d.

n.d.

3,460

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

49.9

3.2

n.d.

0.18

n.d.

n.d.

n.d.

0.3

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.693

0.75

n.d.

0.18

n.d.

n.d.

n.d.

10.4

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.35

0.024

n.d.

Resources

n.d.

n.d.

n.d.

Resources1

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Resources

Resources

n.d.

Fischer 

(1975b)

Hulbert and 

others (1992)

Fischer 

(1975b)

Coveney and 

Nansheng 

(1991)

Zhou and others (2005)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Energizer 

Resources, 

Inc. (2013)

Fischer (1975a); 

Boni and 

others 

(2007)

Fischer 

(1975b)
U6 Critical Mineral Resources of the United States—Vanadium

Table U1. Location, grade, tonnage, and other data for selected vanadium deposits of the world.—Continued

[ID, identifier, shown in figs. U1 and U5; WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern 

Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; %, percent; V O , vanadium oxide; VTM, vanadiferous titano- 2 5

magnetite; n.d., no data available; NA, not available; t, metric ton; SSV, sandstone-hosted vanadium]

ID Deposit Country

District, 

region, 

or State/ 

Province

Deposit 

type

Latitude Longitude Resource 

tonnage 

(million 

metric 

tons)

Grade 

(% V O ) 2 5

V O 2 5

content 

(million 

metric 

tons)

Production, 

resources, 

or reserves

Source 

(decimal degrees, of data

WGS 84)

36

37

38

39

40

41

42

43

44

45

46

47

48

49

50

51

52

Rodsand Norway

Selvag Norway

Tellnes Norway

Kusinskoe Russia

Gusevogorsk Russia

Guseva Gora Russia

Yubryah- Russia

kinskoe

Tsagin Russia

Afrikanda Russia

Lysanovskoe Russia

Visean Russia

Koykara Russia

Koyarsk

Pudozhgorsk Russia

Yelet’ ozero Russia

Pudozhga- Russia

raskoe

Pervouralsk Russia

Kharlovskoe Russia

Møre og 

Romsdal 

County

Nordland 

County

Rogaland 

County

Chelyabinskaya 

Oblast

Kachkanar, 

Sverdlovskaya 

Oblast

Kachkanar, 

Sverdlovskaya 

Oblast

Kachkanar, 

Sverdlovskaya 

Oblast

Murmanskaya Oblast

Murmanskaya Oblast

NA

NA

NA

Republic of 

Kareliya

Republic of 

Kareliya

Republic of 

Kareliya

Sverdlovskaya Oblast

Sverdlovskaya Oblast

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

VTM

62.85

68.66

58.33

55.75

58.70

58.70

59.93

68.00

67.42

54.00

53.00

61.00

61.62

67.07

61.37

56.90

52.00

8.12 n.d.

15.00 n.d.

6.42 n.d.

57.35 n.d.

59.49 16,200

59.49 n.d.

59.17 n.d.

36.00 n.d.

32.68 n.d.

94.00 n.d.

59.00 n.d.

30.00 n.d.

36.24 n.d.

34.21 n.d.

36.48 n.d.

59.53 n.d.

81.00 n.d.

n.d.

n.d.

n.d.

n.d.

0.084

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

13.6

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Resources

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Augé and others (2005); 

Laznicka 

(2010)

Augé and others (2005); 

Laznicka 

(2010)

Augé and others (2005); 

Laznicka 

(2010)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)
Introduction U7

Table U1. Location, grade, tonnage, and other data for selected vanadium deposits of the world.—Continued

[ID, identifier, shown in figs. U1 and U5; WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern 

Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; %, percent; V O , vanadium oxide; VTM, vanadiferous titano- 2 5

magnetite; n.d., no data available; NA, not available; t, metric ton; SSV, sandstone-hosted vanadium]

ID Deposit Country

District, 

region, 

or State/ 

Province

Deposit 

type

Latitude Longitude Resource 

tonnage 

(million 

metric 

tons)

Grade 

(% V O ) 2 5

V O 2 5

content 

(million 

metric 

tons)

Production, 

resources, 

or reserves

Source 

(decimal degrees, of data

WGS 84)

53

54

55

56

57

58

59

60

61

62

63

64

65

66

67

Vantra

Mapochs 

Mine

Rhovan 

Mine

Krokodilkraal 

Mine

Kafferskrall 

Farm

Taberg

Viken

Ruoutevare

Häggån

Union Bay

Klukwan 

Fan

San Gabriel 

Mountains

McClure 

Mountain

Buckmaster 

Draw

Tony M, 

Frank M

South 

Africa

South 

Africa

South 

Africa

South 

Africa

South 

Africa

Sweden

Sweden

Sweden

Sweden

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

Brits, North 

West 

Province

Bushveld 

Igneous 

Complex, 

Limpopo 

Province

Bushveld 

Igneous 

Complex, 

Limpopo 

Province

Bushveld 

Igneous 

Complex, 

North West 

Province

North West 

Province

Jönköping 

County

Myrviken, 

Jämtland 

County

Norrbotten 

County

Storsjön, 

Jämtland 

County

Alaska

Alaska

California

Colorado

Green River, 

Utah

Henry Mtns., 

Utah

VTM

VTM

VTM

VTM

Vanadate

VTM

Shalehosted

VTM

Shalehosted

VTM

VTM

VTM

VTM

SSV

SSV

–25.63

–25.22

–25.57

–25.60

–31.15

57.75

63.12

67.08

63.50

55.77

59.42

34.37

38.34

38.68

38.11

27.78

29.92

27.57

28.18

25.85

14.17

14.36

17.50

12.73

–132.10

–135.88

–118.30

–105.42

–110.03

–110.81

n.d.

100

203

n.d.

n.d.

n.d.

685

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.67

n.d.

n.d.

1.5

0.52

n.d.

n.d.

n.d.

0.29

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.19

n.d.

n.d.

1.5

1.06

n.d.

n.d.

n.d.

1.98

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

0.0012

n.d.

n.d.

Resources 

+ reserves

Resources

n.d.

n.d.

n.d.

Resources

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

n.d.

Fischer 

(1975b)

Reynolds 

(1985); 

Rohrmann 

(1985)

Reynolds 

(1985); 

Rohrmann 

(1985)

Reynolds 

(1985); 

Rohrmann 

(1985)

Fischer 

(1975a)

Fischer 

(1975b)

Aura Energy 

(2012)

Fischer 

(1975b)

Aura Energy 

(2012)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1968)

Johnson 

(1959)
U8 Critical Mineral Resources of the United States—Vanadium

Table U1. Location, grade, tonnage, and other data for selected vanadium deposits of the world.—Continued

[ID, identifier, shown in figs. U1 and U5; WGS 84, World Geodetic System of 1984; negative values for latitude indicate that the deposit is in the Southern 

Hemisphere; negative values for longitude indicate that the deposit is in the Western Hemisphere; %, percent; V O , vanadium oxide; VTM, vanadiferous titano- 2 5

magnetite; n.d., no data available; NA, not available; t, metric ton; SSV, sandstone-hosted vanadium]

ID Deposit Country

District, 

region, 

or State/ 

Province

Deposit 

type

Latitude Longitude Resource 

tonnage 

(million 

metric 

tons)

Grade 

(% V O ) 2 5

V O 2 5

content 

(million 

metric 

tons)

Production, 

resources, 

or reserves

Source 

(decimal degrees, of data

WGS 84)

68

69

70

71

72

73

74

75

76

77

78

79

80

81

Phosphoria

Mecca 

Quarry 

shale

Mike, Pandora Mine

TiTac

Gibellini

Carlin Vanadium

King Tut, 

Carrizo, 

Lukachukai Mountains

Ossining

Diana

Tahawus 

Mine, 

MacIntyre 

stone pit

Slick Rock 

Mill, 

Uravan

Iron Mountain

Broken Hill

Bulawayo

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

United 

States

Zambia

Zimbabwe

Idaho

Illinois and 

Indiana

La Sal 

district, 

Utah

Minnesota

Nevada

Nevada

New Mexico 

and 

Arizona

New York

New York

Sanford 

Lake, 

New York

Uravan Mineral Belt, 

Colorado

Wyoming

Kabwe, 

Central 

Province

Bulawayo

Shalehosted

Shalehosted

SSV

VTM

Shalehosted

Shalehosted

SSV

VTM

VTM

VTM

SSV

VTM

Vanadate

Vanadate

42.70

40.02

38.31

47.58

39.21

40.61

36.74

41.17

44.15

44.00

38.37

41.38

–14.42

–20.17

–111.83

–87.53

–109.25

–92.07

–116.09

–116.12

–109.01

–73.83

–75.25

–74.08

–108.74

–104.83

28.55

28.58

n.d.

n.d.

0.989

n.d.

0.082

25.4

0.846

n.d.

n.d.

n.d.

13.99

n.d.

n.d.

n.d.

n.d.

n.d.

1.46

n.d.

0.29

0.51

1.15

n.d.

n.d.

n.d.

1.29

n.d.

n.d.

n.d.

n.d. n.d.

n.d. n.d.

0.0144 Production

n.d. n.d.

0.00017 Resources

0.13 Resources

0.01 Production

n.d. n.d.

n.d. n.d.

n.d. n.d.

0.18 Production

n.d. n.d.

n.d. n.d.

n.d. n.d.

Love and others (2003); 

Jasinski 

(2004)

Fischer 

(1975b)

Fischer (1968); 

Shawe 

(2011)

Fischer 

(1975b)

American 

Vanadium 

Corp. (2012)

Scandium International 

Mining 

Corp. (2010)

McLemore 

and Chenoweth 

(1997)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer 

(1975b)

Fischer (1968); 

Shawe 

(2011)

Fischer 

(1975b)

Fischer 

(1975a)

Fischer 

(1975a)

1

Also includes some production and reserves.
Introduction U9

Vanadium figure 2. 

China

37%

South Africa

35%

Russia

25%

Other countries

3%

0

1,000

2,000

3,000

4,000

5,000

6,000

7,000

1970 1975 1980 1985 1990

Year

1995 2000 2005 2010

Steel Cast iron and other alloys

EXPLANATION

Chemicals and ceramics Superalloys

Consumption of vanadium concentrate, in metric tons

Vanadium figure 3.

Figure U2. Pie chart showing percentage of world 

vanadium production in 2012, by country. Compiled 

using data from Polyak (2013, table 7).

Figure U3. Graph showing major end uses of vanadium in the United States from 1970 to 2011. The 

layers of the graph are placed one above the other, forming a cumulative total. Compiled using data from 

U.S. Bureau of Mines (1972–76, 1992–96) and U.S. Geological Survey (1997–2013, 2014).

Demand and Availability of Supply

World vanadium resources in 2012 were estimated 

to total more than 63 million metric tons of vanadium. 

Reserves were estimated to be 14 million metric tons of 

vanadium. Because vanadium is usually recovered as a 

byproduct or coproduct, the identified world resources of 

this mineral commodity are understated and therefore not 

fully indicative of available supply. Although domestic 

resources and secondary recovery are adequate to supply 

a large portion of domestic needs, a substantial part of 

U.S. demand is currently met by foreign material (Polyak, 

2013). Future demand is expected to increase primarily 

because such countries as China and Japan are increasing 

the amount of vanadium used in steelmaking to match 

the quality of steel produced from other countries, as well 

as in anticipation of the probable increase in alternative 

renewable sources of energy that require the use of VRBs 

(Roskill Information Services, Ltd., 2013).
U10 Critical Mineral Resources of the United States—Vanadium

Geology

Geochemistry

Vanadium is a trace element that is widely distributed 

in nature. The average abundance of vanadium in the upper 

continental crust is approximately 60 parts per million (ppm) 

(Taylor and McLennan, 1995). Vanadium is produced from 

the combustion of fossil fuels, the mining of ores, and hightemperature industrial activities, including steel refining and 

the processing of phosphate ores.

The natural oxidation states of vanadium are +2, +3, 

+4, and +5. The trivalent ion V3+ has an octahedral radius 

(0.061 nanometers [nm]) that is almost identical with that 

of ferric iron (Fe3+, 0.063 nm). As a consequence, vanadium 

substitutes in iron-rich minerals, such as amphibole, biotite, 

magnetite, and pyroxene, and for aluminum (Al3+, 0.054 nm) 

in iron and ferromagnesian minerals (Fischer, 1973). Vanadium is present in large amounts in mafic igneous rocks (about 

250 ppm), is less abundant in ultramafic and intermediatecomposition rocks (about 50 ppm), and is present in only 

small amounts in felsic rocks (about 20 ppm). Vanadium is 

concentrated in magmatic magnetite deposits, especially those 

that are titaniferous; the vanadium concentrations in these 

deposits commonly range from 1,000 to 5,000 ppm.

Vanadium is also closely associated with organic-rich 

sediments. The average concentrations in shale are about 

130 to 205 ppm, but some carbonaceous shales may have 

up to 5,000 ppm vanadium (for example, shales in eastern 

Kentucky; Robl and others, 1983). The addition of vanadium 

to sediments rich in organic carbon is owing to reduction, 

adsorption, and complexation. Dissolved vanadate species (5+

) 

in oxic seawater are reduced to vanadyl ions (4+) by organic 

compounds or hydrogen sulfide (H2S) and are readily adsorbed 

to particle surfaces as they settle during sedimentation. 

Vanadium3+ may also substitute for aluminum in the octahedral sites of clays (Breit and Wanty, 1991).

Because most vanadium in primary minerals is in 

the weakly soluble +3 valence state, very little vanadium 

is transported by hydrothermal fluids; therefore, most 

hydrothermal ore deposits contain low concentrations 

(10 to 100 ppm) of vanadium. Some hydrothermal deposits 

related to alkaline igneous rocks contain high concentrations 

of vanadium, however. Altered titanium-rich syenite intrusions 

at Potash Sulphur Springs in Garland County, Arkansas, 

contain high concentrations of vanadium (McCormick, 1978) 

as do some gold-quartz veins associated with alkaline igneous 

rocks, especially those with gold-telluride minerals (for 

example, Spry and Scherbarth, 2006).

During weathering in humid climates, much of the 

vanadium contained in ferromagnesian minerals apparently 

partitions into newly formed clay minerals, but it remains in 

the +3 valence state or oxidizes to the +4 valence state, both 

of which are relatively insoluble. If erosion is insignificant 

but chemical leaching is intense, the residual material may 

be enriched in vanadium, as are some bauxites and laterites 

(Patterson, Kurtz, and others, 1986). During the weathering 

of igneous, residual, or sedimentary rocks, some vanadium 

oxidizes to the +5 valence state (vanadate), especially in the 

intensive oxidizing conditions that are characteristic of arid 

climates. In this oxidized valence state and under relatively 

alkaline conditions, vanadium is enriched in surface waters or 

groundwaters and remains in solution (Wright and Belitz, 2010). 

It can be precipitated from solution and locally concentrated 

in rocks by the following processes (Fischer, 1973, p. 682):

(a) Coprecipitation and adsorption with hydroxides of 

aluminum or ferric iron. This process forms or enriches the 

vanadium concentration in some bauxites and in residual or 

sedimentary iron ores; 

(b) Reaction with cations of heavy metals, such as

copper, lead, uranium, and zinc. This process forms epigenetic 

vanadate minerals in the oxidized zones of base-metal 

deposits; and 

(c) Reduction in the presence of organic material or 

biologically generated H2S. Interactions of vanadium-bearing 

solutions with wall rocks during fluid flow may produce 

epigenetic ore deposits, such as those occurring in sandstones 

of the Colorado Plateau province in the United States. On the 

other hand, if the vanadium is carried to the oceans by surface 

waters, it may concentrate syngenetically in carbonaceous 

phosphorites, marls, and shales. Furthermore, if the organic 

materials in these shales are converted to liquid hydrocarbons, 

the vanadium may be moved with the petroleum in vanadiumorganic compounds and in fine inorganic particulates; some 

crude oils contain as much as several hundred parts per 

million vanadium. Vanadium accumulates in the ashes of these 

oils and also in the residues resulting from their natural or 

industrial distillation. These ashes and residues have been used 

as commercial sources of vanadium (Fischer, 1973, p. 682; 

Breit, 1992).

Mineralogy

Vanadium occurs in nature in a wide variety of minerals 

(table U2). The following four principal types of mineral 

deposits are recognized: vanadiferous titanomagnetite (VTM) 

deposits, sandstone-hosted vanadium (SSV) deposits, shalehosted deposits, and vanadate deposits. Magnetite (Fe3

O4) 

and ilmenite (FeTiO3) are the principal vanadium-bearing ore 

minerals in VTM deposits (fig. U4A), but hematite (Fe2O3), 

perovskite (CaTiO3), and rutile (TiO2) are present in some 

deposits (Fischer, 1975b). These minerals occur in medium- to 

fine-grained intergrowths and in exsolution and solid-solution 

relations; ulvospinel and titanomagnetite are two mineral 

names commonly applied to some of these exsolution and 

solid-solution forms. Small blebs and exsolution blades of 

coulsonite (FeV2O4 ) in magnetite have been recognized in a 

few VTM deposits (Balsley, 1943).

Ore minerals below the zone of oxidation in SSV deposits 

are low-valence oxides and silicates of uranium and vanadium. 

Coffinite, montroseite, paramontroseite, uraninite, and 
Geology U11

Mineral name Chemical formula or description

Vanadiferous titanomagnetite (VTM) deposits

Coulsonite (Fe,V)3

O4

Hematite Fe2O3

Ilmenite1 FeTiO3

Magnetite1 Fe3O4

Perovskite CaTiO3

Rutile TiO2

Sandstone-hosted vanadium (SSV) deposits

Carnotite K2

(UO2 )2(VO8 )• 1–3H2O

Coffinite U(SiO4)1–x (OH)4x

Corvusite (Na,Ca,K)V8O2 • 4H2O

Doloresite H8V6O16

Hewettite CaV6

5+O16 •9H2

O

Montroseite (V3+,Fe3+,V4+)O(OH)

Paramontroseite VO2

Pascoite Ca3(V10O28 )• 17H2O

Roscoelite K(V3+,Al,Mg)[(Si,Al)4O10 ](OH)2

Tyuyamunite Ca(UO2)2V2O8 • 5–8H2O

Uraninite UO2

 (with Pb, Th, V, Zr)

Vanadium clays Vanadium-bearing hydrous mica

Vanadian chlorite Vanadium-bearing chlorite

Volborthite Cu3

V2O7(OH)2 • 2H2O

Shale-hosted vanadium deposits

Illite-smectite3 K0.8(Al2.8Mg0.5Fe0.4V0.3)(Si7.2Al0.8)O20(OH)4

Metahewettite CaV6O16 H2O

Quisqueite2 Vanadium-bearing organic matter

Patronite2 V4+(S2

2–)2

Table U2. Selected vanadium-bearing minerals, by deposit type.

Mineral name Chemical formula or description

Vanadate deposits

Brachebuschite Pb2

(Mn,Fe)(VO4)2 •H2O

Calciovolborthite CaCuVO4

(OH)

Chervetite Pb2V2O7

Curienite Pb(UO2)2(VO4 )2 • 5H2O

Descloizite1 PbZn(VO4 )(OH)

Francevillite (Ba,Pb)(UO2)2

(VO4 )2 •5H2O

Heyite Pb5

Fe2(VO4 )2O4

Mottramite1 PbCu(VO4)(OH)

Mounanaite PbFe2(VO4 )2(OH)

Pyrobelonite PbMnVO4

(OH)

Sengierite Cu(UO2)

2

(VO4 )2 •8 or 10H2O

Turanite Cu5

(VO4 )

2

(OH)4

Vanadinite1 Pb5(VO4 )

3

Cl

Volborthite Cu3

(VO4 )

2 • 3H2

O

Epithermal/porphyry/magmatic deposits

Duttonite V4+O(OH)2

Fervanite Fe4

3+(VO4)4 • 5H2

O

Hewettite CaV6

5+O16 • 9H2

O

Karelianite4 V2

O3

Metatyuyamunite6 Ca(UO2)2VO4)2 •5–7H2O

Montroseite (V3+,Fe3+,V4+)O(OH)

Nolanite4 (V,Fe,Al)10O14(OH)2

Roscoelite K(V3+,Al,Mg)[(Si,Al)4O10 ](OH)2

Rutile, brookite TiO2

Schreyerite4 VSiO3(OH)

Vanadium andradite5 Ca3

(Fe3+,V3+)2(SiO4 )3

Vanadium muscovite4 Vanadium-bearing muscovite

Vanadium silicates4 Si3O9V2

Vanadium- and 

titanium-bearing 

mixed layer clay7

(Ca0.08K0.02Na0.03 )0.13

(Al0.11V0.90Ti0.19Fe0.64Mg0.14 )1.98

(Si3.79Al0.21 )4O10(OH)2 • nH2O

1Most common.

2Mina Ragra, Peru (Fischer, 1973).

3Mecca Quarry Shale, Illinois and Indiana (Peacor and others, 2000).

4Tuvatu deposit, Fiji (Spry and Scherbarth, 2006).

5Potash Sulphur Springs or Wilson Springs, Arkansas (Howard and 

Owens, 1995).

6

Weeks and others, 1959.

7Potash Sulphur Springs or Wilson Springs, Arkansas (McCormick, 1978).
U12 Critical Mineral Resources of the United States—Vanadium

Vanbadium Figure 4.

A B

C D

Figure U4. Photographs showing four examples of vanadium. A, Magnetite layer (dark) from the Bushveld Complex, South Africa. 

It is these layers that commonly host vanadium. B, Natural vanadinite, which is a main source of vanadium from vanadate deposits. 

C, Vanadium metal crystals made by electrolysis (the largest crystal is 2 centimeters in length). D, Common hand wrench made 

with vanadium alloy steel. Photograph A courtesy of Kevin Walsh/CC BY 2.0 (https://creativecommons.org/licenses/by/2.0/). 

Photographs B and C courtesy of Juergen Kummer, Jumk.de Web Projects/CC BY 3.0 (http://images-of-elements.com/vanadium.php). 

Photograph D courtesy of MrX/CC-BY-SA-3.0 (https://commons.wikimedia.org/wiki/File:Chrome_Vanadium_Adjustable_Wrench.jpg).

vanadium-rich clay or mica (roscoelite), vanadian chlorite, and 

vanadium-rich clay minerals are the most common (table U2). 

Chalcopyrite, clausthalite, ferroselite, galena, naumannite, 

pyrite, sphalerite, and other sulfides and selenides are common 

trace minerals (Shawe, 2011). Copper sulfides are widespread 

in sparse amounts but are abundant enough in a few deposits 

to constitute ore minerals (Fischer, 1968). Partially oxidized 

orebodies (for example, those immediately above the water 

table) contain minerals similar to those of deposits below the 

zone of oxidation, but with the addition of corvusite-group 

minerals and hewettite (Shawe, 2011). Oxidized deposits 

contain carnotite and tyuyamunite, together with vanadium 

clays and chlorite. The most common authigenic gangue 

minerals in SSV deposits are barite, carbonates (mostly 

calcite), and quartz (Breit and Goldhaber, 1996). Ore minerals 

are typically disseminated in the sandstone or form radiating 

acicular aggregates that fill open pore spaces (Wanty and 

others, 1990). The ore minerals are typically concentrated 

near carbonaceous material (Fischer, 1968; Shawe, 2011).

Vanadate deposits contain a wide variety of vanadiumbearing minerals (table U2). Vanadate minerals are formally 

described as containing an oxoanion of vanadium generally in 

its highest oxidation state of +5. Most vanadate minerals are 

compounds of copper, iron, or lead because these deposits form 

within the oxidized zones of base-metal deposits in areas of 

arid climate and deep oxidation (Fischer, 1975a). Descloizite, 

mottramite, and vanadinite (fig. U4B) are the most common 

vanadium minerals, although numerous others are known 

(table U2). Wulfenite (PbMoO4), which is a molybdate of lead, 

is also common in vanadate deposits (Fischer, 1975a). Where 

paragenetic relations are reported, the vanadium-bearing 

minerals are late in the sequence. For example, the association 
Geology U13

of vanadinite and descloizite is paragenetically late in the 

Berg Aukas and Abenab deposits in the Otavi Mountainland, 

Namibia (Boni and others, 2007). Fischer (1975a) suggests 

that vanadate minerals typically coat and partly replace 

supergene base-metal minerals in deposits in Arizona, but in 

most cases, the vanadates overlap or precede the last stage of 

supergene mineral formation (Boni and others, 2007). Once 

formed, vanadate minerals are stable in the environment of 

the oxidized zone; these minerals persist from the surface to 

the bottom of the oxidized zone, and leaching and corrosion 

of them are rarely reported (Fischer, 1975a). In the Kabwe 

district (formerly known as Broken Hill), Zambia, the vanadate minerals are persistent to depths of 350 meters (m) below 

the surface, which defines the lowest depth of oxidation.

Some hydrothermal deposits contain vanadium-bearing 

minerals, particularly deposits genetically and (or) spatially 

associated with alkaline igneous rocks. In epithermal 

gold deposits, roscoelite is the most common mineral, but 

vanadium-bearing rutile or brookite, karelianite, nolanite, 

shreyerite, and vanadium-rich muscovite and silicate minerals 

have also been identified (Spry and Scherbarth, 2006).

Deposit Types

Vanadium is present in economic concentrations in four 

main types of mineral deposits and as a minor constituent in 

several other types. Fossil fuels are another important source 

of vanadium.

Vanadiferous Titanomagnetite Deposits

Vanadiferous titanomagnetite (VTM) deposits (Fischer, 

1975b) are found throughout the world and are the principal 

source of vanadium (fig. U1). The most economically 

significant VTM deposits or regions, both in the past and 

currently, include the Bushveld Complex in South Africa 

(Reynolds, 1985); the Panzhihua layered intrusion in Sichuan 

Province, China (Zhou and others, 2005); the Kachkanar massif 

in the Ural Mountains in Russia; the Windimurra Complex 

in Western Australia (Ivanic and others, 2010); and the 

Bell River Complex (Matagami deposit) and the Lac Doré 

Complex in Quebec, Canada (fig. U1; Taner and others, 1998).

The VTM deposits consist of magmatic accumulations 

of magnetite and ilmenite, defined arbitrarily as having 

grades of more than about 1 percent rutile (Fischer, 1975b). 

They commonly contain 0.2 to 1 percent V2O5, but some 

zones (for example, the Bushveld Complex) contain greater 

than 1.5 percent V2O5 (Reynolds, 1985). Most exposed 

VTM deposits are Archean or Proterozoic in age, having 

formed in intraplate continental tectonic settings; a few 

deposits (for example, the Panzhihua region in Sichuan Province, China) are younger (table U1; fig. U1). Most deposits 

are associated with large igneous province (LIP) magmatism; 

some layered intrusive complexes are linked to mantle plumes 

and hotspot tracks (Hatton, 1995; Ernst and others, 2005).

VTM deposits are hosted mainly within mafic and 

ultramafic igneous rocks, most commonly anorthosite and 

gabbro. Lithologies within the igneous complexes that contain 

the vanadium-rich ores vary considerably, however. For 

example, in the Bushveld Complex, lithologies range from 

dunite and pyroxenite to anorthosite and pure oxide layers 

(Eales and Cawthorn, 1996). Some vanadiferous deposits 

are hosted in zoned mafic to ultramafic complexes with high 

levels of chromium and platinum-group elements (PGEs); 

these complexes are sometimes referred to as Alaska-type 

PGE deposits; and examples include the Union Bay deposit 

in Alaska (United States), and the Kachkanar complex in 

Sverdlovsk Oblast, Russia. A few deposits are associated 

with alkalic igneous rocks (for example, syenodiorite is 

genetically related to layered gabbro in the Panzhihua district 

of Sichuan Province, China [Shellnut and Jahn, 2010] and in 

the Ganjang deposit in Assam, northeastern India [Saha and 

others, 2010]). The mafic to ultramafic igneous host rocks are 

typically of deep-seated origin, and they occur in stratiform 

tabular bodies, such as sills and laccoliths that are thick and 

laterally extensive. For example, the Mesoarchean gabbroic 

Windimurra Complex in the Yilgarn craton (Western Australia, 

Australia) covers an area of about 2,500 square kilometers 

(km2) and contains layers of gabbroic and ultramafic 

rocks with cumulative thicknesses of 13 kilometers (km) 

(Ivanic and others, 2010). The Paleoproterozoic Bushveld 

Complex (dated at 2.06 giga-annum [Ga]) in the Transvaal 

basin in South Africa contains layers with cumulative 

thicknesses (about 9 km) that are comparable to those of the 

Windimurra Complex, and it has an areal extent of about 

65,000 km2 (Eales and Cawthorn, 1996).

In contrast to laterally extensive and thick tabular bodies, 

some titaniferous magnetite deposits are hosted in relatively 

complex intrusive or lens-shaped bodies (Fischer, 1975b). 

Such deposits are variable in size and shape. In the Panzhihua 

region of China, the maximum size of the lens-shaped ore 

bodies is 160 m long and 30 m wide (Zhou and others, 2005). 

Vanadium-bearing titanomagnetite previously mined in the 

Sanford Lake District of New York and the Taberg deposit in 

Jönköping County, Sweden, was contained in small (550 m 

by 180 m) bodies of gabbro and anorthosite (Balsley, 1943; 

Gross, 1968; Fischer, 1973, 1975b).

The textures and mineralogy of VTM ores are remarkably similar among the largest known deposits, including the 

Panzhihua deposit and the Bushveld and Windimurra deposits 

(Fischer, 1975b; Reynolds, 1985; Rohrmann, 1985; Zhou and 

others, 2005; Ivanic and others, 2010). Ores typically form 

discrete layers that are concordant with the igneous layering, 

which varies between 0.1 and 10 m in thickness, although 

some oxide layers in deposits in the Panzhihua region of 

southwestern China attain thicknesses of 60 m (Zhou and 

others, 2005). Similar to the silicate rocks that host them, the 

oxide layers are laterally extensive, and they can be traced for 

hundreds of kilometers in the case of the Bushveld Complex 

(Rohrmann, 1985; Reynolds, 1985). Most oxide layers have 

sharp lower boundaries that host silicate rocks and have 
U14 Critical Mineral Resources of the United States—Vanadium

gradational tops; some oxide layers contain thin intercalations 

of gabbro (Eales and Cawthorn, 1996; Zhou and others, 2005). 

Most of the vanadium in titaniferous magnetite deposits is 

concentrated as a solid solution in magnetite-ulvospinel, where 

V3+ has replaced Fe3+ (Fischer, 1975b); the vanadium-rich 

spinel mineral coulsonite (FeV2O4) was reported as small 

blebs and exsolution blades in magnetite in a few deposits 

(Balsley, 1943; Fischer, 1975b). Ilmenite, hematite, rutile, and 

perovskite commonly accompany magnetite (Fischer, 1975b). 

Where exposed to weathering, the magnetite may oxidize 

to vanadomaghemite ((FeTi)2O3) and small concentrations 

of hematite without any change in the texture of the ore 

(Rohrmann, 1985).

Ores may be either massive or disseminated. Massive 

ores typically consist of closely packed, nearly equant grains 

of more than 80 percent titanomagnetite and contain variable amounts of clinopyroxene, olivine, and plagioclase. If 

silicate minerals are present, they are completely surrounded 

by oxides (Zhou and others, 2005). Disseminated ores are 

generally coarse grained and are composed of about 50 percent 

titanomagnetite, about 20 percent clinopyroxene, about 

20 percent plagioclase, about 10 percent ilmenite, and small 

amounts of olivine (Rohrmann, 1985; Eales and Cawthorn, 

1996; Zhou and others, 2005).

The mechanisms by which millions of tons of vanadium 

become concentrated in massive iron-titanium oxide deposits 

remain poorly understood, but most researchers agree that 

partial melting of mantle rocks and extensive fractionation 

of the derivative magma are critical early-stage processes. 

Large-scale, in situ crystallization of plagioclase and other 

anhydrous phases (olivine, pyroxene) in the basal parts of 

magma chambers results in an increase in total iron and 

water contents of the residual magma with eventual formation of immiscible oxide melts (Reynolds, 1985; Eales and 

Cawthorn, 1996; Ivanic and others, 2010; Shellnut and Jahn, 

2010). Such oxide ore melts are denser than silicate melts and 

therefore settle to the bottom of the magma chamber, where 

the oxide melts accumulate (Eales and Cawthorn, 1996; Zhou 

and others, 2005). Many layered intrusions show evidence 

of multiple magma injections (Eales and Cawthorn, 1996), 

suggesting that magma mixing may have played a role in 

the development of some titaniferous magnetite deposits 

(Von Gruenwaldt, 1993). Other factors involved in forming 

immiscible oxide melts from silicate magmas are abrupt 

changes in oxygen fugacity, and (or) an introduction of fluids 

(Reynolds, 1985; Zhou and others, 2005). The presence of 

minor disseminated sulfides and apatite (for example, the 

Panzhihua deposits) suggests that sulfur and phosphorus may 

have acted as fluxing agents that promoted the concentration 

of immiscible liquids that led to formation of the Panzhihua 

deposits (Zhou and others, 2005).

Associated metals.—The igneous rocks that host significant 

vanadium resources are commonly temporally and spatially 

associated with magmatic chromium, copper, nickel, and 

PGE deposits (Cawthorn and others, 2002; Naldrett, 2010; 

Zientek, 2012). At least minor amounts of copper-nickel-PGE 

enrichment occurs in igneous complexes adjacent to the vanadiferous zones in the Windimurra region of Australia (Ivanic and 

others, 2010), the Panzhihua region of China (Zhou and others, 

2005), and Kachkanar in the Urals region of Russia (Augé and 

others, 2005). The largest resources of PGEs and chromium 

in the world are contained within the Bushveld Complex of 

South Africa in layers that lie below the vanadiferous zones 

(Eales and Cawthorn, 1996; Naldrett and others, 2009). 

Additionally, high contents of scandium (up to 500 grams per 

metric ton of scandium oxide [Sc2O3]) occur in some of the 

VTM deposits in Russia (Bykhovsky and Tigunov, 2008).

Sandstone-Hosted Vanadium Deposits

Sandstone-hosted uranium deposits have been identified 

on all continents, and many are known to have enrichments 

of vanadium (Dahlkamp, 2010). These deposits of vanadiumand uranium-bearing sandstone (known as sandstone-hosted 

vanadium [SSV] deposits) have average resource and ore 

grades that range from 0.1 to 1 weight percent vanadium 

(George Breit, U.S. Geological Survey [retired], written 

commun., 2013). On a global scale, the United States has 

been and is currently the main producer of vanadium from 

SSV deposits, particularly from those on the Colorado Plateau. 

Additionally, these SSV deposits are the chief domestic source 

of vanadium in the United States (Fischer, 1968, 1973; Polyak, 

2012, 2013). Most deposits are located in western Colorado 

and eastern Utah, although some are also located in northern 

Arizona and New Mexico (fig. U1). Vanadium enrichments 

in sandstones elsewhere in the world include the Bigrlyi 

deposit in Northern Territory, Australia, which has a grade of 

0.13 weight percent vanadium (McKay and Miezitis, 2001); 

deposits in the Tonco-Amblyo district in Argentina, which 

have grades of 0.1 to 0.3 weight percent vanadium (Dahlkamp, 

2010); and a deposit in the Karamurun district of Almaty 

Oblysy, Kazakhstan (Dahlkamp, 2009).

The Colorado Plateau province covers an area of 

337,000 km2 and was developed through a series of geologic 

and tectonic events. Important processes in this region that 

promoted the formation of SSV deposits include deposition 

of evaporitic-sapropelic strata from which brine fluids were 

derived; deposition of fluvial, fine-grained, locally carbonaceous sandstone that served as host rocks for the deposits; and 

magma generation and emplacement of laccoliths during the 

Tertiary that may have heated and driven brine fluids upwards 

along favorable structures (Shawe, 2011).

The SSV deposits occur in fluvial sandstone lenses of the 

Chinle Formation (Upper Triassic) and the Jurassic Morrison 

and aeolian Entrada Formations. The most economically 

significant vanadium deposits occur within the Salt Wash 

Member, which is the basal unit of the Morrison Formation 

(Northrop and others, 1990; McLemore and Chenoweth, 

1997; Shawe, 2011). The Salt Wash Member is 30 to 150 m 

thick and contains laterally continuous fluvial sandstone and 

interbedded mudstone. This member was deposited by a 

west-to-east aggrading alluvial system (Shawe, 2011).
Geology U15

The SSV deposits form subhorizontal lenses or tabular 

bodies that are variable in thickness but are typically less than 

a few meters thick (Northrop and others, 1990; McLemore 

and Chenoweth, 1997) and up to 150 m long (Shawe, 2011). 

Some elongate mineralized zones are parallel to paleostream 

channels in the fluvial sandstone (McLemore and Chenoweth, 

1997; Shawe, 2011). Deposits are typically concordant with 

bedding, although discordant lenses of uranium-vanadium 

minerals cut bedding planes locally. Most commonly, the 

orebodies are entirely within sandstone, although some occur 

at the interface between sandstone and less permeable shale 

or siltstone (Northrop and others, 1990; McLemore and 

Chenoweth, 1997). Tabular bodies display sharp to gradational 

transitions into unmineralized sandstone, and generally 

terminate abruptly against mudstone or claystone seams 

(Shawe, 2011). The SSV deposits are typically found in areas 

with local accumulations of detrital carbonaceous material. 

Pyrite is common within all forms of carbonized plant material 

(Shawe, 2011). Northrop and others (1990) reported that pyrite 

from mineralized intervals has distinctly lower sulfur isotope 

compositions (–4.8 to –2 per mil) than pyrite in unmineralized 

rocks (which average 12.1 per mil). Some tabular deposits 

have been reworked over time by groundwater, resulting in the 

formation of redistributed orebodies (Shawe, 2011).

Vanadium concentrations in SSV ores (expressed as 

V2O5) are commonly 1 percent or greater, and some deposits in 

southwestern Colorado have grades of more than 2.5 percent 

(Shawe, 2011). Vanadium-to-uranium weight ratios in deposits 

within the Morrison Formation vary from 2:1 to 6:1 among 

most areas of production, and therefore, are considered vanadium deposits with accessory uranium (Northrop and others, 

1990; Shawe, 2011). Deposits with high vanadium-to-uranium 

weight ratios are located in southeastern Utah and southwestern Colorado; those in east-central Utah and west-central 

Colorado are characterized by lower vanadium-to-uranium 

weight ratios (George Breit, U.S. Geological Survey [retired], 

written commun., 2013). Many deposits have multiple ore 

zones, consisting of two (or more) closely spaced intervals 

enriched in vanadium and uranium, separated by an interval 

(or intervals) containing no uranium but enriched in vanadium 

(Northrop and others, 1990). The uranium-vanadium minerals 

form the matrix of the mineralized sandstones and locally 

replace detrital quartz and feldspar grains (Kovschak and 

Nylund, 1981; McLemore and Chenoweth, 1997). Vanadium, 

as +2, +3, +4, or +5, occurs as oxide phases or is combined 

with other elements, forming more than 40 different minerals 

in SSV ores (Weeks and others, 1959). Primary SSV ores are 

characterized by a consistent black-mineral suite composed 

of coffinite, montroseite, uraninite, and vanadium aluminosilicates (that is, low-valence V3+ minerals). These primary 

minerals are modified by progressive secondary oxidation 

above the water table to form an oxidized mineral assemblage 

dominated by carnotite, corvusite, and tyuyamunite (Weeks 

and others, 1959). Tabular vanadium-uranium deposits in 

the Salt Wash Member are unusual in that one of the major 

ore phases is typically roscoelite (a vanadium-bearing mica) 

accompanied by chlorite, mixed layer “chloritemontmorillonite” (chlorite-smectite), mixed-layer illite-montmorillonite, 

and “vanadium-hydromicas” (vanadium-bearing illite) 

(Northrop and others, 1990; Wanty and others, 1990). Gangue 

minerals that have been introduced or redistributed during 

formation of the deposits are anatase, barite, carbonates 

(mostly calcite and dolomite), and quartz (mostly calcite and 

dolomite) containing appreciable amounts of iron and (or) 

magnesium (Breit and Goldhaber, 1996).

Most proposed models for the formation of SSV deposits 

suggest that deposition occurred at an interface between two 

fluids of different chemical compositions and (or) states of 

oxidation-reduction (Fischer, 1968; Northrop and others, 1990; 

Shawe, 2011). Deposition involving two fluids was proposed 

many years ago during the early stages of exploration and 

production, and subsequent models, such as the brine-interface 

model, have refined or incorporated portions of these early 

theories. Although details differ, most models suggest that 

these deposits form by multistage processes: (1) shallow burial 

and diagenesis (<400 m), and reduction of dissolved uranium 

and vanadium by organic matter and bacteriogenic hydrogen 

sulfide that create local ore-grade accumulations; (2) progressive burial of the sandstone, creating changes in pore-water 

composition, pressure, and temperature and recrystallization 

of ore phases; and (3) migration of brines that move upward 

along faults and outward into permeable sandstone, mix with 

dilute meteoric water, and result in alteration of existing 

uranium and vanadium deposits (Breit and Goldhaber, 1996). 

The source of the vanadium is not well constrained, but is 

likely from the dissolution of iron-titanium-oxide minerals 

(ilmenite and magnetite) in volcanic detritus and sandstones 

within the Morrison Formation, or from unspecified source 

rocks that could be either younger or older than the host 

formation (Thamm and others, 1981; Northrop and others, 

1990). Critical to the development of brines and the location of 

the SSV deposits in the Salt Wash Member was the deposition 

of evaporitic-sapropelic sedimentary units of Middle and 

Late Pennsylvanian age beneath the Morrison Formation 

(Shawe, 2011). Determinations of the age of formation of the 

deposits varies greatly; some estimates are that primary ores 

formed close to the time of deposition of the host rock (about 

130 mega-annum [Ma]), whereas others suggest a much later 

time (about 30 Ma or latest Oligocene) (Shawe, 2011).

Associated metals.—Uranium occurs with vanadium in 

all SSV deposits, although the uranium-to-vanadium ratio 

varies greatly. Most deposits in the southwestern part of the 

United States (New Mexico and Arizona) have high uraniumto-vanadium ratios that decrease to the north (because of the 

increasing contents of vanadium). After initial mining for 

uranium, these deposits were sought as a source of radium. In 

the 1930s and 1940s, the deposits were of interest mainly for 

vanadium. Most mines were closed for nearly a decade, but 

mining resumed again in 1949 (until about 1983) for major 

production of uranium, with vanadium as a byproduct (Shawe, 

2011). Copper accompanies uranium or vanadium in many of 

the deposits (Shawe, 2011).
U16 Critical Mineral Resources of the United States—Vanadium

Shale-Hosted Vanadium Deposits

Vanadium-rich metalliferous black shales occur primarily 

in late Proterozoic and Phanerozoic marine successions. The 

term shale is used here broadly to include a range of carbonaceous rocks that include marls and mudstones. These finegrained sedimentary rocks were deposited in epeiric (inland) 

seas and on continental margins. They typically contain high 

concentrations of organic matter (greater than 5 percent) and 

reduced sulfur (greater than 1 percent; mainly as pyrite), as 

well as a suite of metals, such as copper, molybdenum, nickel, 

PGEs, silver, uranium, vanadium, and zinc (Desborough and 

others, 1979; Coveney and Martin, 1983; Coveney and others, 

1992; Hatch and Leventhal, 1992; Piper, 1999). Concentrations regularly exceed 0.18 percent V2

O5 and can be as high 

as 1.7 percent V2O5. Vanadiferous black shales are commonly 

found with phosphorite deposits and marine oil shales and, 

in North America, are marine equivalents of coal-bearing 

cyclothems. Well-characterized vanadiferous black shales 

include the Woodruff Formation in Nevada (Desborough and 

others, 1979), the Meade Peak Phosphatic Shale Member of 

the Phosphoria Formation in Idaho and Wyoming (McKelvey 

and others, 1986; Love and others, 2003), the Mecca Quarry 

Shale Member of Illinois and Indiana (Coveney and others, 

1987), the Doushantuo Formation in Hubei Province in 

southern China (Fan and others, 1992), and portions of the 

Toolebuc Formation in Queensland, Australia (Lewis and 

others, 2010). Although these black shales have long been 

recognized as potential sources of vanadium, they are not 

currently exploited. Project development is underway at the 

Gibellini vanadium prospect in Nevada (Woodruff Formation), 

and if production begins, it will be the first primary shalehosted producer of vanadium in the United States. The Julia 

Creek deposit (Toolebuc Formation) is also in the planning 

stages. The Green Giant deposit in southern Madagascar 

(Energizer Resources, Inc., 2013) consists of metamorphosed 

vanadiferous shale that extends for at least 21 km along strike 

and is reported to contain about 350,000 metric tons of V2

O5

(table U1; fig. U1).

The ultimate source of vanadium in metalliferous black 

shales is dissolved vanadium in seawater (Breit and Wanty, 

1991; Piper, 1994). Whereas the specific mechanisms of 

enrichment are disputed, all require the reduction of dissolved 

V5+ (Breit and Wanty, 1991), which is the predominant redox 

state in the oceans (Collier, 1984). Vanadium is used by 

various phytoplankton species (Robson and others, 1986; 

Moore and others, 1996), and sedimentation of phytoplankton 

debris likely acts as a minor source of vanadium in black 

shales (Piper, 1994). In oxygen-deficient bottom waters and 

pore waters, dissolved vanadium is reduced to particle-reactive 

V4+ and is incorporated into the sedimentary fraction (Wehrli 

and Stumm, 1989). Further reduction to V3+ requires the 

presence of dissolved aqueous sulfide (H2S) and promotes the 

incorporation of vanadium into sedimentary organic matter 

and authigenic clays (Lewan and Maynard, 1982; Lewan, 

1984; Breit and Wanty, 1991).

Vanadium concentrations correlate with organic carbon 

in black shales, suggesting that vanadium is incorporated 

into organic matter upon burial (Breit and Wanty, 1991). 

Black shales that have been buried to depths sufficient to pass 

through the oil window typically produce petroleum that has 

high vanadium concentrations (Lewan and Maynard, 1982; 

Lewan, 1984). Conversely, vanadium can become incorporated 

into illite upon burial (Peacor and others, 2000). Because no 

modern analogues for vanadiferous black shales are known, the 

processes of vanadium enrichment are not well understood.

Vanadium-rich black shales of North America occur 

in Devonian to Permian marine successions. Black shale of 

the Upper Devonian Woodruff Formation (Nevada) contains 

0.5 to 1.2 percent V2O5 in unaltered rocks containing high 

concentrations of organic matter (greater than 10 percent) 

(Desborough and others, 1979, 1981). Oxidized zones of 

the Woodruff Formation contain 1.1 to 1.4 percent V2O5, 

reflecting secondary enrichment of vanadium during 

oxidation; the principal vanadium mineral in the oxidized 

shales is metahewettite (CaV6O162 • H2O).

Vanadiferous black shales are also found with Pennsylvanian cyclothems of North America (Coveney and Martin, 

1983; Coveney and others, 1987; Coveney and Glascock, 

1989; Hatch and Leventhal, 1992). These typically thin, 

metalliferous black shales occur throughout the midcontinent 

region. The Mecca Quarry Shale of Illinois and Indiana is 

conspicuously rich in various metals, including vanadium, 

with concentrations of up to 10,000 ppm (Coveney and 

Martin, 1983). These concentrations of vanadium exceed that 

of many VTM deposits; however, because the Mecca Quarry 

Shale is only a few tens of centimeters thick, it is not considered an economically viable vanadium resource (Coveney and 

Martin, 1983). Vanadium enrichments in the Mecca Quarry 

Shale could be related to the action of basinal brines, similar 

to processes that concentrate lead-zinc in Mississippi Valleytype (MVT) deposits (Coveney and Glascock, 1989), although 

direct evidence for this mineralizing process is lacking.

The Permian Phosphoria Formation in Idaho and Wyoming 

contains a world-class phosphate deposit that also includes 

vanadium-rich strata (McKelvey and others, 1986; Piper, 1999). 

The black-shale interval within the Meade Peak Phosphatic 

Shale Member contains an average of 1.2 percent V2

O5

(Piper, 1999) and, since the early 1940s, has been considered 

a potential economic source of vanadium (Love and others, 

2003). In the 1960s, vanadium and uranium were produced 

from ferrophosphorus, a byproduct of an elemental phosphorus 

plant in southeastern Idaho (McKelvey and others, 1986). The 

occurrence of vanadium-rich shale with black-shale-hosted 

phosphorite deposits is common worldwide and suggests that 

steep and (or) fluctuating redox gradients existed within bottom 

waters and pore waters of the sedimentary basin (Piper, 1994).

The Julia Creek deposit in the Cretaceous Toolebuc 

Formation in Queensland, Australia, is an example of a 

vanadium-rich oil shale (Riley and Saxby, 1982; Patterson, 

Ramsden, and others, 1986). The oil shale was deposited in 

a shallow, epicontinental sea under reducing bottom water 
Geology U17

conditions that promoted the enrichment of vanadium. 

Concentrations range from 0.1 to 1.0 percent V2O5. Other 

known vanadium-rich (greater than 0.17 percent V2O5) oil 

shales include the Mississippian Heath Formation in Montana 

(Desborough and others, 1981; Derkey and others, 1985) 

and the Cretaceous La Luna Formation, which is a major 

petroleum source rock in Colombia and Venezuela (AlberdiGenolet and Tocco, 1999).

Associated metals.—Vanadium-rich black shales 

commonly contain high contents of other metals, such as 

silver, barium, cobalt, copper, molybdenum, nickel, phosphorus, PGEs, uranium, and zinc. Well-characterized deposits 

include the following (metals associated with vanadium are 

in parentheses): Julia Creek (molybdenum) in Queensland, 

Australia (Lewis and others, 2010); Nick (nickel, PGEs, and 

zinc) in Yukon Territory, Canada (Hulbert and others, 1992); 

and Viken (molybdenum, nickel, phosphorus, and uranium) 

and Häggån (uranium, molybdenum, and nickel) in Sweden 

(Hallberg, 2012; Aura Energy, 2012). As reported by Coveney 

and others (1992), carbonaceous and phosphatic black shales 

of Cambrian age in China, which have reported values 

exceeding 4 percent V2

O5, contain exceptionally high grades 

of nickel (from 2 to 4 percent) and molybdenum (>2 percent), 

and high concentrations of PGEs (20 to 80 parts per billion 

for platinum and palladium combined). Although vanadium 

has not been recovered from these strata, molybdenum and 

nickel have been mined on a small scale in China since about 

1985. Metamorphosed sulfide- and phosphate-rich black 

shale in southwestern Catalonia, Spain, contains unusually 

high concentrations of vanadium and chromium (as silicates 

and oxides) together with palladium and platinum minerals 

(Canet and others, 2003). The Talvivaara deposit in Finland 

has elevated contents of vanadium (averages 600 ppm) 

(Loukola-Ruskeeniemi and Heino, 1996; Loukola-Ruskeeniemi and Lahtinen, 2013), and until recently, was mined for 

cobalt, copper, nickel, and zinc, all of which were extracted 

using a bioheapleach mineral processing method (Jowitt and 

Keays, 2011; Saari and Riekkola-Vanhanen, 2012; LoukolaRuskeeniemi and Lahtinen, 2013). Vanadium is not currently 

recovered by this process, however.

Vanadate Deposits

Vanadates of lead, zinc, and copper (vanadinite and minerals 

of the descloizite-mottramite series) form in the oxidized zones 

of base-metal deposits, especially in areas of arid climate and 

deep oxidation (Fischer, 1975a). The copper-lead-zinc vanadate 

ores in the Otavi Mountainland of northern Namibia were 

once considered to be among the largest vanadium deposits 

in the world, with an estimated resource of several million 

metric tons (Boni and others, 2007). Other areas with known 

vanadate deposits include Angola, South Africa, Zambia (Broken 

Hill district, now known as the Kabwe Mine), and Zimbabwe 

(fig. U1). Small deposits occur in Argentina, Mexico, and the 

United States (Arizona, California, Nevada, and New Mexico), 

but are unlikely ever to be economically significant resources. 

Vanadates as a supply of vanadium essentially ceased in 1978 

when the last producing vanadium mine at Berg Aukas (Otavi 

Mountainland) in Namibia was closed.

Vanadate minerals and wulfenite (a lead molybdate 

mineral) within these deposits form crusts on open cavities 

or are intergrown with residual clays (Fischer, 1975a). The 

vanadate ores in the Otavi Mountainland occur in collapse 

breccias and solution cavities related to karst development in 

Neoproterozoic carbonate rocks of the Otavi Supergroup and 

are spatially associated with primary sulfide orebodies within 

the carbonate strata (Boni and others, 2007). Mottramite and 

copper descloizite are particularly abundant around copper 

sulfide deposits (Tsumeb type), whereas descloizite occurs in 

areas surrounding primary sphalerite-willemite (a zinc silicate 

mineral) orebodies (Berg Aukas type).

Vanadate deposits are secondary accumulations that form 

during supergene processes. The vanadate ores of the Otavi 

Supergroup are interpreted to have formed during several 

stages, preferentially within a karstic network. The vanadate 

ores, formed by low-temperature fluids related to weathering, 

are clearly distinct in age from that of the associated primary 

sulfide concentrations (Boni and others, 2007). The source 

of vanadium in such deposits is most likely the surrounding 

country rocks, especially shales (Fischer, 1975a), or in the 

case of the Otavi Mountainland deposits, mafic rocks of the 

older Paleoproterozoic basement (Boni and others, 2007).

Other Magmatic-Hydrothermal Vanadium Resources

Some magmatic-hydrothermal niobium-titanium deposits 

contain elevated concentrations of vanadium. Deposits at 

Potash Sulphur Springs (also called Wilson Springs) in 

Arkansas were the most important sources of vanadium in 

North America in the 1970s and 1980s, and nearly 4.3 million 

metric tons of 1.2 percent V2O5 was produced. By 1990, all 

the mines at Wilson Springs were closed (Howard and Owens, 

1995). The deposits are located within secondary enrichment 

zones and fenite that formed during and after intrusion of 

syenite and mafic alkalic igneous rocks (McCormick, 1978). 

The Wilson Springs deposits host a variety of vanadiumbearing minerals (Howard and Owens, 1995), including 

several minerals specific to these deposits (table U2). 

Adjacent carbonatite and alkaline igneous complexes at 

the Christy deposit within the Magnet Cove complex in 

Arkansas have high concentrations of vanadium together with 

titanium, niobium, and (or) rare-earth elements (Verplanck 

and Van Gosen, 2011; Flohr, 1994), and carbonatites and 

related rocks in Kenya are enriched in vanadium (Barber, 

1974). Typical vanadium concentrations in such deposits are 

about 1 percent and are contained in magnetite and titanium 

minerals. At Magnet Cove and Wilson Springs, sodic pyroxene 

and magnetite contain up to 3.19 and 1.43 weight percent 

V2O3, respectively, and high concentrations occur in goethite 

(Flohr, 1994; Howard and Owens, 1995).

Several other deposit types contain vanadium concentrations that are noteworthy, but all are presently uneconomic and 
U18 Critical Mineral Resources of the United States—Vanadium

are unlikely to be considered vanadium resources in the future. 

For example, heavy-mineral-concentrate samples from iron 

ores (Kiruna-type apatite-magnetite deposits) in Sweden and 

Chile have reported high concentrations (1,000 to 2,000 ppm) 

of vanadium (Nyström and Henríquez, 1994) and iron oxide 

mineral separates (magnetite and hematite) from such deposits 

contain up to 0.479 weight percent vanadium (Dupuis and 

Beaudoin, 2011). Vanadium concentrations of a few tenths of a 

percent are also common in titanium-bearing minerals, such as 

rutile and brookite (table U2) in some epithermal gold-silver 

and porphyry copper deposits (Fischer, 1973). For example, 

porphyry deposits in Australia contain rutile with vanadium 

contents of 0.2 and 1.3 weight percent (Scott, 2005), and rutile 

from the Pebble porphyry deposit in Alaska has vanadium 

concentrations that average 6.3 weight percent (Kelley and 

others, 2011). Even iron-rich minerals that do not contain 

titanium, such as magnetite and hematite, in some porphyry 

deposits contain elevated concentrations (up to 0.619 weight 

percent) of vanadium (Dupuis and Beaudoin, 2011).

In some gold-quartz veins, especially those containing 

gold-telluride minerals, roscoelite and other vanadium-bearing 

minerals are common as fine-grained intergrowths with quartz 

and other gangue minerals (Richards, 1995). In the Tuvatu 

gold-telluride deposit in Viti Levu, Fiji, vanadium occurs 

in roscoelite together with karelianite, vanadian muscovite, 

titanium-free nolanite, vanadian rutile, schreyerite, and an 

unnamed vanadium silicate mineral (Spry and Scherbarth, 

2006). Rutile in the Tuvatu deposit contains up to 5.2 weight 

percent V2

O3; roscoelite contains 32.71 weight percent 

V2O3, which is among the highest reported vanadium value 

for roscoelite from an epithermal gold-tellurium deposit 

(Spry and Scherbarth, 2006). The source of the vanadium 

is probably magnetite-bearing mafic alkalic igneous rocks 

that are spatially associated with the gold ores (Spry and 

Scherbarth, 2006).

Fossil Fuels

Vanadium closely correlates with organic carbon and, 

therefore, is enriched in many oil shales. It follows that 

significant amounts of vanadium are available for commercial 

use as a byproduct of petroleum, and minor amounts are 

produced as byproducts of coal and tar sands (Breit, 1992; 

Polyak, 2012, 2013). At least 10 percent of the world’s supply 

of vanadium comes from coal, crude oil, and petroleum 

(Polyak, 2012, 2013). The highest concentrations of vanadium 

are in heavy crude oils (Breit, 1992). Most of the world’s 

heavy oil and vanadiferous petroleum resources are located 

in Venezuela. These oilfields contain consistently high (up to 

1,400 ppm) concentrations of vanadium (Kapo, 1978). Other 

oils with greater than 50 ppm vanadium are produced in Iran 

and Japan, and from several fields within the United States, 

including Alaska, Arkansas, California, Louisiana, Mississippi, 

Oklahoma, Texas, and Wyoming. Vanadium is recovered from 

oil by processing ash generated in thermoelectric powerplants, 

petroleum coke residues generated during refining of heavy 

oils, and residues plated onto catalysts (Breit, 1992).

Vanadium abundances in ashes formed by burning coals 

generally range from 0.01 to 0.3 weight percent, with some 

as high as 8 percent (Reynolds, 1948). The lowest values are 

contained in coals formed from subaerial plant material; the 

highest concentrations are in marine sapropelic coals (Breit, 

1992). Although China is the only producer of vanadium 

from coal, coal deposits in Venezuela contain high vanadium 

contents. The average vanadium content of coal in the 

United States is 20 ppm (Swanson and others, 1976).

Tar sands are large deposits of bitumen or extremely 

heavy crude oil. The sands were originally named for those 

in the Athabasca region in northeastern Alberta, Canada, and 

tar sand deposits in this region are the best known examples 

in the world. Other documented occurrences are in Alabama, 

Alaska, California, Kansas, Kentucky, Oklahoma, Missouri, 

New Mexico, Texas, and Utah in the United States (Wever 

and Kustin, 1990), and in other countries, such as Jordan 

(Breit, 1992; Dill and others, 2009). The oil sands consist 

of a mixture of crude bitumen (a semisolid form of crude 

oil), silica sand, clay minerals, and water. Production of 

refinery-grade oil from the tar sand deposits generates a 

substantial amount of petroleum coke fly ash that may contain 

appreciable amounts of valuable metals, such as nickel, 

titanium, and vanadium. The amount and form of vanadium 

varies, depending on the nature of the sands. Tar sands in the 

Athabasca region average several hundred ppm vanadium, 

whereas other localities contain lower concentrations 

(Wever and Kustin, 1990; Breit, 1992).

Resources and Production

Identified Resources (United States and World)

A majority of the world’s supply of vanadium (approximately 80 to 85 percent) is derived from mined ore that 

comes either directly from deposits or from steelmaking slags 

produced by processing the ores mined from VTM deposits. 

The remaining 15 to 20 percent of the world’s supply of 

vanadium comes from (a) spent catalysts that collected 

vanadium during the refining of crude oils; (b) residues from 

the production of alumina, uranium, and some hydrocarbons; 

and (c) ash derived from burning high-vanadium-content 

coal or petroleum. Although the leading vanadium-producing 

countries in 2012 (fig. U2) were China (37 percent), 

South Africa (35 percent), and Russia (25 percent), Australia 

is poised to become a major world producer of vanadium in 

the future (Moskalyk and Alfantazi, 2003). World vanadium 

resources in 2012 were estimated to be 63 million metric 

tons of vanadium, and world reserves were estimated to be 

14 million metric tons (Polyak, 2013). Because vanadium 

is usually recovered as a byproduct or coproduct, the 

demonstrated world resources of this mineral commodity are 
Resources and Production U19

understated and. therefore. not fully indicative of available 

supply. Although domestic resources and secondary recovery 

are adequate to supply a large portion of domestic needs, a 

substantial part of U.S. demand is currently met by foreign 

material (Polyak, 2013).

Among the vanadium-rich ore deposits currently being 

mined, VTM deposits contain the largest tonnages of ore 

(table U1; fig. U5) and are the reason that China, South Africa, 

and Russia lead the world in vanadium production. Sandstonehosted deposits commonly have the highest grades of vanadium, but most are small; that is, they have less than 1 million 

metric tons of ore (table U1; fig. U5). Among the three deposit 

types, however, the leading producers of vanadium in the 

United States have been SSV deposits. In 2010, production 

from a mill facility in Utah that recovered vanadium from 

SSV ores was sufficient to meet 20 percent of the vanadium 

demand for the United States, which amounted to nearly 

2 percent of global vanadium production (Polyak, 2011).

Some carbonaceous and phosphatic shale-hosted deposits 

(Julia Creek, Australia; Viken, Sweden; and Green Giant, 

Madagascar) have reported high tonnages of vanadium, but 

these deposits are not currently in production. Technology to 

recover vanadium profitably in current markets is still being 

developed. Small amounts of vanadium were produced during 

World War II from the organic-rich Alum Shale in southern 

and central Sweden (Dyni, 2006). Also, until 1999, approximately 2,000 metric tons of V2

O5 was produced annually from 

the ferrophosphorus slag generated during the reduction of 

phosphate to elemental phosphorus in ore from the Phosphoria 

Formation in Idaho and Wyoming (Jasinski, 2004).

Bauxite is another source of vanadium. The 

vanadium content depends on the origin and nature of the 

bauxite, but average V2O5 contents typically range from 

0.05 to 0.25 percent. France, Germany, and India host 

important occurrences of vanadium-rich bauxite (Patterson, 

Kurtz, and others, 1986). During the processing of bauxite and 

the production of alumina, vanadium accumulates in residual 

sludge (red mud) and other byproducts that contain as much 

as 10 to 18 percent V2O5.

Vanadium is present in crude oil or the residues (bitumen 

or asphalts) of crude oils remaining in petroleum source 

rocks. Japan and the United States have recovered significant 

quantities of vanadium from petroleum residues. Asphaltine 

or bitumen in the Mina Ragra deposit in the Pasco Region of 

Peru, was extensively exploited for vanadium from 1907 until 

1955. It was the principal source of vanadium in the world in 

the early 1900s (Fischer, 1973). Unoxidized ore at Mina Ragra 

consists of quisqueite (a vanadium hydrocarbon) and patronite 

(vanadium sulfide) (table U2). Vanadates and vanadium 

oxide minerals form the near-surface oxidized ore, some of 

which is extremely rich and contains as much as 40 percent 

V2O5 (Fischer, 1973). Bitumen and asphaltine that contain 

vanadium-rich minerals occur in fractures and as disseminations in the host shale, forming a lens-shaped body (Fischer, 

1973).

Undiscovered Resources

Given that most of the present-day vanadium comes 

from VTM deposits and steelmaking slags produced from 

these deposits, it is likely that additional vanadium resources 

will come predominantly from VTM deposits and districts. 

VTM deposits in Australia, Brazil, Canada, Chile, India, 

and Malaysia have been identified only recently, and the 

full extent of the resources has not yet been evaluated fully. 

Among the largest regions with potential for widespread VTM 

resources is the 5,000-km2

 region known as the “Ring of Fire” 

in northern Ontario (fig. U1). Results from early exploration 

programs indicate that vanadium grades for deposits within 

mafic and ultramafic complexes in this region (for example, 

0.64 percent V2

O5 at the Butler Lake area, Ontario, Canada; 

MacDonald Mines Exploration Ltd., 2013) are comparable to 

economic deposits elsewhere in the world.

Some uranium SSV ores in southwestern South Dakota 

and northeastern Wyoming are part of the Wyoming basin’s 

uranium province and have recently been distinguished from 

Deposit tonnage, in million metric tons

0.01

0.1

1

10

0.0001 0.001 0.01 0.1 1 10 100

Sandstone-hosted vanadium

Shale-hosted vanadium

Vanadate

Vanadiferous titanomagnetite

Grade, in weight percent V O2 5

Vanadium figure 5

EXPLANATION

Deposit type

46

71

72

45

6

33 77

73

74 75

15

56

7 51

42

2

8 3

80

5 4

1 44 59 53

Figure U5. Plot of grade and tonnage of vanadium 

deposits for which data were available. Each symbol 

represents an individual deposit. Numbers next to the 

symbol correspond to the deposit numbers given in table U1 

and figure U1. The deposits with the highest grades 

(greater than or equal to 0.3 percent vanadium oxide [V2

O5]) 

and tonnages (greater than 1 million metric tons of V2

O5) 

are located in Africa, Australia, Canada, and China.
U20 Critical Mineral Resources of the United States—Vanadium

deposits elsewhere in the province by their substantially higher 

vanadium contents (George Breit, U.S. Geological Survey 

[retired], written commun., 2013). These are possible deposits 

for future vanadium production in the United States.

The amount of vanadium recovered from processing 

crude oils, coals, and tar sands will undoubtedly increase with 

future technological advances. Oil from large fields in the 

Caribbean Basin, the Middle East, and Russia are known to 

contain anomalously high vanadium contents (Mukhametshin 

and Punanova, 2011). Tar sands in North America may 

become an important resource of vanadium in the future. 

The effectiveness of recovery of vanadium is dependent on 

the silica and alumina contents of the fly ash; high silica and 

alumina contents tie up metal values in a silica-alumina matrix 

(Gomez-Bueno and others, 1981). Results of experimental 

work show that such matrixes can be broken down by 

application of a saline water roast of the carbon-free fly ash 

(Gomez-Bueno and others, 1981). Additional research on 

development of processing techniques of fly ash may lead 

to further enhanced recoveries of vanadium (Holloway and 

Etsell, 2006).

Iron sands are another potential source of vanadium. Iron 

sands are placer deposits that contain abundant concentrations 

of iron-bearing heavy minerals, including vanadium-rich 

magnetite. These iron sands are distributed extensively on the 

west coast of New Zealand’s North Island (Sweatman and 

others, 2012), but other placer occurrences most likely are 

present on the east coast of the conterminous United States 

and the coasts of Africa and Brazil, as well as fluvial irontitanium-vanadium placer deposits (for example, interior Iran; 

Razmara and Asadi, 2010). Vanadium concentrations in the 

New Zealand sands are as high as 0.5 percent. Although vanadium is currently not produced as a byproduct, research that 

is focused on methods to optimize its recovery is reportedly 

underway (Sweatman and others, 2012).

Most known occurrences of vanadium in shale are 

currently uneconomic but are estimated to contain large 

resources (table U1). Targets may become more viable in the 

future with advances in extraction technology. For example, 

development of a method to extract vanadium from the 

Gibellini vanadium deposit in central Nevada (American 

Vanadium Corp., 2012) is underway. Similar efforts to 

produce vanadium from metashales are reported for the 

Green Giant deposit in Madagascar (Energizer Resources, 

Inc., 2013). Graphitic deposits in Alabama and China (Liu 

Mao Mine) contain mica and garnet with elevated vanadium 

concentrations (up to 0.2 percent V2O5) that may be similar 

to the Green Giant deposit (Pallister and Thoenen, 1948; this 

volume, chap. J ). Other vanadiferous shales include the lowto medium-grade Okcheon metamorphic belt in Korea (Jowitt 

and Keays, 2011), where metalliferous black shales were 

used locally as fuel, and contain up to 2.04 percent vanadium, 

together with high concentrations of gold, molybdenum, 

nickel, PGEs, and uranium (Jeong, 2006). The Alum Shale in 

Sweden is an organic-rich marine sequence that is 15 to 60 m 

thick and contains up to 3,100 ppm vanadium and high 

concentrations of molybdenum, nickel, uranium, and zinc 

(Dyni, 2006). Although small amounts of vanadium have 

already been produced from the Alum Shale, new technologies 

may allow industrial-scale extraction from this laterally extensive unit (Dyni, 2006; Aura Energy, 2012; Hallberg, 2012). 

Black shales in Russia that occur within the Mongolia/Ural 

and Pacific gold belts have local concentrations of vanadium 

and uranium and therefore are also potential resources for 

these metals (Karpuzov and others, 2008). Although it lies at 

great depths (about 2,000 m below the surface), the oil-bearing 

Bazhenov Formation in the West Siberian Basin is another 

potential source of byproduct metals (Laznicka, 2010), 

with reported average concentrations of 105 ppm uranium, 

285 ppm molybdenum, and 1,015 ppm vanadium in a laterally 

extensive interval that is at least 15 to 20 m thick.

Exploration for New Deposits

Exploration methods for vanadium-rich ores vary 

with deposit type. An understanding of how each type of 

deposit forms is essential for predicting the potential for 

undiscovered deposits.

Exploring for Vanadiferous Titanomagnetite 

Deposits

Most vanadiferous titanomagnetite (VTM) deposits 

that formed on cratons of Archean to Proterozoic age are 

closely associated with LIPs; consequently, exploration for 

this deposit type is focused on LIPs. Mafic to ultramafic 

complexes that host VTM deposits are generally apparent as 

aeromagnetic anomalies, even if not exposed at the surface. 

Most commonly, airborne geophysical surveys that show 

coincident magnetic and electromagnetic anomalies are further 

investigated with followup ground geophysical surveys, 

including gravity, induced polarization, resistivity, and electromagnetic methods. Modeling of aeromagnetic and gravity 

data can indicate the extent, thickness, and form of intrusions 

(Cawthorn and Webb, 2001; Ivanic and others, 2010), which 

can help delineate intrusive bodies having the potential to host 

vanadium-rich magnetite layers. Paleomagnetic investigations 

of the Bushveld Complex have shown that different mineralized zones have different paleomagnetic signatures (that is, 

ages), allowing for regional correlation and delineation of the 

lateral extent of selected zones (Eales and others, 1993).

An important frontier issue in the exploration for 

VTM deposits is development of better exploration models 

that integrate the characteristics of these deposits with an 

improved understanding of LIP plumbing systems (Ernst and 

others, 2005). LIPs can have direct links to ore deposits (as 

hosts or heat engines) or indirect links, and can be used as 

guides for determining Precambrian paleocontinental reconstructions and related tracing of metallogenic belts between 

formerly adjacent tectonic blocks (Ernst and Peck, 2010). 
Exploration for New Deposits U21

Moreover, the compositions of igneous rocks in LIPs can be 

used to assess fertility and ore potential. For example, specific 

element ratios can be used to identify whether particular 

igneous rocks are permissive for hosting vanadium mineralization (Ernst and Peck, 2010). Lithogeochemistry, therefore, 

provides a guide for selecting the most prospective LIPs and 

LIP segments for exploration for VTM deposits, or for associated chromium, copper-nickel, or PGE deposits (Augé and 

others, 2005; Ivanic and others, 2010).

Major known ore-bearing magmatic intrusions are widespread within, or can be traced to the edge of, host continental 

blocks, suggesting that the intrusions likely continued into a 

formerly adjacent block. Examples include intrusions such 

as the Bushveld Complex (Kaapvaal craton), and the Great 

Dyke of Zimbabwe (Zimbabwe craton). From an exploration 

standpoint, robust Precambrian reconstructions aid in tracking 

well-documented metallogenic belts from heavily explored to 

frontier regions. New methods that use comprehensive data for 

LIPs can make such Precambrian reconstructions much more 

efficient (Ernst and Peck, 2010).

Exploring for Sandstone-Hosted Vanadium 

Deposits

All sandstone-hosted vanadium (SSV) deposits in 

the Colorado Plateau province contain uranium; therefore, 

methods used to explore for uranium deposits of this type are 

applicable to the search for vanadium. The recognition and 

documentation of meander bends and possible bifurcating 

paleochannels in the sandstone host rocks are important broad 

characteristics. Features that may be used as guides to ore 

include sandstone-mudstone facies, trunk channel systems, 

stratigraphic pinchouts, individual channels, thick sandstone 

lenses, and carbonaceous material (Kovschak and Nylund, 

1981). Models for the formation of SSV deposits suggest that 

bacteriogenic hydrogen sulfide acted as a reducing agent for 

ore formation (Northrop and others, 1990; Wanty and others, 

1990). In addition, pyrite in mineralized zones typically has 

lower sulfur isotope values than pyrite above or below the 

zone (Northrop and others, 1990), indicating that isotopically 

light pyrite is an indicator of mineralizing processes in an area.

Some deposits have readily recognizable alteration zones, 

whereas others show only subtle differences between unmineralized, altered, and mineralized zones. Principal criteria used 

to distinguish the alteration zones are color, pyrite morphology, 

kaolinization of feldspar, and low gamma-ray counts commonly 

accompanied by gamma-ray anomalies at the upper and lower 

margins of the altered zone (Rackley and others, 1968). The 

presence of limonite stains, green and blue secondary copper 

minerals, gray-green alteration of brown or reddish mudstone 

seams or lenses, and iron and (or) copper sulfide minerals 

were used as guides to ore in the 1950s (Johnson, 1959).

Geophysical methods assist in determining regional 

subsurface geology as it may relate to SSV deposits. 

Conspicuous gravity lows may indicate evaporite units that 

typically underlie the deposits, although many units may be 

too thin for detection. Magnetic anomalies highlight depths 

to basement, major faults that serve as conduits for ore fluids, 

and the presence of laccolithic intrusions that are interpreted 

as a possible source of metals in some deposits (Case and 

Joesting, 1972). More refined geophysical methods are needed 

to resolve the shape and position of channel sands. Airborne 

electromagnetic techniques have recently been applied in 

Nebraska to map the three-dimensional configuration of 

aquifers and paleochannels (Abraham and Cannia, 2011) and 

may be useful for delineating channels that are SSV ore hosts.

Exploring for Shale-Hosted Vanadium Deposits

High vanadium concentrations in shales are closely 

correlated with elevated contents (greater than 5 percent) 

of organic carbon. Many vanadiferous shales with more 

than 20 percent organic carbon are considered oil shales 

(for example, Julia Creek deposit, Queensland, Australia). 

Deposits of oil shale occur worldwide, and include major 

deposits in the United States (INTEK, Inc., 2011). Some of 

these oil shales are known to have high vanadium contents (for 

example, Heath Formation in Montana; La Luna Formation in 

Venezuela), but many others likely have not been analyzed for 

trace metals. Existing and new geochemical data for vanadium 

and other trace elements in shales would need to be examined 

to locate potential shale-hosted vanadium deposits.

A good understanding of how vanadiferous shales form is 

critical for exploration. It is generally accepted that the source 

of vanadium is seawater, but the enrichment mechanism is 

uncertain. The close occurrence of vanadium-rich shales with 

black-shale-hosted phosphate deposits suggests control by 

fluctuating redox gradients within a sedimentary basin (Piper, 

1994). If this suggestion is correct, then exploration will likely 

be focused in marine successions that were deposited along 

continental margins or inland seas, where redox gradients are 

likely to have occurred. Furthermore, because shales can be 

laterally extensive but have variable grades, understanding 

the distribution of metal enrichments may help target areas 

of interest.

National-scale geochemical datasets also may be useful 

in highlighting regions with black shales that are permissive 

for hosting vanadium deposits. Preliminary soil and sediment 

maps of the United States (Smith and others, 2015) show 

arcuate highs for a range of elements, including cobalt, 

molybdenum, nickel, and vanadium in central Kentucky—

an area that is underlain by the New Albany Shale (Ripley 

and others, 1990) and stratigraphically equivalent shales that 

are known to contain high concentrations of these metals. 

Although less pronounced, similar high concentrations are 

evident in east-central Nevada and could be related to shales 

of the metal-rich Woodruff Formation.

Geophysical methods may prove valuable for delineating 

the distribution of black shales, although many of these shales 

may not contain significant concentrations of vanadium. 
U22 Critical Mineral Resources of the United States—Vanadium

Shales are good conductors, particularly if graphite is present, 

and therefore, electromagnetic and resistivity surveys may 

prove useful in delineating the extent of shale units under 

shallow cover. Recent studies show that combined seismic and 

electromagnetic methods are effective in highlighting zones of 

high gas potential in shale reservoirs (Kumar and Hoversten, 

2012); such methods could be tested for use in the search for 

high-organic intervals in shales. Because many vanadium-rich 

carbonaceous shales also contain high uranium concentrations, 

downhole geophysical techniques (gamma logs) may be useful 

in identifying the position of metalliferous shales. Aerial 

radiometric surveys may also resolve the outcrop position of 

uranium-rich (and, by association, other metal-rich) shales 

(Pirkle and others, 1982). The arcuate trend of metals in 

surficial materials described above for Kentucky is readily 

mapped in an aeroradiometric survey (Hill and others, 2009).

Environmental Considerations

Sources and Fate in the Environment

Vanadium commonly occurs in one of three oxidation 

states in weathering environments: +3, +4, and +5. Vanadium 

+3 and +4 are relatively insoluble ions because they tend 

to form solid (oxy)hydroxides (for example V(OH)3 and 

VO(OH)), respectively, but the V 5+ state is generally dissolved 

in solution as various oxyanions (for example VO4

3–, HVO42–,

and H2VO4

–

). Under a wide pH range, these high-valence 

vanadium oxyanions may sorb to iron and aluminum 

(oxyhydr)oxide minerals. Most dissolved vanadium in rivers 

and streams derives from the weathering of silicate minerals 

(Shiller and Mao, 2000). Dissolved V4+ and V3+ are known 

to form strong complexes with organic compounds, many 

of which originated as V5+-organic complexes that were 

reduced. The formation of strong vanadium-organic compound 

complexes is one of the mechanisms by which fossil fuels and 

black shales may become enriched in vanadium.

Examples of natural concentrations of vanadium in rocks, 

soils, water, and air are given in table U3. Vanadium contents 

in soils away from known vanadium deposits, mines, or 

smelters range from 13 to 227 ppm (Shacklette and Boerngen, 

1984; Tyler, 2004), whereas streams and rivers contain vanadium concentrations that range from 0.23 to 3.7 micrograms 

per liter (μg/L) or ppb (Shiller and Mao, 2000; Gaillardet and 

others, 2003). The concentration of vanadium in suspended 

particulates in world rivers averages 129 mg/kg (Viers and 

others, 2009). Dissolved vanadium in coastal seawater ranges 

from 0.31 to 2.8 μg/L (Shiller and Mao, 1999; Wang and 

Sañudo-Wilhelmy, 2008; Strady and others, 2009). In North 

Pacific seawater, dissolved vanadium concentrations generally increase with depth with a total range of 1.5 to 1.9 μg/L 

(Collier, 1984). In contrast, concentrations in particulate 

matter (with diameters greater than 53 micrometers) decrease 

in the upper 100 to 200 m of seawater, then remain relatively 

constant over a range of 0.0001 to 0.0004 μg/L (Collier, 1984). 

Among nearly 9,000 samples of filtered and unfiltered groundwaters collected in California, 90 percent contained vanadium 

concentrations ranging from 3 to 24 μ/L (Wright and Belitz, 

2010). Vanadium occurs naturally in the atmosphere as part 

of mineral dust particles, and ranges in concentration from 

0.0006 to 0.002 nanogram per cubic meter (ng/m3) over the 

South Pole and from 0.8 to 1.4 ng/m3 over Greenland (KabataPendias and Pendias, 2001).

Mining and industrial activities can lead to abovebackground concentrations of vanadium in the environment 

(table U3). For example, the concentration of vanadium in 

soils (548 to 7,160 ppm) collected near a vanadium mine in 

the North West Province of South Africa (Mandiwana and 

Panichev, 2004) is roughly an order of magnitude greater than 

natural concentrations in soils. In the Panzhihua region, China, 

soil vanadium concentrations depend upon the predominant 

land use and decrease in the following order: smelting 

(208 to 938 ppm), mining (112 to 591 ppm), agricultural 

(86 to 227 ppm), and urban use (94 to 184 ppm) (Teng and 

others, 2011). The Bílina River in the Czech Republic receives 

drainage from municipal and industrial areas; vanadium 

concentrations range from 2.5 to 85.6 μg/L, with the highest 

concentrations observed in the industrial region (Kohušová 

and others, 2011). Primary inputs of vanadium to the atmosphere result from mining, ore processing, and combustion of 

fuel oils and coal. Vanadium concentrations in air in urban and 

industrial areas range from 5 to 200 ng/m3 (Kabata-Pendias 

and Pendias, 2001).

Mine Waste Characteristics

Mine waste is generally considered to be the material that 

originates and accumulates at a mine site but has no current 

economic value (Lottermoser, 2010), and it includes both 

solid and liquid waste. Because vanadium can be recovered 

as a byproduct of the mining of bauxite, carnotite, phosphate, 

and titanomagnetite ores, the character of the mine waste will 

vary according to the methods used to extract the primary 

ore. Vanadium mining of black shales and copper-lead-zinc 

vanadate deposits, as well as recovery of vanadium from fly 

ash, generates additional types of mine and processing waste. 

Mining of vanadium from VTM deposits in Australia is 

projected to generate 12 million cubic meters of waste rock at 

the Windimurra deposit (Environmental Protection Authority, 

2008) and 59 million metric tons of tailings at the Balla 

Balla deposit (Environmental Protection Authority, 2009). 

American Vanadium Corp. is planning to mine vanadium 

from the Gibellini vanadium deposit in Nevada. Production 

is to involve open pit mining and processing on site by heap 

leaching; the total amount of ore and waste to be extracted 

is projected to be 24 million metric tons (Hanson and others, 

2011). Plans to mine vanadium at the metashale-hosted Green 

Giant deposit in Madagascar includes a well-known process 

called alkaline press leaching, which is expected to produce 

a battery-grade form of V2

O5 that has a purity of greater than 

99.5 percent.
Environmental Considerations U23

Table U3. Vanadium concentrations in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency; V, vanadium; cm, centimeter; μg/L, microgram per liter; μm, micrometer; 

ng/m3, nanogram per cubic meter; ppm, part per million; %, percent]

Environment

and (or) location

Vanadium

 concentration Unit Comments Reference(s)

Rocks

Upper continental crust 60 ppm Average Taylor and McLennan (1995)

Bulk continental crust 230 ppm Average Taylor and McLennan (1995)

Lower continental crust 285 ppm Average Taylor and McLennan (1995)

Basalt 250 ppm Average Levinson (1974, p. 44)

Black shale 205 ppm Average Ketris and Yudovich (2009)

Granite 20 ppm Average Levinson (1974, p. 44)

Limestone 15 ppm Average Levinson (1974, p. 44)

Shale 130 ppm Average Levinson (1974, p. 44)

Soils

Western United States 70 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

Eastern United States 43 ppm Mean for 20 cm depth Shacklette and Boerngen (1984)

North West Province, 

South Africa

548 to 7,160 ppm Composite of upper 20 cm; 

proximal to V mining

Mandiwana and Panichev (2004)

Panzhihua, China 94 to 184 ppm Composite of upper 20 cm; 

urban soil

Teng and others (2011)

Panzhihua, China 86 to 227 ppm Composite of upper 20 cm; 

agricultural soil

Teng and others (2011)

Panzhihua, China 112 to 591 ppm Composite of upper 20 cm; 

proximal to V mining

Teng and others (2011)

Panzhihua, China 208 to 938 ppm Composite of upper 20 cm; 

proximal to V smelting

Teng and others (2011)

Sweden 13 to 47 ppm Range of profile developed

on quartzite and gneiss

Tyler (2004)

Proposed DOE benchmark 2 ppm Screening benchmark for 

terrestrial plants

Efroymson and others (1997)

Soil-quality guideline 130 ppm Canadian agricultural soil 

guideline

Canadian Council of Ministers 

of the Environment (2007)

Waters

Seawater, North Pacific 1.5 to 1.9 μg/L Dissolved Collier (1984)

Seawater, North Pacific 0.0001 to 0.0004 μg/L Particulate (<53 μm) Collier (1984)

Seawater, coastal New York, 

United States

0.31 to 1.78 μg/L Dissolved (<0.2 μm) Wang and Sañudo-Wilhelmy (2008)

Seawater, Louisiana Shelf, 

United States

0.32 to 1.70 μg/L Dissolved and colloidal 

(<0.45 μm)

Shiller and Mao (1999)

African rivers 0.59 to 0.65 μg/L Dissolved load (< 0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

European rivers 0.4 to 2.9 μg/L Dissolved load (< 0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

North American rivers 0.4 to 1.84 μg/L Dissolved load (< 0.2 μm); 

nonpolluted

Gaillardet and others (2003) 

and references therein

Amazon River 0.703 μg/L Mean dissolved load 

(<0.2 μm); nonpolluted

Gaillardet and others (2003) 

and references therein

Bílina River 2.5 to 85.6 μg/L Receives municipal and 

industrial drainage

Kohušová and others (2011)
U24 Critical Mineral Resources of the United States—Vanadium

Table U3. Vanadium concentrations in rocks, soils, waters, and air.—Continued

[DOE, U.S. Department of Energy; EPA, U.S. Environmental Protection Agency; V, vanadium; cm, centimeter; μg/L, microgram per liter; μm, micrometer; 

ng/m3, nanogram per cubic meter; ppm, part per million; %, percent]

Environment 

and (or) location

Vanadium 

 concentration Unit Comments Reference(s)

Waters—Continued

Mississippi River tributaries, 

United States

Stream water, California, 

United States

Gironde Estuary, France

Groundwater, California

Sediment, world river (average)

Proposed EPA benchmark

Proposed EPA benchmark

0.23 to 1.77

0.29 to 3.7

0.89 to 2.76

3 to 24

129

280

20

μg/L

μg/L

μg/L

μg/L

ppm

μg/L

μg/L

Dissolved and colloidal 

(<0.4 μm)

Dissolved and colloidal 

(<0.4 μm)

Dissolved

Filtered and unfiltered; range

is for 90% of samples

Suspended sediment

Tier II Secondary acute value

Tier II Secondary chronic value

Shiller and Mao (2000)

Shiller and Mao (2000)

Strady and others (2009)

Wright and Belitz (2010)

Viers and others (2009)

Suter and Tsao (1996)

Suter and Tsao (1996)

Air

South Pole

Greenland

Urban/industrial

0.0006 to 0.002

0.8 to 1.4

5 to 200

ng/m3

ng/m3

ng/m3

None

None

None

Kabata-Pendias and Pendias (2001)

Kabata-Pendias and Pendias (2001)

Kabata-Pendias and Pendias (2001)

The mineralogy of solid mine waste derived from 

vanadium mining is similar to that of the deposit from which 

it is extracted, but the proportion of vanadium minerals is 

smaller. Tailings from the VTM deposit in Panzhihua, China, 

contain ilmenite (15 to 18 mass percent), augite (46 mass 

percent), plagioclase (31 to 34 mass percent), and pyrite 

(2 to 3 mass percent) (Dahe, 2004). The Slick Rock SSV 

deposit in Colorado typically contains low contents of uranium 

and vanadium ore minerals, as well as anatase, barite, calcite, 

jordisite, pyrite, and quartz; selenium- and copper-bearing 

minerals are found in some deposits (Shawe, 2011). In the 

Gibellini vanadium deposit, vanadium is concentrated in 

organic material; associated phases are apatite, calcite, clay, 

microcline, pyrite, and sphalerite (Desborough and others, 

1979). Copper-lead-zinc vanadate deposits may contain 

the ore minerals descloizite, mottramite, and vanadinite 

(table U2), which at the Otavi Mountainland deposits in 

Namibia host trace amounts of arsenic, cadmium, chromium, 

iron, manganese, molybdenum, nickel, and phosphorus (Boni 

and others, 2007). The most common gangue minerals in the 

Otavi Mountainland deposits are calcite and dolomite, which 

have high concentrations of iron, phosphorus, and lead and 

trace amounts of arsenic, barium, cadmium, copper, nickel, 

strontium, and vanadium (Boni and others, 2007).

In weathering environments, trace metals in vanadium 

deposits are expected to remain immobile under oxic, nearneutral pH conditions (5<pH<8), especially in the presence 

of iron-rich solids (Smith and Huyck, 1999). Under anoxic 

and (or) acidic conditions, however, vanadium minerals, 

some gangue minerals, and associated trace metals may 

dissolve. In those deposits that contain sulfide minerals (for 

example, chalcocite, pyrite, and sphalerite), generation of 

acidity is possible if sulfide dissolution is not balanced by the 

acid-neutralizing capacity of carbonate minerals (for example 

calcite and dolomite). The dissolution of sulfide minerals 

releases metals and produces sulfuric acid, and the subsequent 

acidic pH values of the solutions allow higher concentrations 

of metals to be dissolved—potentially causing the environmental problem known as acid mine drainage (AMD). Mining 

of VTM, SSV, and black shale deposits that contain sulfides 

and lack appreciable quantities of carbonate minerals increases 

the potential for AMD generation. Alternatively, some of these 

deposit types, particularly some SSV and black shale deposits, 

contain appreciable amounts of carbonate minerals, thereby 

lowering the acid-generation potential.

Recovery of vanadium from mine tailings and other 

wastes is becoming an increasingly important source, 

particularly because the fraction of vanadium in discarded 

products that are recycled is less than 1 percent (Graedel and 

others, 2011). The main challenge of vanadium recycling is 

that it is often included in alloys in small amounts, making 

recovery technologically and economically unfeasible (Reck 

and Graedel, 2012).

Human Health Concerns

The Agency for Toxic Substances and Disease Registry 

provides a useful summary of the human toxicology of vanadium (Agency for Toxic Substances and Disease Registry, 2012). 
Environmental Considerations U25

The general public is most likely to be exposed to vanadium 

through consumption of contaminated food. Occupational 

exposure to vanadium usually results from the inhalation of 

V2O5 dust during the production of FeV and steel. Exposure 

to high oral doses of vanadium may lead to nausea, diarrhea, 

stomach cramps, decreased numbers of red blood cells, and 

increased blood pressure. Inhalation of V2O5 may cause 

extended coughing and is suspected to cause cancer.

Vanadium is believed to be a micronutrient, with a postulated requirement for humans of less than 10 micrograms per 

day, which can be met through dietary intake (Anke, 2004, and 

references therein). Vanadium in the form of vanadyl sulfate 

and sodium metavanadate has been administered to diabetic 

patients as a dietary supplement because these compounds 

have been observed to mimic the actions of insulin in isolated 

cell systems (Anke, 2004, and references therein), but clear 

therapeutic benefit has yet to be established (Wiernsperger and 

Rapin, 2010). Primary and secondary drinking-water regulations for vanadium currently do not exist in the United States, 

but the Occupational Safety and Health Administration has set 

an exposure limit of 0.05 milligrams per cubic meter (mg/m3) 

for V2O5 dust (Occupational Safety and Health Administration, 

2013b) and 0.1 mg/m3 for V2O5 fumes (Occupational Safety 

and Health Administration, 2013a) in workplace air over 

an 8-hour workday. Vanadium toxicity is believed to result 

from an intake of more than 10 to 20 milligrams per day 

(Anke, 2004).

Ecological Health Concerns

Vanadium is essential for some biological processes 

and organisms, as well as a potential toxicant. For example, 

some nitrogen-fixing bacteria require vanadium for producing 

vanadium nitrogenase, an enzyme used to convert nitrogen 

(N2) from the atmosphere into ammonia, which is a more 

biologically accessible form of nitrogen (Madigan and others, 

2003). This process is critical to the health of the biosphere. 

Vanadium is also essential to certain species of algae for 

chlorophyll production and overall growth (Anke, 2004).

Compared with other metals and metalloids, the ecological impacts of vanadium in the environment are not well 

known. Because fish tend to be sensitive to low concentrations 

of dissolved metals, they are often indicators of contamination 

in aquatic systems. One of several useful endpoints used in 

toxicity tests is that which determines the lethal concentration 

that leads to 50 percent mortality (LC50) after exposure to a 

substance for a specified amount of time. Chronic toxicity tests 

(that is, lower doses over longer time periods) of dissolved 

vanadium in developing rainbow trout (Oncorhynchus 

mykiss) revealed a mean LC50 value of 0.17 milligrams per 

liter (mg/L) after 28 days of exposure (Birge and others, 

1980). Acute toxicity tests (higher doses over shorter times) 

of vanadium to a fish known as the threespine stickleback 

(Gasterosteus aculeatus) revealed LC50 values that ranged 

from 2.4 to 4.1 mg/L after 4 days of exposure (Gravenmier 

and others, 2005). Similar acute toxicity tests of vanadium to 

three species of endangered fish from the Green River, Utah, 

which runs through carnotite deposits, showed no difference 

in sensitivity among species, but LC50 values ranged from 

5.3 to 8.8 mg/L for young specimens and 2.2 to 5.1 mg/L for 

older stages (Hamilton, 1995) after 4 days of exposure. These 

dissolved vanadium concentrations are much greater than 

those observed in many rivers and streams (table U3). Some 

regions of the United States have adopted secondary acute and 

chronic screening benchmarks of 280 micrograms of vanadium per liter of water (μg V/L) and 20 μg V/L, respectively, 

for aquatic freshwater life (Suter and Tsao, 1996).

The essentiality of vanadium to higher plants is debated, 

but clear evidence exists for vanadium phytotoxicity. After 

amending soils with vanadium in the form of dissolved ammonium metavanadate (NH4

VO3), Wang and Liu (1999) found 

that soybean seedling growth was markedly stunted in flood 

plain soils (fluvaquents) with greater than 30 ppm of vanadium, whereas little to no stunting occurred in red earth soil 

(Oxisol), even at vanadium concentrations as high as 75 ppm. 

Likewise, in phytotoxicity tests with forb (a herbaceous 

flowering plant), crop, and grass species, LC50 values ranged 

from 21 to 59 ppm of V2O5 after exposure for 4 to 5 weeks; 

higher LC50 values (90 to greater than 130 ppm of V2O5) 

were observed under similar conditions, but with greater 

soil nutrient levels (Smith and others, 2013). These studies 

highlight how vanadium toxicity to plants varies, depending 

upon the prevailing soil conditions, the type of plant species, 

and the chemical form of vanadium, which determines 

its bioaccessibility. The chemical form of vanadium in 

ecotoxicity studies may be more bioaccessible than the 

chemical form of vanadium in soils. Furthermore, the amount 

of bioaccessible vanadium is likely to be smaller than the 

amount of total vanadium in soils. For example, the concentration of bioaccessible vanadium in soils of Northern Ireland 

was determined to be about 7.5 to 17 percent of the total soil 

vanadium content (Barsby and others, 2012). Some regions 

of the United States have adopted a soil screening benchmark 

of 2 ppm of vanadium for terrestrial plants (Efroymson and 

others, 1997). The Canadian agricultural soil quality guideline 

for vanadium is 130 ppm of vanadium (Canadian Council of 

Ministers of the Environment, 2007).

Mine Closure

Most recent and new mining operations include closure 

plans that address issues related to the mine footprint. A mine’s 

footprint includes the waste left on site and the locally affected 

soil and water, as well as ecological impacts, such as habitat 

destruction and loss of biodiversity. Mine closure issues 

that could have the greatest environmental impacts depend 

upon the type of deposit being mined, and if applicable, the 

methods employed to process the ore on site. Given the variety 

of deposit types from which vanadium is obtained, all mine 

closure issues related to vanadium mining are too numerous to 
U26 Critical Mineral Resources of the United States—Vanadium

describe here. Instead, the focus is on potential issues related 

to mining of the largest and economically most important 

deposit type (VTM deposits), and to the new mining operation 

planned in the Gibellini vanadium deposit in Nevada.

Some VTM and shale-hosted deposits contain sulfide 

minerals, and, therefore, the potential for AMD exists at both 

types of mine sites. Acidic drainage may seep from waste piles 

or tailings ponds. Common methods for treating AMD include 

active water-treatment facilities, passive limestone-lined 

channels, and constructed wetlands (Plumlee and Logsdon, 

1999). The end result of both active and passive approaches 

is eventual precipitation of dissolved metals. Precipitated 

metals in passive wetland systems tend to be more stable 

under the prevailing anoxic conditions, whereas the metal-rich 

precipitates that result from active treatment facilities form a 

sludge that can cause environmental problems if not disposed 

of responsibly. When the potential for AMD exists at a mine 

site, common preventative measures include conducting water 

quality surveys before, during, and after mining.

The great size of VTM deposits tends to result in large 

mine waste piles. These waste piles have the potential to 

become unstable and can be a source of metal-rich dust. If 

using mine waste as backfill into dry mine workings is not 

an option, problems of mine-waste-pile stability and dust 

generation can often be addressed through grading and 

covering the piles with vegetation.

Vanadium-rich black shale at the Gibellini vanadium 

deposit is planned to be processed on site using acid heap 

leaching and solvent extraction—a recovery process that has 

yet to be applied to vanadium ores (Hanson and others, 2011). 

Acid heap leaching is commonly used on low-grade cobalt, 

copper, and nickel ores. In general, heap leaching involves 

placing crushed ores on top of impervious liners to form a 

slightly sloped leach pad, applying leach solutions to the pad, 

then collecting the leachates delivered by the impervious liners 

into ponds or tanks for further processing. During solvent 

extraction, leachates are mixed with a vanadium-optimized 

organic extractant (that is, a chelating agent), thereby forming 

vanadium-organic complexes that can be separated from 

other undesired metals present in the solution. At the Gibellini 

vanadium deposit, the resulting metal solution is planned to 

be recirculated back through the leaching process, while the 

vanadium-organic solution will undergo acidic stripping of 

vanadium from the organics, followed by precipitation of V2

O5

(Hanson and others, 2011). Potential seepage or spillage of 

leachates into local surface and groundwater can have negative 

impacts on the environment, however, given that leachates 

may contain metals, such as arsenic, cadmium, chromium, 

copper, mercury, lead, and zinc (U.S. Environmental Protection Agency, 1994). Best practices for this type of operation 

include engineering for leak prevention, ongoing seepage 

testing, and post-closure monitoring of mine waste discharges 

and downstream water quality (U.S. Environmental Protection 

Agency, 1994).

Problems and Future Research

Infrastructure development and the need for steel products are the main sources of demand for vanadium worldwide. 

The current worldwide resources of about 63 million metric 

tons of vanadium appear to be adequate to meet current 

demand. Future demand is expected to increase because of the 

following factors (Roskill Information Services, Ltd., 2013):

• China is the top steel producer and consumer of 

vanadium in the world (the country accounted for 

about 34 percent of the world’s vanadium consumption 

in 2012), and China’s demand for vanadium to use in 

the production of steel is expected to remain strong 

for the next 10 to 20 years.

• China and Japan have legislated increased vanadium 

content in steel rebar so that the quality of their steel 

matches that of other major steel-producing countries; 

therefore, use of vanadium for steel production in 

China and Japan is expected to increase dramatically.

• India’s steel production is projected to almost double 

owing to industrialization of that country.

• Development of alternate renewable sources of energy 

will likely require increased use of VRBs.

With the anticipated increase in demand for vanadium and 

the limited supply, maintaining a constant supply would likely 

mean that new sources of vanadium would need to be identified 

and the extraction of vanadium from currently defined sources 

would need to be optimized. Many future resources are likely 

contained within VTM deposits in unexplored regions. Another 

potential source is vanadium-rich shale because known 

deposits of this type have large tonnages and grades that are 

similar to those of presently mined VTM deposits (fig. U5). 

For shale-hosted deposits to be a viable resource, however, 

methods of profitably extracting the vanadium and producing 

a product that could be used in VRBs would be required, and 

future research would need to be designed to optimize these 

methods. Other sources, such as iron sands (placers) likely 

exist in many parts of the world; identification and quantification of these sands in terms of tonnage and vanadium content 

could be undertaken. Research on optimizing the recovery 

of vanadium from crude oil, bauxite, and tar sands also may 

increase global identified resources of this valuable metal.

Acknowledgments

We would like to thank USGS geologists John Slack 

and George Breit for their helpful technical reviews of this 

manuscript. John DeYoung and Suzanne Nicholson also 

provided useful reviews of the manuscript. Discussions with 

George Breit greatly aided our understanding of fossil fuel and 

shale-hosted vanadium deposits. Keith Labay is thanked for 

help with drafting the figures.
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Zirconium and Hafnium

By James V. Jones III, Nadine M. Piatak, and George M. Bedinger

Chapter V of

Critical Mineral Resources of the United States—Economic and 

Environmental Geology and Prospects for Future Supply

Edited by Klaus J. Schulz, John H. DeYoung, Jr., Robert R. Seal II, and Dwight C. Bradley

Professional Paper 1802– V

U.S. Department of the Interior

U.S. Geological Survey
Contents of Chapter V

[Figure and table titles for all chapters are listed in the volume contents before the conversion factors table and list of 

abbreviations and symbols. A periodic table of the elements is on the inside front cover of the printed volume]

Abstract........................................................................................................................................................V1

Introduction..................................................................................................................................................V2

Uses and Applications ......................................................................................................................V2

Recent Aspects of Supply and Demand ........................................................................................V2

Geology.........................................................................................................................................................V5

Geochemistry and Mineralogy ........................................................................................................V5

Deposit Types .....................................................................................................................................V6

Primary Igneous Deposits .......................................................................................................V6

Secondary Deposits—Coastal Placers and Paleoplacers ................................................V7

Resources and Production ......................................................................................................................V12

Identified Resources .......................................................................................................................V12

Other Resources ..............................................................................................................................V13

Exploration for New Deposits .................................................................................................................V13

Environmental Considerations ................................................................................................................V14

Sources and Fate in the Environment ..........................................................................................V14

Mine Waste Characteristics ..........................................................................................................V14

Volume of Mine Waste ...........................................................................................................V14

Mineralogy and Chemistry of Slimes and Tailings ............................................................V15

Acid-Base Accounting ...........................................................................................................V16

Human Health Concerns .................................................................................................................V17

Ecological Health Concerns ...........................................................................................................V17

Mine Closure.....................................................................................................................................V18

Problems and Future Research ..............................................................................................................V18

References Cited.......................................................................................................................................V19

Suggested citation:

Jones, J.V., III, Piatak, N.M., and Bedinger, G.M., 2017, Zirconium and hafnium, chap. V of Schulz, K.J., DeYoung, J.H., Jr.,

Seal, R.R., II, and Bradley, D.C., eds., Critical mineral resources of the United States—Economic and environmental 

geology and prospects for future supply: U.S. Geological Survey Professional Paper 1802, p. V1– V26, http://dx.doi.org/

10.3133/pp1802V.
Zirconium and Hafnium

By James V. Jones III, Nadine M. Piatak, and George M. Bedinger

Abstract

Zirconium and hafnium are corrosion-resistant metals that 

are widely used in the chemical and nuclear industries. Most 

zirconium is consumed in the form of the main ore mineral 

zircon (ZrSiO4

), or as zirconium oxide or other zirconium 

chemicals. Zirconium and hafnium are both refractory 

lithophile elements that have nearly identical charge, ionic 

radii, and ionic potentials. As a result, their geochemical 

behavior is generally similar. Both elements are classified as 

incompatible because they have physical and crystallochemical

properties that exclude them from the crystal lattices of most 

rock-forming minerals. Zircon and another, less common, 

ore mineral, baddeleyite (ZrO2), form primarily as accessory 

minerals in igneous rocks. The presence and abundance of 

these ore minerals in igneous rocks are largely controlled 

by the element concentrations in the magma source and by 

the processes of melt generation and evolution. The world’s 

largest primary deposits of zirconium and hafnium are associated with alkaline igneous rocks, and, in one locality on the 

Kola Peninsula of Murmanskaya Oblast, Russia, baddeleyite 

is recovered as a byproduct of apatite and magnetite mining. 

Otherwise, there are few primary igneous deposits of 

zirconium- and hafnium-bearing minerals with economic value 

at present. The main ore deposits worldwide are heavy-mineral 

sands produced by the weathering and erosion of preexisting 

rocks and the concentration of zircon and other economically 

important heavy minerals, such as ilmenite and rutile (for 

titanium), chromite (for chromium), and monazite (for rareearth elements) in sedimentary systems, particularly in coastal 

environments. In coastal deposits, heavy-mineral enrichment 

occurs where sediment is repeatedly reworked by wind, waves, 

currents, and tidal processes. The resulting heavy-mineral sand 

deposits, called placers or paleoplacers, preferentially form 

at relatively low latitudes on passive continental margins and 

supply 100 percent of the world’s zircon. Zircon makes up a 

relatively small percentage of the economic heavy minerals 

in most deposits and is produced primarily as a byproduct of 

heavy-mineral sand mining for titanium minerals.

From 2003 to 2012, world zirconium mineral 

concentrates production increased by more than 40 percent, 

and Australia and South Africa were the leading producers. 

Global consumption of zirconium mineral concentrates 

generally increased during the same time period, largely as 

a result of increased demand in developing economies in 

Asia and the Middle East. Global demand weakened in 2012, 

causing a decrease in world production of zirconium mineral 

concentrates and delaying the development of several new 

mining projects. Global consumption is expected to increase in 

the future, however, as demand from the ceramics, chemicals, 

and metals industries increases (driven by renewed growth 

in developing economies) and demand for zirconium and 

hafnium metal increases (driven by the construction and 

operation of new nuclear powerplants).

The behaviors of zirconium and hafnium in the 

environment are very similar to one another in that most 

zirconium- and hafnium-bearing minerals have limited 

solubility and reactivity. Anthropogenic sources of zirconium, 

and likely hafnium, are from industrial zirconium-containing 

byproducts and emissions from the processing of sponge 

zirconium, and exposure to the general population from 

these sources is small. Zirconium and hafnium are likely not 

essential to human health and generally are considered to 

be of low toxicity to humans. The main exposure risks are 

associated with industrial inhalation and dermal exposure. 

Because of the low solubility of zirconium and hafnium, 

ecological health concerns in the aquatic environment and 

in soils are minimal. Heavy-mineral sand mining may lead 

to increased erosion rates when the mining is managed 

improperly. In addition, surface mining requires removal of 

the overlying organic soil layer and produces waste material 

that includes tailings and slimes. The soil removal and mining 

activity disturbs the surrounding ecosystem and alters the 

character of the landscape. Dry mineral separation processes 

create high amounts of airborne dust, whereas wet mineral 

separation processes do not. In operations that restore the 

landscape to pre-mining conditions, the volume of waste and 

the impact on the landscape may be relatively temporary.
V2 Critical Mineral Resources of the United States—Zirconium and Hafnium

Introduction

Uses and Applications

Zirconium (Zr) and hafnium (Hf) are metals (figs. V1A

and V1B, respectively) that are used in the chemical and 

nuclear-reactor industries in applications for which corrosion resistance, structural stability at high temperatures, 

and specific alloying properties and (or) specific neutronabsorption characteristics are required. The main ore mineral 

for both elements is zircon (ZrSiO4) (figs. V1C and V1D), 

which typically has a zirconium to hafnium (Zr:Hf) ratio of 

50:1. Baddeleyite (ZrO2) (fig. V1E) is a less common but 

commercially important mineral that contains both zirconium 

and hafnium, but in a ratio of approximately 73:1. Zircon 

and baddeleyite are formed primarily as accessory minerals 

in igneous rocks and are typically present at relatively low 

abundances. They also occur naturally in a wide variety of 

rock types and geologic environments. The main ore deposits 

worldwide are heavy-mineral sands (figs. V1F and V1G) 

produced by the weathering and erosion of preexisting rocks 

and the concentration of zircon in sedimentary systems, 

particularly in coastal environments.

Approximately 95 percent of all zirconium consumed 

is in the form of zircon, zirconium oxide, or other zirconium 

chemicals (Nielson and Wilfling, 2002; Gambogi, 2012b). 

Zircon is highly refractive, with a melting point of 

2,550 degrees Celsius (°C) or greater. Thus, zircon is used 

for facings on foundry molds, and milled or ground zircon 

is used in refractory paints for coating the surfaces of molds. 

Zircon bricks and blocks are used in furnaces and hearths 

for containing molten metals, and glass tank furnaces use 

fused-cast and bonded alumina-zirconia-silica-based refractory 

metal. Zircon is also used in minor quantities as a gemstone 

and may be processed to produce cubic zirconia, which is 

a synthetic gemstone and diamond simulant. Baddeleyite is 

used in the manufacture of alumina-zirconia abrasives and in 

ceramic colors and refractories.

Zirconia, which is a form of zirconium oxide, is produced 

by reacting zircon ore with caustic soda and then chlorinating 

it to produce zirconium oxychloride. The zirconium oxychloride

is then calcined and precipitated to form zirconia powder. 

Zirconia has a high melting point (2,700 °C) and low thermal 

conductivity, and it is an important constituent of ceramic 

colors and glazes. Zirconia undergoes a phase transformation 

during heating, however, and must be stabilized for other uses 

by adding such oxides as calcium oxide (CaO), magnesium 

oxide (MgO), and yttrium oxide (yttria; Y2O3) into the zirconia 

structure. Yttria-stabilized zirconia is used in the manufacture 

of cubic zirconia, fiber-optic connector components, refractory 

coatings, engineering and structural ceramics, and oxygen 

sensors that control combustion in automobile engines. 

Because it has increased fracture resistance and strength 

relative to similar alumina products, yttria-stabilized zirconia 

is also used in dental applications, such as bridges, crowns, 

and inlays, and in joint-replacement procedures. Other 

zirconium chemicals include the following: zirconium boride 

used as a diffusion barrier in semiconductors, as a container 

for molten metals, and as a burnable absorber in nuclear 

reactor cores; zirconium carbonates used in antiperspirants 

and in printing and paper manufacturing; zirconium hydride 

used for flares, fuses, and combustion charges in pyrotechnics 

and as a binding or brazing component in abrasive wheels 

and polishing discs; zirconium nitride used as a wear-resistant 

coating on drill and tool bits; zirconium phosphates used for 

cation exchange applications, such as dialysis, drug delivery, 

and nuclear waste management; and zirconium sulfates used in 

leather tanning, as catalysts, and as coatings for titanium oxide 

pigment powder (Nielson and Wilfling, 2002).

Zirconium and hafnium metals are produced by the Kroll 

process, which involves reduction of zirconium oxychloride 

by magnesium metal in an inert atmosphere. The resulting 

metal contains a mixture of zirconium and up to 2 percent 

hafnium and is used for nonnuclear applications in corrosive 

environments and in specialty alloys. For nuclear-grade 

applications, the metals must be separated because of their 

different neutron absorption characteristics. Hafnium-free 

zirconium metal has a low thermal neutron-absorption cross 

section, so it is used for structural materials in nuclear reactors 

and as cladding for nuclear fuel rods. Hafnium has a relatively 

high thermal neutron absorption cross section and thus is 

used in nuclear control rods. Reactor-grade zirconium and 

hafnium are produced by dissolving zirconium oxychloride 

and chemically separating the hafnium from the solution. The 

zirconium and hafnium components are separately processed 

to form chlorides, which are then reduced to metals by the 

Kroll process. Hafnium is also used in high-temperature 

ceramics, nickel-base superalloys, and nozzles for plasma 

arc metal cutting.

Recent Aspects of Supply and Demand

Global production of zirconium mineral concentrates 

increased by more than 50 percent between 2003 and 2012 

(fig. V2A); the main source of the output was zircon from 

heavy-mineral sands. A relatively small quantity of zirconium 

was derived from baddeleyite produced from a single source on 

the Kola Peninsula of Murmanskaya Oblast, Russia. Excluding 

the United States, world production reached approximately 

1.68 million metric tons of zirconium mineral concentrates 

in 2011, followed by a significant decrease to 1.45 million 

metric tons in 2012 (fig. V2A; U.S. Geological Survey, 2014). 

In 2012, Australia and South Africa were the leading world 

zircon producers, accounting for approximately 70 percent 

of all production outside of the United States (fig. V2B; 

Bedinger, 2015a). China was the leading consuming country.

Global consumption of zircon increased in 2011, largely 

as a result of increased demand in developing economies in 

Asia and the Middle East (Gambogi, 2012b), and the prices of 

zirconium ores and concentrates rose dramatically during the 

same period. In 2011, prices of domestic standard-grade bulk 

zircon increased to between $2,550 and $2,750 per metric ton 
Introduction V3

A

F

Zirconium figure W1

G

C D E

B

Figure V1. Examples of sources and uses of zirconium and hafnium. A, Cube and crystal bars of highly pure zirconium 

metal (greater than 99.95 percent purity). The cube is 1 cubic centimeter in volume. B, (left) A pure (99.98 percent) melted 

tip of a hafnium electrode used in an electron beam remelting furnace for manufacturing metal parts; (right) an oxidized 

hafnium electron beam remelted ingot (color is from thin film effects in the oxide layer); and (center) a cubic centimeter 

of hafnium metal. Photographs A and B are by Heinrich Pniok (http://pse-mendelejew.de). C, Primary igneous zircon 

grains separated from granite. Note the prismatic character of most of the grains and the double termination on some. 

The wire (for scale) in the lower left-hand corner of the photograph is approximately 0.1 millimeter (mm) in diameter. 

D, Detrital zircon grains separated from metamorphosed sandstone. The rounding of the grains likely occurred during 

sedimentary transport and reworking. The wire (for scale) at the top of the photograph is approximately 0.1 mm in 

diameter. Photographs C and D are by James V. Jones. E, Baddeleyite crystals from the Palabora carbonatite deposit, 

Limpopo Province, South Africa. Photograph by Rob Lavinsky (www.iRocks.com). F, Heavy-mineral concentrations (dark 

seams) in weakly indurated outcrops of quartz beach sand from Chennai, India. The penny embedded in the sand (lower 

right) is for scale. Photograph by Mark Wilson (College of Wooster, Ohio). G, Photomicrograph of heavy-mineral sands 

from Sri Lanka that are rich in red to brown spinel. Note the larger, well-rounded pink zircon grain indicated by the white 

arrow. The width of the view is 20 mm. Photograph by Simm Sepp (www.sandatlas.org).
V4 Critical Mineral Resources of the United States—Zirconium and Hafnium

from between $830 and $890 per metric ton (fig. V2C; 

Loferski, 2013a). The average unit value of imported 

zirconium ore and concentrates (primarily zircon sand) 

was $2,170 per metric ton in 2011 (fig. V2C; Loferski, 

2013a). The subsequent decrease in world production of 

zirconium mineral concentrates in 2012 was in response 

to weakened demand, particularly in China, because of a 

slowdown in housing construction. Global consumption 

of zirconium mineral concentrates is expected to increase 

in the future, as demand from such consumers as the 

ceramics, chemicals, and metals industries increases, 

driven by renewed growth in developing economies. 

Construction and operation of new nuclear powerplants 

throughout the world are also expected to result in future 

demand increases for zirconium and hafnium metal. 

Although new zircon mines were brought online in recent 

years, the global financial situation led to decreased 

production and delayed the development of several 

mining projects. Thus, constricted supply together with 

the projected increase in demand could create a scarcity 

of zircon in the future (Loferski, 2013a). Despite this 

possibility, zirconium ranks relatively low on the British 

Geological Survey’s risk list, possibly reflecting the 

political and (or) economic stability of the countries that 

are major sources of this mineral commodity and the 

fact that production comes from several areas worldwide 

(British Geological Survey, 2012).

2003 2004 2005 2006 2007 2008 2009 2010 2011 2012

Production, in

million metric tons

A. World zirconium mineral concentrates mine production1

0

20

40

60

80

100

2007 2008 2009

Year

Year

2010 2011 2012

Other

Mozambique

India

Indonesia

China

South Africa

Australia

Zirconium-producing

 countries1

Percent of global production1

B. Zirconium mineral concentrates mine production by country1

0

25

50

75

100

125

0

25

50

75

100

C. Amount, value, and price of U.S. exports and imports of 

 zirconium ores and concentrates

Amount, in

thousand metric tons 

0

1

2

3

2003 2004 2005 2006 2007

Year

2008 2009 2010 2011

Value, in

thousand U.S. dollars Price per ton,2 in thousand U.S. dollars

Zirconium Figure W2.

Domestic, standard-grade, bulk

Australian standard-grade, 

free on board, bulk

U.S. exports

U.S. imports

U.S. imports

U.S. exports

1Does not include the United States

2Average price

1.0

1.5

2.0

Figure V2. Graphs showing global zirconium 

production (not including U.S. production) 

and U.S. trade information for zirconium 

ores and concentrates. A, World zirconium 

mine production from 2003 to 2012. Data are 

from Gambogi (2010), U.S. Geological Survey 

(2010–13), and Bedinger (2014). B, Global 

zirconium mine production, by country, from 

2007 to 2012. Data are from U.S. Geological 

Survey (2010–13) and Bedinger (2014). 

C, U.S. exports and imports of zirconium ores 

and concentrates, by amount, value, and price, 

from 2003 to 2011. Data are from U.S. Geological 

Survey (2006–9, 2010–13), Gambogi (2010), and 

Bedinger (2014).
Geology V5

Geology

Geochemistry and Mineralogy

Zirconium and hafnium are refractory lithophile 

elements that have nearly identical charge, ionic radii, and 

ionic potentials. They are classified as incompatible elements 

because they have physical and crystallochemical properties 

that exclude them from the crystal lattices of most rockforming minerals. In particular, zirconium and hafnium cations 

have relatively high charges (Zr4+ and Hf 4+) and small ionic 

radii. Thus, they and other similar elements, such as niobium, 

thorium, and uranium, develop intense electrostatic fields and 

do not substitute for the major elements in ordinary minerals. 

Instead, these elements, known as high field-strength elements 

(HFSE), concentrate in less common accessory minerals. 

Zirconium and hafnium are enriched in Earth’s crust relative 

to the mantle and have estimated average crustal abundances 

of 100 parts per million (ppm) and 3 ppm, respectively 

(Taylor and McLennan, 1995). Both elements are concentrated 

in the upper crust relative to the lower crust by a factor of 

approximately 1.5 (Taylor and McLennan, 1995).

Zircon (ZrSiO4) is the most common naturally occurring 

zirconium- and hafnium-bearing mineral. Most zircon forms as 

a product of primary crystallization in igneous rocks; in most 

rocks, zircon is the main mineral and zirconium is an essential 

structural constituent (Hoskin and Schaltegger, 2003). Most 

natural zircon has a Zr:Hf ratio of 33.6:1, which is about the 

same as the average crustal abundance ratio of those elements 

(Taylor and McLennan, 1995). The hafnium content of zircon 

can vary widely, however (from 6,000 to 11,600 ppm) (Ahrens 

and Erlank, 1969; Heaman and others, 1990), depending on the 

composition of the magma from which the zircon crystallized.

The natural color of zircon crystals varies from colorless 

to yellow, reddish-pink, purple, brown, blue, and green. 

Crystal sizes range in diameter and length from microns to a 

few centimeters, and well-formed zircon crystals are foursided prisms terminated by pyramids on each end (fig. V1C). 

Zircon exhibits a wide range of morphologies, however, 

which can provide clues about the conditions of formation or 

the subsequent history of the grain (Pupin, 1980; Corfu and 

others, 2003). In igneous rocks, zircon is generally present as 

small, early formed crystals enclosed in later-formed minerals. 

In some cases, such as granite pegmatites and nepheline 

syenites, zircon forms large, well-developed crystals. Once 

zircon has crystallized, its refractory nature allows it to persist 

through repeated cycles of igneous activity, and inherited 

zircon is common in a variety of igneous rocks, either as 

complete grains or as internal cores of younger zircon crystals. 

In metamorphic rocks, zircon can form by recrystallization 

of existing grains or by growth of new material. At lower 

temperature and pressure, new zircon is rare, and metamorphic 

zircon is formed by partial to complete recrystallization of 

existing grains. Under more intense metamorphic conditions, 

new zircon forms more readily at subsolidus conditions and 

also during partial melting (Schaltegger and others, 1999; 

Rubatto and others, 1999). Rare authigenic zircon has been 

reported in sedimentary rocks (Saxena, 1966), but sedimentary 

zircon is derived primarily from the weathering of preexisting 

igneous, sedimentary, or metamorphic rocks. Detrital zircon 

(fig. V1D) in sediment and sedimentary rocks is highly 

durable and chemically inert and, therefore, can survive 

numerous cycles of reworking and concentration in the 

sedimentary environment.

Baddeleyite (ZrO2 ) is a less common zirconium ore 

mineral that occurs in alkaline rocks, mafic and ultramafic 

rocks, and some metamorphosed carbonate rocks. It may 

occur with zircon and has been recognized as an accessory 

phase in a wide variety of rock types. Baddeleyite is most 

common in the late-stage, most chemically fractionated 

portions of mafic magmas (Heaman and LeCheminant, 1993), 

and it is the main zirconium-bearing phase in carbonatite 

rocks (fig. V1E) because of the generally lower silica activity 

in carbonatite melts. As with zircon, the Zr:Hf ratio in 

baddeleyite can vary considerably, and baddeleyite has been 

shown to fractionate zirconium from hafnium, particularly in 

carbonatite systems (Scharer and others, 1997; Klemme and 

Meyer, 2003). In pristine mafic rocks, baddeleyite crystals are 

usually light tan to dark brown in color and form euhedral, 

wafer-thin blades that commonly contain striated crystal faces 

(Heaman and LeCheminant, 1993). Crystals are typically very 

small, often less than 30 microns in width (Söderlund and 

Johansson, 2002).

Eudialyte (Na15Ca6Fe3Zr3Si26O73(OH)4Cl2) is a relatively 

rare zirconium-bearing mineral that is locally abundant in 

some types of potassium- and sodium-rich rocks (Kogarko, 

1990). It is an essential mineral in a variety of peralkaline 

nepheline syenites and nepheline syenite pegmatites, particularly those classified as agpaitic (Sørenson, 1997; Le Maitre, 

2002). Eudialyte-bearing rocks have been identified in 

multiple alkalic intrusive complexes worldwide (Marks and 

others, 2011); the Lovozero Massif on the Kola Peninsula in 

Murmanskaya Oblast, Russia, hosts nearly monomineralic 

eudialyte ores that contain 10 percent ZrO2 (Kogarko, 1990).

Hafnon ((Hf,Zr)SiO4 ) is a rare hafnium-rich mineral 

that exhibits continuous solid solution with zircon. Synthetic 

hafnon has been produced with 100 percent hafnium substitution (Speer and Cooper, 1982), but naturally occurring 

hafnon crystals typically have Zr:Hf ratios of 0.03 to 0.08. 

Hafnon occurs in tantalum-bearing granite pegmatites in the 

Zambézia district of Mozambique (Correia Neves and others, 

1974), and hafnian zircon has been identified in a niobiumtantalum-rich granitic complex in China (Wang and others, 

1996) and in several occurrences of granitic pegmatite worldwide (for example, Levinson and Borup, 1960; Černý and 

Siivola, 1980; Nickel and Robinson, 1985).

Zirconium and hafnium ore minerals are all primarily 

formed by crystallization from magma. The stability of 

zirconium- and hafnium-bearing phases in igneous systems is 

largely controlled by the concentration of the low-abundance 
V6 Critical Mineral Resources of the United States—Zirconium and Hafnium

elements (Watson and Harrison, 1984), so the element concentrations depend, in large part, on the magma source. Trace 

element patterns from basalts generated along mid-ocean 

ridges indicate that the upper mantle—the main source of melt 

generation beneath oceanic spreading centers—is relatively 

depleted in zirconium, hafnium, and other incompatible 

elements. Partial melting of the depleted upper mantle 

produces mainly tholeiitic or subalkaline basalts that have 

relatively low concentrations of zirconium and hafnium. In 

contrast, alkaline basalts formed in tectonic settings, such as 

oceanic islands and continental rifts, are relatively enriched in 

zirconium and hafnium because they are derived from deeper, 

more heterogeneous, undepleted, or locally enriched parts 

of the mantle (for example, Pearce and Norry, 1979; Fitton, 

1987; Weaver and others, 1987). Mantle enrichment occurs 

through metasomatism; that is, through chemical alteration 

that results from interactions between rocks and fluid. This 

process can replenish lithophile elements in the mantle prior to 

melt extraction; however, the nature of mantle metasomatism 

and its role in the generation of alkaline basalts is still under 

debate (Bailey, 1987; Menzies, 1987; Hoffman, 1997).

Zirconium and hafnium concentrations in igneous 

systems are also a function of the processes of melt generation 

and liquid evolution. Alkaline magmas are derived from 

enriched or less depleted sources, but they also are formed 

from relatively low fractions of partial melt (Thompson and 

others, 1984; McKenzie, 1985). Small fractions of partial melt 

preferentially extract alkalis and other incompatible elements, 

thus concentrating these elements even more relative to the 

magma source. As a result, zirconium concentration in igneous 

rock suites derived from alkaline basalts is significantly 

greater than igneous rock suites derived from tholeiitic or 

subalkaline basalts (Engel and others, 1965). Zirconium 

and hafnium are further concentrated during crystallization 

of alkaline systems, as these elements tend to remain in the 

residual liquid because of their higher solubility in magmas 

rich in sodium, potassium, fluorine, chlorine, and hydroxide. 

In the Skaergaard intrusion of East Greenland, late-stage 

differentiates of the layered mafic-ultramafic complex have 

unusually strong enrichment of zirconium; that is, greater than 

1,000 ppm. The enrichment is particularly significant because 

the initial magma concentration was 94 ppm, which is similar 

to the average zirconium content of tholeiitic or subalkaline 

basalts (Wager and Mitchell, 1951; Brooks, 1969).

Zirconium, hafnium, and other HFSEs are generally 

depleted in subduction zone magmas relative to mid-oceanridge, ocean-island, and rift-related basalts (Perfit and others, 

1980). Rubatto and Hermann (2003) demonstrated that zircon 

is stable as a residual phase in subducted basalt and sediment 

and thus could limit the liberation of HFSEs into subduction 

zone fluids and melts. Melts generated in subduction zones 

are relatively rich in water, and the solubility of zirconium, 

hafnium, and other HFSEs in aqueous solutions is much lower 

than that for other incompatible elements, such as potassium 

(Gill, 1981, and references therein). As a result, some incompatible elements are scavenged and relatively enriched in 

subduction zone magmas, whereas others are more immobile 

and remain in the country rocks. As hydrous melts migrate 

through the mantle wedge, interaction with the depleted upper 

mantle leads to preferential HFSE depletion in the ascending 

magma (Kelemen and others, 1980). Less-depleted magmas 

in subduction zone systems may be produced if the melts 

interact with enriched pockets of lithospheric mantle beneath 

the crust (Edgar, 1987) or with metasomatized regions of the 

continental or oceanic crust. These effects are relatively localized, however, and the resulting concentrations of zirconium, 

hafnium, and associated HFSEs are still relatively low.

Because zirconium and hafnium are nearly identical in 

chemical properties, the geochemical behavior of hafnium 

during magma formation and differentiation is generally 

similar to that of zirconium. This is particularly true in mafic 

to intermediate magmas, which represent relatively pure 

silicate melts and have Zr:Hf ratios similar to the chondritic 

model of bulk silicate Earth composition (Bau, 1996; 

David and others, 2000). In other magmas, particularly highly 

evolved melts rich in boron, chlorine, fluorine, lithium, phosphorus, or water, Zr:Hf ratios vary substantially and indicate 

decoupled fractionation of the two elements (Ellison and 

Hess, 1986). The general behavior of trace elements in highly 

evolved melts rich in volatiles is difficult to model quantitatively (Bau, 1996), and yet these are the kinds of magmas in 

which zirconium and hafnium are naturally concentrated. As 

a result, the hafnium content of the resulting ore minerals and 

(or) the presence of hafnon in igneous systems are variable 

and difficult to predict. In the Skaergaard intrusion, differentiation produced Zr:Hf ratios that range from 59 in the most 

mafic portions to 85 in the most silica- and alkali-rich parts of 

the mafic-ultramafic complex (Brooks, 1970). In calc-alkaline 

suites, the Zr:Hf ratios of zircon range from 50 to 70 in 

gabbro to 30 to 35 in granite and suggest hafnium enrichment 

during magma evolution (Condie and Lo, 1971). Zircon from 

silica-undersaturated rocks, such as syenite and nepheline 

syenite, has low hafnium content and high Zr:Hf ratios relative 

to granite, whereas peralkaline granite, granitic pegmatite, 

and lithium-tantalum-bearing granite and pegmatite are the 

only rocks known to contain both hafnon (Correia Neves 

and others, 1974) and also zircon with high concentrations 

of hafnium (>30 percent) (von Knorring and Hornung, 1961; 

Wang and others, 1996) and also unusually low Zr:Hf ratios 

(<20) (Wang and others, 1996).

Deposit Types

Primary Igneous Deposits

The world’s largest primary deposits of zirconium and 

hafnium are associated with alkaline igneous rocks (Kogarko, 

1990). These rocks also host much of the niobium, tantalum, 

and rare-earth elements (REEs) on Earth (Kogarko, 1990; 

Chakhmouradian and Wall, 2012), and they have a strong 

association with economic deposits of apatite (Kogarko, 1987; 

Singer, 1992) and diamond (Bergman, 1987; Dawson, 1987; 
Geology V7

Cox, 1992). Volumetrically, alkaline rocks account for less 

than 1 percent of all igneous rocks, however, and individual 

intrusions are typically of limited spatial extent and volume. 

Also, the concentration of zirconium and hafnium in these 

enriched igneous systems is still relatively low. Thus, only 

a few primary igneous deposits of zirconium- and hafniumbearing minerals are economic to mine at present.

The Kovdor deposit lies within the Paleozoic Kola 

alkaline province, which is located on the southwestern 

part of the Kola Peninsula in Murmanskaya Oblast, Russia 

(fig. V3). This deposit, from which baddeleyite is recovered 

as a byproduct of apatite and magnetite mining, is the only 

igneous deposit in the world that is currently producing 

zirconium (Singer, 1992). The Kovdor deposit is made up 

primarily of carbonatite and phoscorite; the carbonatites are 

represented by both calcite and dolomite carbonatite varieties. 

Phoscorites have carbonate mineral assemblages similar 

to those in the carbonatites but also include magnetite and 

silicate minerals, such as apatite, forsterite, and phlogopite 

(Rodionov and others, 2012). The ore complex (which covers 

a 0.8-by-1.3-kilometer [km] area) is part of a 40-squarekilometer (km2

) multiphase, concentric intrusion that also 

includes ultramafic rocks (clinopyroxenite and dunite) and 

nepheline syenite (Verhulst and others, 2000, and references 

therein). The intrusion extends down to 12 km with steeply 

dipping to vertical contacts (Arzamastsev and others, 2000), 

and it is surrounded by Archean biotite gneiss and granitic 

gneiss. The Kovdor intrusive complex was emplaced along 

the Kandalaksha Deep Fracture Zone, which is one of the 

most prominent structural features of the eastern Baltic Shield 

(Bell and others, 1996).

The broader Kola alkaline province consists of more than 

24 intrusive complexes of Devonian age (380 to 360 megaannum [Ma]) (Kramm and others, 1993; Rodionov and others, 

2012) that are exposed across more than 100,000 km2 from 

western Finland to the eastern part of the Kola Peninsula. 

Petrogenetic models for the Kola alkaline province suggest 

melt derivation from lithospheric mantle that was metasomatized prior to melting, and that the trigger for melting 

might have been a mantle plume (Sindern and others, 2004; 

Downes and others, 2005). Melt generation, migration, and 

emplacement were controlled by rifting perpendicular to 

and along regional northwest-southeast-trending structures 

with Precambrian ancestry (Dudkin and Mitrofanov, 1994). 

Associated intrusive complexes—namely the Lovozero 

and the Khibiny complexes––represent the largest agpaitic 

intrusions and contain the largest magmatic apatite deposits 

in the world (Kogarko, 1990). These intrusions also have 

localized phases that are enriched in zirconium, such as veins 

of pegmatitic nepheline syenite in the Khibiny complex that 

contain eudialyte. The Lovozero complex contains more widespread eudialyte-bearing nepheline syenite, called lujavrite, 

as well as ore horizons that contain up to 10 percent ZrO2

and between 50 and 90 percent eudialyte (Kogarko, 1990). 

Locally, lujavrite forms plate-like bodies up to 800 meters (m) 

thick. Despite the enrichment of eudialyte in both of these 

intrusions, neither Lovozero nor Khibiny is presently being 

mined for zirconium.

Secondary Deposits—Coastal Placers 

and Paleoplacers

Zirconium and hafnium ore minerals, particularly zircon, 

are present in a wide variety of rocks, but rarely in economic 

concentrations. Thus, most economic deposits are formed 

in sedimentary environments wherein zircon and other ore 

minerals are separated from their primary host rocks by 

weathering and erosion and transported and concentrated 

in surficial deposits. Zircon is chemically stable and hard 

(has a hardness of 7 on the Mohs scale) and is consequently 

quite durable in the sedimentary environment (Morton and 

Hallsworth, 1999). Zircon and other related ore minerals 

also have high densities and therefore accumulate under 

the influence of gravity in wind and water. The resulting 

heavy-mineral deposits, called placers, exhibit a wide range of 

physical characteristics, form in many different environments, 

and host a variety of minerals; for example, they are valuable 

resources for gold, iron, tin, and titanium (Slingerland and 

Smith, 1986; Garnett and Bassett, 2005). Alluvial placers 

form in rivers and streams, and they contain concentrations 

of zircon and other heavy minerals derived from source 

rocks within a drainage network. Additional concentration 

can take place as sediment moves downstream and as older 

river deposits are reworked by younger systems, although 

downstream dilution is more common.

In sediment deposited along the coasts, heavy-mineral 

enrichment takes place primarily in the foreshore and 

uppermost part of the shoreface environment (fig. V4D) where 

sediment is repeatedly reworked by wind, waves, and waveinduced currents. Coastal placer and paleoplacer systems, or 

beach placers, form along wave-dominated coastlines where 

reworking is most intense. They supply all of the world’s 

current primary production of zircon and constitute the most 

important deposit type for zirconium and hafnium (Garnett 

and Bassett, 2005). Rigorously defined, a coastal placer is a 

concentration of heavy minerals within sediment deposited in 

a range of coastal sedimentary environments that are typically 

dominated by eolian, wave, and tidal processes (Hamilton, 

1995). The concentrating agent at the site of deposition is 

primarily mechanical, and the presence and concentration of 

zircon and other economically important minerals, such as 

ilmenite and rutile (for titanium), chromite (for chromium), 

and monazite (for REEs) are dependent on a variety of factors 

described below. Most deposits range in age from the Eocene 

Epoch to the present, and economically viable deposits occur 

along numerous modern beaches on nearly every continent 

(fig. V3; for example, Peterson and others, 1986; Mallik and 

others, 1987; Hamilton, 1995; MacDonald and Rozendaal, 

1995; Hou and others, 2011). Rare examples of much older 

paleoplacers, such as the Archean Witwatersrand deposit 

in South Africa, have been successfully mined, although 

primarily for gold (Minter, 2006).
V8 Critical Mineral Resources of the United States—Zirconium and Hafnium 180° 135° 90° 45° 0° 45° 90° 135° 180° 60° 50° 40° 30° 20° 10°0° 10° 20° 30°

40°50°

80°

70°

70°

80°

60°

Kovdor

Trail Ridge

Stony Creek

Seven Devils

Coos Bay

Eucla basin

Jacinth-Ambrosia Murray basin

Douglas

Kilimiraka Lethbridge

South

Kwale

Moma

Grande Côte

Namakwa Richards Bay

Sierra Rutile

Srikurmam

Pulmoddai

Ranobe

Guaju

Buena

Odisha

Athabasca

Coburn

Eneabba

Tutnup South

Perth basin

Keysbrook

Mission

Zirconium figure W3.

Base from U.S. Geological Survey Global 30 arc-second elevation data (1996) and from Natural Earth (2014); Robinson projection; World Geodetic System 1984 datum

EXPLANATION

Zirconium and hafnium deposits

IgneousCoastal placer or paleoplacerOil sands

Modern coastal placer regions

rom fig. 2 of World map showing locations of selected zirconium and hafnium deposits and regions with modern coastal placer systems (adapted f Figure V3.

Hamilton, 1995).
Geology V9

Coastal placers are typically fed by detritus derived 

from deeply weathered bedrock source terranes that include 

abundant felsic to intermediate crystalline rocks (fig. V4A). 

Host sands form coast-parallel surficial bodies, such as 

barrier islands (figs. V4B and V4C), and older sand bodies 

form topographic highs inland of the modern coastline that 

represent former barrier islands or eolian dunes (fig. V4A; 

Force and Rich, 1989). Sand bodies are formed in the 

foreshore, shoreface, dune, tidal delta, and washover environments (figs. V4B and V4D; Roy and others, 1994); the 

economic heavy-mineral concentrations within these systems 

are typically about 10 m thick, 1 km wide, and more than 

5 km long. In some cases, economic deposits are made up 

of amalgamated smaller sand bodies of similar origins or 

superimposed composite bodies from different environments 

(Force, 1991). The mineral grains in most economic deposits 

are medium- to fine-grained sand-sized, well sorted, generally 

well rounded, and typically unindurated (fig. V1G). Welldeveloped weathering profiles are superimposed on most sand 

bodies, and local cementation by iron oxides, clay minerals, 

or humate is common. The proportions of heavy minerals vary 

considerably both within deposits and among sand bodies 

(fig. V1F ); the average concentration of total heavy minerals 

in economic deposits ranges from less than 1 percent to about 

25 percent. Zircon makes up a relatively small percentage of 

the economic heavy minerals in most deposits; thus, zircon is 

primarily a byproduct of heavy-mineral sand mining for the 

titanium minerals ilmenite and rutile.

In a review of some of the world’s major coastal 

titanium- and zirconium-placer-forming regions, Hamilton 

(1995) noted the following three major factors that were 

common to all of them: (a) placer-forming regions lie on a 

trailing continental margin (Inman and Nordstrom, 1971), 

(b) margins are bounded by escarpments, and (c) heavy 

minerals were derived from Precambrian metamorphic 

bedrock regions and intermediate sedimentary hosts. 

Otherwise, individual deposits vary considerably in terms 

of their structure, location, and conditions of formation. 

Hamilton (1995) examined the controls on placer formation 

that act at different spatial and temporal scales and developed 

a conceptual model focused on the genetic characteristics of 

the placer-forming regions. The conceptual model is shown as 

a flow chart in figure V5A, with broader controls, such as the 

tectonic setting and bedrock geology of a region, at the top and 

more localized controls, such as coastal plain morphology and 

coastal physiography, at the bottom. The shading and hatched 

patterns indicate the time scales at which the different controls 

occur, and the arrows indicate potential interactions between 

different controls within the placer-forming system. The rest 

of this section discusses some of these key controls and how 

they influence placer formation.

The most important long-term control on placer formation is tectonic setting. In general, major placer-forming 

regions form on passive margins, as illustrated by modern 

coastal placer regions on passive tectonic margins, such as the 

eastern coast of North America and South America and the 

western and southeastern coasts of Africa (fig. V3). Coastal 

placers can also form along both convergent and transform 

margins, as evidenced by placers along the Oregon coast 

(Coos Bay, Seven Devils Mine) (Peterson and others, 1986) 

and both the main islands of New Zealand (Brathwaite and 

Christie, 2006). The size of these deposits is small compared 

with those in some other regions, such as southern Australia 

and South Africa, however. Most major placer-forming 

regions overlie broad coastal plains of Mesozoic sediments, 

and all have an escarpment that separates the low-lying coastal 

plain from the hinterland, or the broad inland region behind 

the coastal plain that includes the piedmont and more distal 

regions of basement bedrock outcrops (fig. V4A). Escarpments 

are relatively dynamic features that act over intermediate 

time scales, and they are important because they (a) provide a 

mechanism for initiating erosion of bedrock through headwall 

retreat, (b) provide a mechanism to rejuvenate fluvial landscapes and evacuate stream valleys during periods of low sea 

level, thus providing sediment to the outer edge of the coastal 

plain for subsequent reworking and enrichment, (c) provide a 

distinct landward barrier to marine transgression that restricts 

the location of repeated reworking of coastal sediments during 

periods of rapid sea-level change, and (d) indicate tectonic 

instability or flexure along the margin that can elevate and 

preserve heavy-mineral concentrations near the escarpment 

(Hamilton, 1995).

The lithology of the hinterland region is the most 

variable aspect of the tectonic setting of major placer-forming 

regions. Most placers are adjacent to provinces underlain 

by Precambrian crystalline bedrock, whereas others were 

derived from basaltic terranes or Phanerozoic orogenic belts. 

Thus, Hamilton (1995) points out that, although lithology 

is an important control on the presence and abundance of 

economic minerals, there are other key influences on the 

degree of concentration and, therefore, the economic potential 

of placers and paleoplacers. The weathering history of the 

source region is perhaps most important in determining the 

likelihood of concentration of a suitable mineral suite. The 

most enriched modern placer systems are located at latitudes 

lower than 35° (fig. V3; Force, 1991). Climatic change during 

the Cenozoic Era played an important role in the development 

of paleoplacer deposits worldwide directly through more 

intense weathering and, more indirectly, through repeated 

cycles of sea-level change (Force, 1991). Many of the major 

placer coastlines of the world underwent a shift from arid to 

tropical conditions during the Quaternary Period. This longer 

term trend, coupled with more rapid and repeated shifts in 

climate, produced large volumes of regolith in source areas, 

preferentially removed unstable minerals from existing 

coastal placers, and led to alternating periods of high and low 

sediment delivery to the coast. Repeated sea-level change 

associated with glacial-interglacial cycles caused multiple 

marine excursions across the continental margins, leading to 

extensive sediment reworking, heavy-mineral enrichment, 

and preservation of strandlines associated with sea-level 

highstands (fig. V4A). These processes were most significant 
V10 Critical Mineral Resources of the United States—Zirconium and Hafnium

Piedmont

Escarpment

Crystalline bedrock

containing zircon Saprolite and heavily

weathered

crystalline rocks

Active barrier 

island systems

A

B

Modified from Reinson (1979)

Older strand lines, beach ridges,

and barrier islands elevated by

active faulting, marine regression,

and (or) isostatic adjustments

Crystalline basement

NOT TO SCALE

Coastal plain sediments

Coastal

plain

Modified from Markewich and others (1990)

Modified from Reinson (1979)

Ocean

River

River

NOT TO SCALE

Marsh

Tidal flats

Washover Dunes

Beach

Dune

Berm

Beachface Low tide

terrace

Longshore 

trough

Longshore

bar

Swash

zone

Surf

zone

Breaker

zone

Shoaling

zone

Wave base

10

5

Sea

level

–5

–10

Backshore—Dune Fore- Shoreface

shore

Lagoon

Shoreface

Barrier

island

Main tidal channel

Flood

tidal

delta

Secondary

tidal 

channel 

Ebb tidal

delta

METERS

High-water level

Low-water level

Littoral drift and

dominant sediment

transport direction

Washover

Accretional spit

Pro-grading

beach ridges

River

input

Swash

bar

Sediment

transport

reversal

Dominant

wave

angle

C

D

NOT TO SCALE Modified from Elko and Davis (2006)

NOT TO SCALE

Zirconium Figure W4

Figure V4. Examples of coastal depositional systems with an emphasis on barrier island location, morphology, and depositional 

environments. A, Schematic model for coastal placer systems. Zircon is derived from deeply weathered crystalline bedrock, delivered 

to the coast by river systems, and concentrated by a variety of factors, including wind and waves, storm events, tidal action, and ocean 

currents. Zircon may also be derived from older coastal plain sediment or offshore sediment. B, Depositional environments associated 

with barrier island systems. The highest concentration of zircon and other heavy minerals are found primarily in the beach, dune, and 

washover environments. C, Process-response model and features of a drumstick barrier island system along coasts, such as that 

of the Southeastern United States, that have a consistent supply of sediment from major rivers. Sediment is redistributed from river 

input in the direction of littoral drift (bold arrows), and heavy-mineral concentration occurs in the beach and washover environments. 

D, Generalized profile of barrier beach and shoreface environments.
Geology V11

4 Disseminated heavy minerals

Transgressive barrier

Prograded barrier

Seabed

Sea level

Dune

Erosional

discontinuity

Bedrock or

coastal plain sediment

2 1 3

Modified from Roy (1999)

Modified from Roy and others (2000) Modified from Roy and others (2000)

NOT TO SCALE

NOT TO SCALE

x

x

x x

x

x

x x x

x x x x x x xx x x x x x x x x x x x x x x x

Pleistocene

strand plain Bedrock headlands

Holocene

beach

Sea bed

Erosion surface

Small Holocene barrier

Heavy-mineral concentration

Littoral drift

EXPLANATION

Littoral drift

EXPLANATION

B C

D

Seabed 

erosion

Lagoon

Coastal plain

Rising

sea level

Net sand

loss

Washover

Tectonic setting

Wave and

tide energy

Coastal plain

morphology

Geology Relative

sea level

Geologic

structure

Continental shelf

morphology

Coastal

physiography

Placer formation

A component of

Modified from figure 8 of Hamilton (1995)

A

Seabed

Relict land surface

Long-term

Short-term

Quasiinstantaneous

Time scales of variability

Transgressive barrier

(bold gray arrow indicates

washover direction)

Sea level

Heavy-mineral enrichment

Cross-section of barrier and strand plain

Estuary between

older (landward)

and younger

(seaward) barriers

Older barriers

Younger barriers

Heavy-mineral

concentrations

Disseminated heavy minerals

Lagoon

deposits

Heavy-mineral

concentrations

Zirconium figure W5.

EXPLANATION

Figure V5. Influences on and locations of placer 

formation in coastal environments. A, Conceptual 

model showing influences on placer formation in 

coastal environments. Influences of the different 

components of the coastal system are indicated by 

arrows, and relative time scales for variability of select 

components are indicated by shading and diagonal 

lines. B and C, Models for enrichment of heavy minerals 

in the coastal environment by (B) littoral drift and 

(C) washover in transgressive barriers. D, Schematic 

cross-section through an idealized barrier island 

complex showing four different types of heavy-mineral 

deposits and their association with host barrier 

environments: (1) lag deposits, which are formed 

along erosional discontinuities or surfaces between 

barrier islands of different ages; (2) transgressive 

deposits, which are formed at the rear of barrier islands 

immediately landward of the strand plain during periods 

of rising sea level; (3) regressive deposits, which are 

formed in barrier islands that have a consistent supply 

of sediment during periods of falling sea level; and 

(4) eolian dune deposits with relatively low-grade 

disseminated heavy-mineral deposits formed downwind 

from heavy-mineral-rich beaches.
V12 Critical Mineral Resources of the United States—Zirconium and Hafnium

for placers lying near the coast during the Quaternary Period, 

but similar processes have also been cited as important in 

Eocene to Pliocene deposits as well (Puffer and Cousminer, 

1982; Carpenter and Carpenter, 1991; Hou and others, 2011).

Factors that vary over short-term, quasi-instantaneous 

time periods include tides, storms, winds, and waves. 

Microtidal coasts—that is, those with tidal ranges of less 

than 2 m, such as those in southwestern Australia—have 

long been considered to be the most favorable for coastal 

placer development (Sutherland, 1985). In these systems, 

placers of beach (foreshore environment) and eolian 

(backshore or dune environment) origin are common 

(figs. V4B and V4D; for example, Hou and others, 2011). 

In mesotidal (2 to 4 m) to macrotidal (>4 m) systems, such 

as those of the Southeastern United States and eastern 

India, placers are more likely to be found in finer grained 

progradational beach-ridge systems associated with river 

mouths (fig. V4C; for example, Pirkle and others, 1984). 

Storms, winds, and waves can have more variable effects 

on coastal morphology and placer development. Storm 

events can significantly modify coastal geomorphology, 

and storm surges can erode offshore sand deposits along 

the continental shelf and produce substantial heavy-mineral 

enrichment in the shoreface and washover environments 

(figs. V4B and V4D; Leatherman and others, 1977; Roy, 

1999; Roy and others, 2000; Buynevich and others, 2004). 

Strong onshore winds winnow shoreface deposits at low tide 

and ultimately transport sediment inland from the beach or 

foreshore environment, trapping deposits in the backshore 

environment well above sea level (figs. V4D and V5D; 

Roy, 1999). Economically important dune deposits formed 

by strong onshore winds are found in all regions except 

South Africa, where longshore winds are more significant 

for placer formation (MacDonald and Rozendaal, 1995). 

Trail Ridge, which is a prominent physiographic feature that 

extends for more than 160 km across Florida and Georgia, 

formed as a coast-parallel dune complex during a major 

Pleistocene sea-level highstand and has been an important 

domestic source of titanium minerals and zircon production 

since 1949 (fig. V3; Force and Rich, 1989).

Waves and ocean currents are perhaps the most 

important agents in producing economic heavy-mineral sand 

deposits. In areas such as southeastern Australia, where 

sands are not derived directly from rivers, waves drive large 

sand fluxes onshore and in longshore transport currents 

that move the sand along the coast from south to north in 

a process known as littoral drift (figs. V5B, and V5C; Roy, 

1999). Prominent bedrock headlands along the drift-aligned 

coastlines trap heavy minerals on the up-drift side of embayments, and enrichment takes place along erosional surfaces 

between different-aged barriers (fig. V5B; Roy, 1999). 

Similar processes have been described in beach placers 

along the Oregon coast, where reversing seasonal currents 

serve to both winnow and subsequently preserve heavymineral sands in pockets adjacent to the headland (Komar 

and Wang, 1984; Peterson and others, 1986, 1987).

In areas with more consistent sediment supply from 

external sources, such as inland river systems or scouring 

of the sea bed, heavy minerals are concentrated in sand-rich 

barrier-island systems (figs. V4A and V4B). Reworking and 

concentration takes place primarily in the upper shoreface 

environment in the beach and washover environments, 

and heavy-mineral-rich sands are distributed and further 

fractionated during transport along the length of the barriers 

by littoral drift (figs. V4C and V5C; Roy, 1999). Many 

economic deposits in the Florida and Georgia coastal plain 

of the Southeastern United States occur where Pleistocene 

and Holocene beach ridges intersect ancestral to modern river 

systems, thus illustrating the importance of rivers as a source 

of heavy minerals in this particular coastal system (Hails 

and Hoyt, 1972; Pirkle and others, 1974, 1984). The deposits 

extend in the direction of littoral drift (south) for kilometers 

from the river intersection and help to define a series of long 

linear ridges that represent relict shorelines formed during 

a succession of marine transgressions from the Pleistocene 

Epoch to the present (Winker and Howard, 1977). Economic 

deposits occur locally in modern barrier islands adjacent to 

river mouths along the Florida and Georgia coasts, and some 

of these have been mined in the past (Pirkle and others, 2007).

Other factors may also exert local control over placer 

formation and distribution. These include tectonic effects, 

such as active faulting (Peterson and others, 1987; Roy and 

others, 2000), flexure (Cronin and others, 1984), or isostatic 

uplift (Opdyke and others, 1984) that might increase sediment 

supply, promote reworking, or enhance the potential for 

preservation. These effects, however, must be analyzed in the 

context of the other variables described above to adequately 

assess the prospectivity of a particular deposit or region.

Resources and Production

Identified Resources

World zircon reserves are estimated to be 78 million metric 

tons of contained zirconium oxide (ZrO2

) (Bedinger, 2015b). 

Domestic zirconium reserves are estimated to be 500,000 metric 

tons of ZrO2 , and estimated resources include approximately 

14 million metric tons of zircon associated with titanium 

resources in heavy-mineral sand deposits (Bedinger, 2015b).

Zircon production in the United States is a byproduct 

of mining and processing heavy-mineral sands for titanium 

minerals in Florida and Virginia. U.S. companies usually 

consider data on domestic production and consumption of 

zircon concentrates to be proprietary and generally withhold 

these data. U.S. exports of zirconium ore and concentrates 

generally declined from 2005 to 2011 (fig. V2C), and the 

United States was a net importer of zirconium ore and concentrate in 2011 for the first time since 1998 (Gambogi, 2012a). 

The United States is one of the leading world producers of 

zirconium and hafnium metal, along with France and Russia 

(Gambogi, 2012a).
Exploration for New Deposits V13

New heavy-mineral sand mines have been brought 

online in Kenya, Senegal, and South Africa in recent years. 

The Kwale project in Kenya began production in 2014 and 

was expected to produce 30,000 metric tons per year over 

a 60-year mine life. The Tormin project in South Africa 

was expected to produce 48,000 metric tons per year of 

nonmagnetic concentrate grading 81 percent zircon and 

11.6 percent rutile over a 4-year mine life. In Senegal, the 

Grand Côte project was expected to produce 85,000 metric 

tons per year during a mine life of 20 years. In the 

United States, a new zircon mine started up in southeastern 

Georgia in 2014, and a second mine was expected to begin 

production in late 2015 (Bedinger, 2015b).

Other Resources

Because of its relative ubiquity in crustal rocks and its 

high specific gravity, zircon is likely present in most if not all 

placer deposits, but it is not always recovered during placer 

mining. Placers and paleoplacers have been mined since 

humans first used metals (Garnett and Bassett, 2005), but 

the first economic extraction of zircon took place only in the 

first half of the 20th century (Elsner, 2010). Global zircon 

production has historically been a byproduct of titanium 

mineral extraction, and, thus, potentially significant economic 

concentrations of zircon might remain in tailings processed 

exclusively for titanium and other commodities, such as gold, 

iron, REEs, and tin. In KwaZulu, Natal, South Africa, a mine 

tailings treatment plant was commissioned at Richards Bay 

(fig. V3) to recover heavy-mineral concentrates, including 

zircon, from approximately 30 years of accumulated mine 

tailings initially processed primarily for titanium minerals 

(Gambogi, 2012a). Similar treatment plants are in development to recover titanium minerals and zircon from tailings 

of the Athabasca oil sands project in the Province of Alberta, 

Canada (fig. V3). Depending on future price fluctuations, 

economic concentrations of zirconium and hafnium ore 

minerals might also be recovered from tailings of primary 

igneous deposits, particularly in alkaline igneous provinces, 

such as in central Arkansas (Flohr and Ross, 1990). These two 

examples highlight the potential for recoverable resources 

from tailings in a wide variety of settings where zircon is 

not or was not the primary commodity of interest in the past 

but may be present at economic concentrations and is also 

relatively easy to extract.

In addition, older, lithified sedimentary units sourced 

from more fertile bedrock regions (Dickinson, 2008) could 

yield economic concentrations of zircon, especially in 

areas where the bedrock was deeply weathered. One such 

example is the Cretaceous Tokio Formation that is exposed 

in southwestern Arkansas, which contains ilmenite-bearing 

sands derived in part from deeply weathered alkaline intrusions in the surrounding area (Hanson, 1997). Zircon could 

be recovered during quarrying operations for sand, gravel, 

and aggregate in environments or regions where relatively 

fertile bedrock might be exposed (for example, continental 

rift systems or uplifts of crystalline bedrock) or where surficial 

heavy-mineral concentrations might exist (for example, river 

valleys, dunes, or high-energy coastal deposits).

Exploration for New Deposits

The process of identifying and evaluating heavy-mineral 

sand deposits requires careful consideration and assessment of 

the many variables outlined above. Exploration of modern to 

recent coastal placer systems is aided by the relative ease of 

identifying the coastal morphology and physiography, sediment supply and distribution systems, geologic setting, and 

climate history. Identifying and understanding the distribution, 

geometry, and evolution of paleoshorelines is more difficult 

because of the complex interactions between medium- to 

short-term factors and increased potential for reworking and 

burial. One notable example––the discovery and development 

of the Ambrosia and Jacinth heavy-mineral sand deposits in 

the Eucla basin of southern Australia (fig. V3; Hou and others, 

2011)––provides a possible template for exploration. The 

Eucla basin initially was recognized as an extensive area of 

preserved Cenozoic marine sediment spanning approximately 

50 million years (m.y.) of deposition with a high degree of 

preservation and high prospectivity for heavy-mineral deposits 

(Clarke and others, 2003, and references therein); however, 

early exploration revealed a need to better understand the 

geometry and evolution of the paleoshorelines to provide 

a focus for exploration. The reconstruction of complex 

sequences of coastal deposition constrained by a detailed 

chronology of major sea-level events (McGowran and 

others, 1997) and models for heavy-mineral concentration 

(Roy, 1999; Roy and others, 2000) led to successful prediction 

of new prospective sites and key discoveries (Hou and others, 

2011). Mining at the Jacinth deposit was commissioned in 

2009, and more than 12 separate heavy-mineral prospects 

where zircon is the dominant component of the heavy-mineral 

assemblage have been identified (Hou and others, 2011). The 

Eucla basin represents an important new region of heavymineral sand resources; deposits in the region are capable of 

producing approximately 300,000 metric tons of zircon in 

their initial years of production.

Other regions with long records of Cenozoic coastal 

deposition remain the most prospective worldwide. Similar 

to the Eucla basin in Australia, the Gulf of Mexico Basin 

in the southern United States records more than 60 m.y. 

of sediment deposition and coastline evolution. Sufficient 

outcrop, seismic, and borehole data have been collected to 

describe the extent of various sedimentary facies associated 

with numerous depositional episodes in the Gulf of Mexico 

Basin. The basin includes volumetrically significant Paleocene 

through Miocene coastal plain and shore zone facies deposits 

(Galloway and others, 2000). The Gulf of Mexico Basin has 

been extensively evaluated for petroleum resources, but units 

representing paleoshorelines have not been thoroughly studied 

for the occurrence of heavy-mineral resources.
V14 Critical Mineral Resources of the United States—Zirconium and Hafnium

Environmental Considerations

Sources and Fate in the Environment

The behaviors of zirconium and hafnium in the environment are very similar to one another—most notably, that 

zirconium- and hafnium-bearing minerals have limited 

solubility and reactivity in the environment. More literature is 

available on the behavior of zirconium than of hafnium, but 

the studies on zirconium are generally applicable to hafnium 

for the following reasons: the elements are extremely similar 

chemically (that is, they have almost identical atomic sizes 

and analogous electronic structures); they nearly always occur 

together; hafnium occurs in solid solution with zirconium 

in zircon; and hafnium is recovered as a byproduct during 

zirconium extraction. Anthropogenic sources of zirconium, 

and likely those of hafnium, are from industrial zirconiumcontaining byproducts and emissions from the processing of 

sponge zirconium; exposure to the general population from 

these elements is generally considered small (Schaller, 2004).

The concentrations of zirconium in soils, waters, and the 

air are significantly higher than those of hafnium because of 

the higher crustal abundance of zirconium. The mean concentration in soils worldwide is 300 ppm (or 300 milligrams per 

kilogram [mg/kg]; the range is from 30 to 2,500 ppm) for 

zirconium compared with a mean concentration of 3 ppm 

(and a range of from 2 to 20 ppm) for hafnium (KabataPendias and Mukherjee, 2007). According to Shacklette and 

Boerngen (1984), soils in the conterminous United States have 

zirconium concentrations that average 230 ppm and range 

from less than 20 to 2,000 ppm; hafnium concentrations were 

not reported (Shacklette and Boerngen, 1984). According to 

Govindaraju (1994), reference soils from the United States 

contain between 2.5 to 17 ppm hafnium, and average 7.3 ppm. 

The ratio of zirconium to hafnium concentrations in waters is 

usually similar to their ratio in soils, but their concentrations 

are commonly very low because they are relatively immobile 

compared with some other elements (Gaillardet and others, 

2003). Concentrations of less than 1 part per billion (ppb) (or 

1 microgram per liter) for zirconium are common for natural 

waters, and concentrations for hafnium are typically less by 

an order of magnitude. According to Reimann and de Caritat 

(1998), the median dissolved and total concentrations in 

stream waters worldwide range from 0.01 to 0.9 ppb for zirconium and from 0.009 to 0.03 ppb for hafnium. Gaillardet and 

others (2003) reported average global dissolved concentrations 

in rivers of 0.04 ppb for zirconium and 0.006 ppb for hafnium. 

Industrial emissions likely do not increase these concentrations significantly (Schaller, 2004). Zirconium and hafnium 

are found in higher proportions in sediment in rivers than in 

the dissolved fraction. Globally, the average concentration in 

suspended sediments in rivers is 160 ppm for zirconium and 

4.4 ppm for hafnium (Viers and others, 2009). In the atmosphere, zirconium concentrations for industrial regions range 

from 0.2 to 7 parts per trillion (ppt) (or 0.7 to 26 nanograms 

per cubic meter [ng/m3

]) for zirconium and reach more than 

0.09 ppt (0.32 ng/m3

) for hafnium (Reimann and de Caritat, 

1998; Kabata-Pendias and Mukherjee, 2007).

Little reported information is available on the pre-mining 

geochemical signatures of soils, sediment, and waters around 

heavy-mineral sand deposits and around the carbonatite 

deposit in Russia that is currently being mined for zirconium 

and hafnium. In general, carbonatites, soils, stream sediments, 

and weathered rock may be enriched in iron, niobium, phosphorus, REEs, thorium, and uranium before mining (Modreski 

and others, 1995); if baddeleyite is present, zirconium 

and hafnium may also be enriched. Because most of these 

elements (that is, hafnium, niobium, REEs, thorium, uranium, 

and zirconium) occur in relatively stable insoluble minerals, 

the stream water and groundwater around the deposits may not 

display the same geochemical signature as the soils and stream 

sediments (Modreski and others, 1995).

Only small amounts of zirconium and hafnium are 

recycled. Less than 1 percent of discarded zirconium and 

hafnium metal or alloy is recycled or reused worldwide, and 

the recycled content of zirconium in products globally is only 

between 1 and 10 percent; estimates were not available for 

hafnium (Graedel and others, 2011). The recycled content 

includes both new (loss during manufacturing products that is 

transferred to the scrap market) and old (post-consumer) scrap 

(Graedel and others, 2011). Most of the zirconium recycled is 

from scrap generated during metal production and fabrication, 

and hafnium metal recycling is insignificant (Loferski, 2013b). 

The low proportions of these minerals that are recycled may 

be a reflection of the complex materials, such as electronic 

equipment, that they are used in; the technical difficulty of 

recycling them; and the small amounts used in these products. 

Also, the long lifetime of some products in which zirconium 

and hafnium are used make the minerals’ return to production 

very slow. One recent research study suggests a new method 

for economically recovering and recycling zirconium from 

used nuclear fuel cladding (Collins and others, 2011).

Mine Waste Characteristics

The volume and characteristics of mine waste produced 

from the recovery of zirconium and hafnium depend on 

the deposit type exploited and its size. Currently, the most 

important deposits for zirconium and hafnium are heavymineral sands and carbonatites.

Volume of Mine Waste

Processing heavy-mineral sands at a mine site produces 

waste material that includes tailings and slimes. Mechanical 

separation produces quartz-rich tailings after the heavy 

minerals are separated out. During wet concentration, 

slimes, which consist of a slurry of fine-grained materials 

(commonly clay minerals) are separated from the quartz 

and heavy minerals and mixed with a flocculent to cause 

settlement of the clay minerals and other clay-sized particles. 
Environmental Considerations V15

The percentages of heavy minerals and clay minerals in 

a deposit are variable and consequently result in various 

amounts of the different types of waste. Generally, heavy 

minerals make up from 1 to 25 percent of the material 

in economic deposits but additional natural sorting can 

concentrate them further. Zircon is only a small fraction of 

the total heavy-mineral content of a deposit. Heavy-mineral 

sand deposits in Western Australia, two deposits in India 

(Sattankulam and Kuttam), and one deposit in southeastern 

Virginia (Old Hickory) contain up to 10 percent heavy 

minerals (Murty and others, 2007; Zimmerman and others, 

2008). In contrast, heavy-mineral sands on the northeast 

coast of Sri Lanka (the Pulmoddai deposit) are very high 

grade and contain up to 80 percent heavy minerals, including 

ilmenite (70 percent), zircon (8 to 10 percent), rutile (8 

percent), sillimanite (1 percent) and monazite (0.3 percent); 

approximately 2 percent of the sand is clay sized (Premaratne 

and Rowson, 2003). Approximately 15 percent, by volume, 

of the two heavy-mineral sands deposits in India is clay 

sized and would be separated out and disposed of as slime 

(Murty and others, 2007). The measured reserves of these two 

deposits are estimated to be 400 million metric tons of raw 

sand that covers approximately 120 km2 (Murty and others, 

2007). The amount of waste generated by these two deposits 

would be much greater than for the smaller (3.2 km2) but 

higher grade deposit in Sri Lanka (Premaratne and Rowson, 

2003). For another comparison, up to 50 km2 of land could 

potentially be mined for heavy minerals d in the coastal plain 

region of the Southeastern United States (Daniels and others, 

2003). The modern mining of these deposits, along with many 

other heavy-mineral sand deposits worldwide, commonly 

involves removing the organic soil layer (if present), mining 

the sands, and then backfilling with tailings and slimes and 

replacing the organic soil layer. For operations that restore 

the landscape to pre-mining conditions, the volume of waste 

and its impact on the landscape is generally only a temporary 

land-use issue.

A variety of types of mine waste have been produced 

at the Kovdor Mine in Murmanskaya Oblast, Russia, which 

is the only carbonatite deposit from which zirconium and 

hafnium are currently being recovered. The Kovdor carbonatite complex in Russia has been mined for a variety of ores 

that include apatite-baddeleyite-magnetite, phlogopite, and 

vermiculite, and estimating the large expanses of mine waste 

within the complex is challenging. One report by the Kola 

Science Center of the Russian Academy of Sciences (1999) 

reported volume estimates for waste produced before the end 

of 1994. The amount of overburden rock stripped and stored at 

the mine site was estimated to be more than 700 million metric 

tons; a small fraction of the overburden was used for construction purposes. In addition to overburden, mine waste included 

more than 61 million metric tons of low-iron-content rock 

and approximately 10 million metric tons of apatite-staffelitebearing rock. (Staffelite is carbonate-rich fluorapatite.) The 

site contained a total of nearly 200 million metric tons of 

tailings, including 64 million metric tons from the processing 

of magnetite ore that was available for recovery of apatite and 

baddeleyite. As of 1994, 6 million metric tons of tailings had 

already been processed to recover apatite, baddeleyite, and 

magnetite (Russian Academy of Sciences, 1999).

Mineralogy and Chemistry of Slimes and Tailings

The mineralogy and chemistry of the slimes and tailings 

produced at heavy-mineral sand mines worldwide have many 

similarities; however, differences in the proportions of some 

minerals result in some variations. For example, slimes from 

the mineral-sand deposits in India contain mostly hematite, 

kaolinite, and quartz (Murty and others, 2007), whereas slimes 

from the Yoganup North Mine in Western Australia contain 

predominantly kaolinite (80 percent) and lesser amounts of 

quartz (10 percent) and goethite (10 percent) (Fujiyasu and 

others, 2000). Another example is the heavy-mineral sand 

deposit on the northeast coast of Sri Lanka, where the separation process for the concentrate generates a low-density 

fraction (24 percent) of quartz-rich tailings, a nonmagnetic 

heavy fraction (21 percent) that contains rutile and zircon, and 

a magnetic heavy fraction (55 percent) of mostly ilmenite and 

pseudorutile (Premaratne and Rowson, 2003).

Several studies have been conducted that characterize 

the composition of waste dumps in various stages of reclamation (that is, of reclamation pits) at heavy-mineral sand 

processing sites in Florida and Virginia (Daniels and others, 

1992, 2003; Orndorff and others, 2005). In 2007, the deposit 

at the Old Hickory Mine in Virginia contained an average of 

7 percent heavy minerals and an average of 35 percent clay 

minerals (the range of clay minerals was from 5 to 50 percent) 

(Zimmerman and others, 2008); the heavy minerals found at 

this mine were very similar to those contained in the nearby 

Brink and Concord deposits and included ilmenite, leucoxene, 

rutile, and zircon. One study examined soil profiles from 

within the reclamation pits at the Old Hickory Mine (these 

materials are referred to as “soils” despite being altered by 

mining). The reclamation pit soils were mixed slimes and 

tailings deposited as a slurry and left to dewater; then, soil 

amendments, including biosolids, fertilizer, and lime, were 

added. The slimes, which were composed mostly of kaolinite 

and lesser amounts of other clay minerals (vermiculite and 

smectites), were shown to separate from the quartz-rich 

tailings in the pits, which leads to lateral and vertical variability within the pits. The surface horizon (that is, the upper 

10 to 24 centimeters) in the pits, most of which had been 

amended with biosolids and lime, had near neutral pH values 

and contained up to 1.5 percent organic matter. In contrast, 

samples collected from below the surface material were generally acidic and contained little organic matter (Orndorff and 

others, 2005). Extractable concentrations of several elements, 

including phosphorus (<100 ppm), magnesium (<300 ppm), 

potassium (<400 ppm), and calcium (<3,000 ppm), were 

generally low in both surface and subsurface samples, despite 

fertilizer and lime additions (Daniels and others, 2003). 

Although no obvious deficiency symptoms were noted in 
V16 Critical Mineral Resources of the United States—Zirconium and Hafnium

vegetation growing on the pits, the acid extractable amounts 

of some micronutrients, such as copper (< 2 ppm), zinc 

(< 6 ppm), and boron (<1 ppm) were low (Daniels and others, 

2003). Another study (Daniels and others, 1992) examined 

mine soils from the Green Cove Springs Mine in Florida. 

The reclamation pit soils there are similar to those at the 

Old Hickory Mine in that the amended surface layers are 

more organic-rich than the subsurface samples; in contrast 

to the Virginia sites, however, very little clay material was 

found in the Florida deposit (Daniels and others, 1992). Both 

the natural and mine soils in the subsurface are acidic at the 

site in Florida.

The composition of mine waste produced from the 

recovery of baddeleyite from igneous deposits is very 

different from that produced from the recovery of zircon 

from heavy-mineral sand deposits. At the Kovdor deposit in 

Russia, baddeleyite is recovered as a byproduct of apatite and 

magnetite mining. Currently, the minerals are recovered from 

the ore in three sequential steps that begin with separating 

out the magnetite by magnetic concentration. The tailings 

from this first step are then processed by flotation to recover 

apatite and, finally, the post-apatite tailings are used to recover 

baddeleyite. According to Petrov (2004), the final tailings 

at the Kovdor Mine consist almost entirely of calcite and 

forsterite. Tailings produced before the recovery of apatite 

(recovery began in 1975) and baddeleyite (recovery began 

in 1977) are currently being reprocessed to recover these 

minerals. Kononov and others (1992) reported that these 

tailings contain the following: forsterite (50 to 60 weight 

percent), carbonates (20 to 25 weight percent), apatite 

(5 to 10 weight percent), and baddeleyite (<0.15 weight 

percent). Similarly, Zosin and others (2011) reported that the 

mineralogy of the tailings currently being produced at the site 

is dominated by apatite, carbonates, and forsterite (greater 

than 25 weight percent each), along with some phlogopite 

(approximately 12 weight percent), and lesser amounts of 

diopside and magnetite. In addition to tailings, overburden 

is an additional waste product and contains mostly pyroxene 

and amphibole (approximately 40 weight percent combined 

total), carbonate (approximately 20 weight percent), and minor 

amounts of apatite, forsterite, mica, and other minor minerals 

(<12 weight percent each) (Zosin and others, 2011).

In addition to the Kovdor deposit, baddeleyite has been 

recovered in the past as a byproduct of copper and phosphate 

(apatite) mining of the Palabora carbonatite complex in 

Limpopo Province, South Africa. Tailings from the main 

operation circuit were reprocessed by gravity separation to 

recover baddeleyite and uranium oxides. In 1990, approximately 15,000 metric tons of baddeleyite were recovered from 

29 million metric tons of ore (Towner, 1992, p. 4). The mine 

waste and tailings produced from exploiting this complex have 

a different mineralogy and chemistry than those generated at 

Kovdor because of differences in the compositions of the ores. 

In addition to apatite, baddeleyite, magnetite and vermiculite, 

the Palabora ore also contains economic amounts of copper 

sulfides (primarily chalcopyrite and bornite) and uranium 

minerals (uraninite-thorite).

Acid-Base Accounting

Little information quantifying the acid-generating 

and acid-neutralizing potentials of mine waste from heavymineral sands and from the Kovdor carbonatite complex is 

available. One study in Australia examined the potential of 

heavy-mineral sands to generate acid at a specific deposit 

(Landers and Usher, 2012). This work revealed that potentially 

acid-generating material was found within the deposit owing 

to the presence of pyrite. Sulfur at this deposit is present 

almost entirely as sulfide and occurs in the organic-rich layers 

of overburden materials (at a median grade of 0.66 percent 

sulfur) and in the orebody (at a median grade of 1.10 percent 

sulfur). The maximum potential acidity (that is, the maximum 

amount of sulfuric acid that could be produced) was low 

to moderate, with median values of 20 and 34 kilograms 

of calcium carbonate per metric ton (kg CaCO3/t) reported 

for organic-rich overburden and ore, respectively. The acid 

neutralizing capacity (that is, the amount of acid the sample 

can neutralize) was extremely low, and the paste pH (that is, 

the pH of water after it is mixed with a sample) values were 

highly acidic for ore and organic-rich overburden (Landers 

and Usher, 2012). Overall, much of the heavy-mineral sands 

ore material and some of the overburden have the potential 

to generate acid; the overburden that is not organic-rich is 

unlikely to generate acid. In general, heavy-mineral sands do 

not contain significant amounts of carbonates and offer little 

buffering capacity. Therefore, even trace amounts of sulfides 

can result in an acid-generating material. At the Green Cove 

Springs Mine in Florida, natural as well as reclamation pit 

soils were acidic (Daniels and others, 1992) and, at the Old 

Hickory Mine in Virginia, the subsurface material in the 

reclamation pits was generally acidic (pH<5.5) (Orndorff and 

others, 2005). At the Old Hickory Mine, the clay-rich layers 

within the waste piles were commonly strongly acidic and any 

neutral soil pH was at the surface because of the addition of 

lime to the piles.

In contrast to heavy-mineral sand deposits that commonly 

lack carbonate minerals, carbonatite deposits are carbonate 

rich. Pyrite, pyrrhotite, and several other sulfides that may 

be present in carbonatites have the potential to produce 

metal-rich acidic drainage but, in most cases, carbonates in 

the host rocks neutralize the acidity (Modreski and others, 

1995). The Kovdor carbonatite deposit contains chalcopyrite, 

pyrite, and pyrrhotite but is dominated by calcite and dolomite 

(Petrov, 2004). Reshetnyak and others (2008) noted high 

concentrations of sulfate as well as high pH for groundwater 

draining from the Kovdor Mine site. The source of sulfate is 

likely from the oxidation of sulfides, which can produce acid, 

depending on the sulfide mineralogy; the high pH results from 

buffering during the dissolution of carbonates.
Environmental Considerations V17

Human Health Concerns

Zirconium and hafnium are likely not essential for 

human health and generally are considered elements of low 

toxicity to humans (Reimann and de Caritat, 1998). They 

can enter the human body by ingestion, inhalation, or dermal 

contact. Guidelines for drinking water and the consumption 

of aquatic organisms have not been established by the 

U.S. Environmental Protection Agency (EPA) for zirconium 

or hafnium and their toxicities have not been reviewed by the 

Agency for Toxic Substances and Disease Registry (ATSDR) 

or the World Health Organization (WHO). Also, the carcinogenic potentials of zirconium and hafnium have not been 

examined by the International Agency for Research on Cancer 

(IARC) (U.S. Environmental Protection Agency, 2012).

Human health concerns regarding exposure to zirconium 

and hafnium may be associated with industrial inhalation and 

dermal exposure. The EPA, in an attempt to define a toxicity 

value for soluble zirconium to be used at hazardous waste 

sites, reviewed relevant toxicity literature on human and 

animal studies from 1900 to 2012 and found that not enough 

information was available to derive a Provisional PeerReviewed Toxicity Value for Superfund (U.S. Environmental 

Protection Agency, 2012). The EPA has not conducted a 

comparable investigation on hafnium toxicity. The following 

is a summary of the findings reported by the EPA on zirconium 

(U.S. Environmental Protection Agency, 2012):

• Ingestion of large quantities of zirconium was shown 

to cause health issues in humans in one study, but this 

type of oral exposure is not likely to occur even in 

industrial settings.

• Some inhalation studies of workers producing or 

working with zirconium and hafnium metals and 

compounds have reported respiratory tract irritation, dermatitis, pulmonary fibrosis, and one case 

of an acute allergic hypersensitivity reaction that 

led to death.

• Several other industrial inhalation exposure studies 

on zirconium compounds found no abnormities or 

health issues.

• A few dermal exposure studies suggest zirconium 

may cause dermal lesions.

In addition to the EPA report (U.S. Environmental 

Protection Agency, 2012) and references therein, evidence that 

human exposure to zirconium compounds may cause inflammation of the skin was reported in Schaller (2004). The only 

toxicity symptoms reported from hafnium were for individual 

allergic sensitivity (Kabata-Pendias and Mukherjee, 2007).

Mining and dry separation processes conducted on 

heavy-mineral sands creates high amounts of airborne dust, 

which leads to inhalation and, to a lesser extent, dermal exposure. Elevated concentrations of zirconium have been detected 

in the lungs, blood, and urine of heavy-mineral sands miners 

(Patty and others, 1981). In the United States, the permissible 

exposure limits for zirconium and hafnium compounds in an 

8-hour workday are 1.3 and 0.07 ppm (5 and 0.5 milligrams 

per cubic meter), respectively (Occupational Safety and Health 

Administration, 2013). The radioactivity of the dust is also a 

concern during processing steps that concentrate radioactive 

minerals. For example, dry processing heavy-mineral sands 

to concentrate monazite increases radiation levels. Control 

measures, such as adequate ventilation and air filters, reduce 

the exposure levels to just the dust encountered during mining 

and milling operations.

Ecological Health Concerns

Because of the low solubility of zirconium and hafnium, 

ecological health concerns in the aquatic environment 

are minimal. Suter (1996) proposed values of between 

982 and 54.9 ppb zirconium as secondary acute and chronic 

toxicological benchmarks, respectively, for freshwater 

biota; hafnium was not considered. The benchmarks were 

established using fewer data than required to establish EPA 

aquatic water-quality criteria (Suter, 1996). For comparison to 

the benchmarks, mean concentrations of zirconium in stream 

waters worldwide are between 0.9 and 90 ppb (Reimann 

and de Caritat, 1998). An estimated lowest chronic toxicity 

value for fish is 548 ppb based on a lethal concentration 

for 50 percent (LC50) of the test animals (Suter, 1996). 

Borgmann and others (2005) reported that the toxicity risk 

of hafnium to freshwater amphipods (Hyalella azteca) is low 

and comparable to that of zirconium. The LC50 in controlled 

laboratory tests was estimated to be greater than 1,000 ppb for 

both zirconium and hafnium (Borgmann and others, 2005)—

values that are more than 10 times greater than the worldwide 

mean concentrations.

Similar to the aquatic environment, ecological health 

concerns from zirconium and hafnium in soils are minimal. 

Only a small proportion of zirconium in soil is taken up by 

plants, and it does not accumulate in the food chain. Also, the 

low mobility of zirconium within the plant organs is evidenced 

by its commonly higher concentrations in the roots than in the 

top vegetation (Kabata-Pendias and Mukherjee, 2007). Despite 

the low soil mobility and bioavailability of zirconium, Shahid 

and others (2013) and references therein reported some cases 

of reduced plant growth and modifications in enzyme activities 

owing to zirconium uptake by plants.

Although zirconium and hafnium are usually not of 

environmental concern from mining, other associated metals 

are potentially deleterious to the environment, especially if 

concentrated in mine waste or tailings. Tailings from some 

carbonatite mines in particular contain significant concentrations of thorium and uranium (Modreski and others, 1995) 

and other heavy metals. For example, waters draining tailings 

piles at the Kovdor Mine in Russia contain copper, manganese, 
V18 Critical Mineral Resources of the United States—Zirconium and Hafnium

and strontium in concentrations higher than local water quality 

guidelines (Zosin and others, 2011). Also, apatite, baddeleyite, 

and zircon from the Kovdor deposit have been shown to be 

enriched in thorium and uranium (Amelin and Zaitsev, 2002) 

and tailings containing these minerals have the potential to be 

radioactive. The final baddeleyite concentrate at the Kovdor 

Mine was reported to contain low radioactivity (59 becquerels 

per gram) (Petrov, 2004), suggesting that the tailings, likewise, 

are only minimally radioactive.

Another ecological concern specifically related to mine 

sites involves the failure of tailings dams. This hazard poses 

an ecological concern at heavy-mineral sand mines and 

igneous rock deposits. In 2010, a settling pond at the Moma 

heavy-mineral sands mine in Mozambique breached its 

retaining wall and released water, sand, and clay that flooded 

a nearby village, causing one fatality and extensive damage to 

property (Gambogi, 2012a). Also, the release of large amounts 

of mining waste into downstream waterways physically 

suffocates aquatic biota and releases metals into waters and 

sediments. No known tailings dam failure at a heavy-mineral 

sand mine has taken place in the United States.

Mine Closure

Heavy-mineral sand mining leads to increased erosion 

rates when improperly managed, disturbs the surrounding 

ecosystem, and alters the character of the land. Tailings and 

slimes are used as backfill or disposed of in large waste piles. 

Commonly, the topsoil initially removed at a site is stored 

during mining and either mixed with tailings and slimes or 

placed on top of waste piles; topsoil and other amendments, 

such as lime, fertilizers, and biosolids, added to the waste 

facilitate the return of vegetation. The disturbed mined lands 

can be successfully returned to local region agricultural levels 

when these sorts of reclamation practices are used based on 

results reported in Orndorff and others (2011) and Zimmerman 

and others (2008) from the Old Hickory Mine in Virginia. At 

the Green Cove Springs Mine in Florida, returning the land 

to pine forest involved rebuilding the landform with tailings, 

replacing topsoil, reestablishing the growth of the pines and 

ensuring surface stabilization from re-colonization of grasses 

(Brooks, 2000). In Australia, farmland in Western Australia 

was reclaimed for grazing by mixing clay slimes into the 

soils to improve their capacity to hold water and adsorb plant 

nutrients. Also in Australia, sand dunes mined in New South 

Wales and Queensland were reclaimed by reestablishing the 

original slope and landforms and performing surface stabilization and revegetation (Brooks, 2000). In general, many of 

the lands mined for heavy-mineral sands have been reclaimed 

to native conditions and returned to productive uses.

Closure plans for the current mining operation at the 

Kovdor complex will involve management of the large open pit 

as well as the extensive waste piles. Future plans for the mine, 

which is expected to continue operations until at least 2032, 

include developing a deeper mine pit that will reach a depth of 

900 m by expanding the current Zhelezny open pit operation 

that is 2.4 km long, 1.6 km wide, and 170 m deep without any 

significant enlargement of the pit rim (Ellenbecker, 2011). At 

mine closure, the large open pit is likely to be partially backfilled and flooded. Also, overburden, mine waste, and tailings 

piles that are presently several kilometers in length are expected 

to be expanded by future operations (Russian Academy of 

Sciences, 1999). Reclamation options for the waste piles 

include regrading, capping, adding soil amendments, revegetating, and possibly treating drainage and seepage waters.

Problems and Future Research

Because of the relative ubiquity of zircon in crustal rocks 

and the worldwide distribution of coastal environments with 

favorable geometries and conditions for placer formation 

(fig. V3), the outlook for zirconium and hafnium supply is 

generally favorable. In modern coastal environments, an integrated understanding of coastal morphology, sediment supply 

and distribution systems, geologic setting, and climate history 

is key for identifying and delineating heavy-mineral concentrations. In older coastal systems, relatively recent discoveries 

of significant heavy-mineral deposits in the Eucla basin of 

southern Australia (fig. V3; Hou and others, 2011) highlight 

the potential usefulness of a detailed reconstruction of major 

sea-level events combined with models for heavy-mineral 

concentration. The Gulf of Mexico Basin of the southern 

United States is a particularly prospective region with a long 

history of Cenozoic deposition and volumetrically significant 

Paleogene through Miocene coastal plain and shore zone facies 

deposits (Galloway and others, 2000). There is also potential for 

recoverable zirconium and hafnium resources in a wide variety 

of settings, including from tailings, where zircon is or was not 

the primary commodity of interest but may still be present at 

economic concentrations and is also relatively easy to extract.

Zirconium- and hafnium-bearing minerals have generally 

low solubility and reactivity; consequently, these elements do 

not pose significant threats to human health or the health of 

ecological systems. In heavy-mineral sand mining operations, 

the main environmental impact is landscape disruption by 

surface mining. It is common for modern mineral-sand mines 

to restore the landscape to pre-mining conditions. Thus, in 

many cases, the volume of mine waste and its impact on the 

landscape is generally only a temporary land-use issue.
References Cited V19

References Cited

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For more information concerning this report, 

please contact:

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Schulz and others, eds.—Critical Mineral Resources of the United States—Economic and Environmental Geology and Prospects for Future Supply—Professional Paper 1802