Bureau of Mines

Report of Investigations 5243

SELECTIVE EXTRACTION OF MERCURY

AND ANTIMONY FROM CINNABAR-STIBNITE ORE

BY E. G. ERSPAMER AND R. R. WELLS

p

i

_I -- United States Department of the Interior- September 1956

... 1 ---
SELECTIVE EXTRACTION OF MERCURY

AND ANTIMONY FROM CINNABAR-STIBNITE ORE

«

BY E. G. ERSPAMER AND R. R. WELLS

* * * * * * * * * Report of Investigations 5243

UNITED STATES DEPARTMENT OF THE INTERIOR

Fred A. Seaton, Secretary

BUREAU OF MINES

Marling J. Ankeny, Director

Work on manuscript completed November 1955. The Bureau of Mines will welcome reprinting of this paper, provided

the following footnote acknowledgment is made: "Reprinted from Bureau of Mines Report of Investigations 5243."

September 1956
SELECTIVE EXTRACTION OF MERCURY AND ANTIMONY

FROM CINNABAR-STIBNITE ORE

by

E. G. Erspamerl/ and R. R. Wells2J/

CONTENTS

Page

Summnary .......................... ,............. 1

Introduction ................... eww=... e=,e, .* .., 1

Concentration of mercury ores ........................ .... 2

Previous laboratory studies of Red Devil ore ................. 4

Current investigations .................................. .... 5

Scope *............. ............. ....... 5

Mineral-dressing studies ................................. 5

Character of the ore .............................. 5

Physical ................... .. ........... 5

Chemical ........................ ........... 5

Preliminary tests ..................... 6

Screen analysis ................................ 6

Heavy-liquid test .............................. 6

Methods of concentration ..... ................. . 6

Gravity concentration ........................ 6

Bulk flotation ................................. 7

Selective flotation ............................ 8

Discussion ........ .9... ........................ 9

1/ Formerly metallurgist, Bureau of Mines, Region I, Juneau, Alaska.

/ Supervising metallurgist, Bureau of Mines, Region I, Juneau,

Alaska.

In preparing this report, the authors were assisted by J. V.

Batty and G. M. Potter, metallurgists, Intermountain Experiment

Laboratory, Salt Lake City, Utah, and C. T. Baroch, metallurgist, Boulder City, Nev.

Report of Investigations 5243

I
ii

CONTENTS (Con.)

Pag

Current investigations (Con.)

Furnace-reduction studies .... ......................... 9

Experimental procedure ..... ............ ...... 9

Results ............ ..................................... 9

Excess air ..... ........................... 9

Limited air ..................................... 9

Discussion .............. ........................ 10

Hydrometallurgical studies ......... .... ........................ 10

Library research .. ............................... 10

Leaching ........... ........................... 11

Results ......... .............................. 11

Precipitation ......... ...... .. 13

Discussion ....................... 13

Bibliography ................................. , ..... 13

TABLES

1. Chemical analysis of Red Devil ore ...................... ...........

2. Spectrographic analysis ............ ................... 5

3. Screen analysis ............................................ 6

4. Jig and table test ...... ............................. 7

5. Locked flotation test .... ....................... .,..... 7

6. Selective flotation, Red Devil ore ........ ................ 8

7. Leaching tests on 1:1 cinnabar-stibnite mixtures .......... 11

8. Leach tests on bulk flotation concentrate ................. 12

ILLUSTRATION

Fig.

1. Location map, Kuskokwim mercury area ...................... 3
a

1

SUMARY

Laboratory investigations were made by the Bureau of Mines on ore containing

the sulfides of both mercury and antimony to develop a method or methods for the

economic recovery of mercury metal and the concurrent recovery of antimony as metal

or salts. If standard mercury-recovery techniques are applied to such ores, the

antimony content interferes with the effective production of mercury metal; furthermore, the antimony is lost.

This report presents the results of tests on a sample of cinnabar-stibnite ore

from the Red Devil mine near Sleitmute, Alaska. Red Devil ore was chosen for the

investigation because it is purported to be typical of ores of the Kuskokwim River

area, which, in their aggregate, constitute large potential reserves of mercury and

antimony. Studies included petrographic, spectrographic, and chemical analyses,

gravity-separation tests, bulk flotation of mercury and antimony sulfides, differential flotation, leaching, and furnacing of crude ore and of flotation concentrate.

Results of this investigation indicate that bulk flotation of sulfides, followed by furnacing of the flotation concentrate with a carefully regulated admission

of air, would yield an overall recovery of more than 95 percent of the mercury in

the ore and would permit possible recovery of antimony from the furnace calcine.

Similar treatment is reported tp have been employed successfully on livingstonite

ore from Guerrero, Mexico (j5) ./

Results of preliminary hydrometallurgical tests indicated that leaching of

cinnabar-stibnite ores or concentrates with sodium sulfide solution is a potential

means of obtaining separate mercury and antimony products.

INTRODUCTION

Quicksilver has long been noted as a metal with tremendous and sudden price

fluctuations. Recent fluctuations in price have included an increase from $75 per

76 -pound flask shortly before Wolrd War II to $196 per flask during that war, a

sharp drop to $70 per flask in June 1950, and an equally sharp rise to $233 per

flask by January 1951. Since 1951 the price has maintained a high level; it has

not dropped below $180 per flask and climbed as high as $325 per flask in October

1954.

The rate of increase in consumption of quicksilver is well illustrated by the

following quotation from C. N. Schuette (25, p. 326).

/7 Underlined numbers in parentheses refer to citations in the bibliography at the

end of this report. Page references refer to pages in the citation and not in

this report.
2

In the five pre-World War II years of 1936 through 1940, the average

annual consumption was 27,340 flasks, the peak being 35,000 flasks in

1937. In the five World War II years of 1941 through 1945, the average

annual consumption was 50,866 flasks, the peak year being 1945 with a consumption of 62,429 flasks. In the five post-World War II years of 1946

through 1950, the average annual consumption was 40,491 flasks, 1950 consumption reaching a 49,215-flask peak.

The two factors - increased consumption and increased average price - have

stimulated domestic quicksilver production and have renewed interest in the mercury

deposits of southwestern Alaska, especially those of the Kuskokwim Basin. (See

fig. 1.) Field examinations, coupled with significant ore discoveries made as a

result of an exploration program at the Red Devil mine during the summer of 1955,

indicate that the Kuskokwim area has large potential reserves of mercury ore.

In some Kuskokwim deposits cinnabar is associated intimately with significant

quantities of stibnite. Arsenic minerals are present in smaller amounts. The presence of these volatile impurities is detrimental to economic recovery of quicksilver

by standard mercury-recovery techniques. At the Red Devil plant it was determined

that nearly complete oxidation of the antimony and arsenic was required to effect

good recovery of quicksilver. A large amount of the oxides of these impurities

therefore was collected in the condensing system. As a result of the presence of

the bulky oxide precipitate, treatment of the soot for recovery of mercury was difficult and costly.

No records of mercury losses at thB Red Devil recovery plant were maintained.

It is generally estimated, however, that recovery has averaged less than 70 percent

and has been especially poor on ores of high antimony content.

The laboratory study was conducted with a twofold purpose. First, it was desired to devise a method that would effect more complete and more economic extraction of mercury from the ore. Second, it was deemed desirable to determine the

feasibility of making an antimony byproduct and thus to utilize a commodity that is

wasted by the standard furnacing methods now used.

CONCENTRATION OF MERCURY ORES

Most mercury ores are roasted directly or are retorted for recovery of the

metal without prior concentration. Recoveries are high; costs usually are nominal.

SchueTte (25, p. 325) reports that the New Almaden mine in California operated at a

profit during World War II on ore averaging only 3 pounds of mercury per ton.

Walter W. Bradley (1) conducted a large number of laboratory concentration

tests in 1919; he concluded that the ultimate decision between straight furnace reduction and combined concentration and reduction treatment was a matter of comparative costs and comparative extractions and that usually the combined treatment had

doubtful economic value. Under most conditions, it is probable that the cost of the

added grinding and concentrating operations is greater than the saving inherent in

roasting only a portion of the ore. For this reason, aside from a few small

operations, concentration of mercury ores never has been accepted for plant practice.

When unusual conditions exist, however, a combined treatment process deserves

consideration. In the Kuskokwim area the cost of erecting and equipping a large

plant and the cost of fuel for roasting are major items. The possibility exists,

therefore, of saving enough to offset the cost of concentration, provided the

concentration is efficient both in recovery and in ratio of concentration.
0- 0 0 0

Figure 1. - Location map, Kuskokwim mercury area ao'lJ
4

In spite of the high specific gravity of cinnabar, gravity methods rarely are

applicable for concentrating mercury ores. Cinnabar slimes readily; thus, considerable loss is involved if the slimes are discarded. Addition of slimes to the gravity concentrate results in a low-grade product containing large amounts of fine

gangue; such products are difficult to filter and dry and, in addition, usually

present difficulties in subsequent reduction treatment.

Laboratory work by Shaffer, Newton, and Fahrenwald (27), Rey and Brevers (23),

Wells (41), and others indicates that cinnabar is readily amenable to flotation.

Tailings containing 0.2 to 1.0 pound of IHg per ton are not uncommon. High-grade

concentrates, amenable to reduction by retorting or roasting, usually are obtained.

PREVIOUS LABORATORY STUDIES OF RED DEVIL ORE

Early in 1947 a sample of cinnabar-stibnite ore from the Red Devil mine was

submitted to the Salt Lake City laboratories to determine a practical method of

separating the two minerals into a furnaceable mercury concentrate and a marketable

antimony concentrate. Unfortunately, the sample, taken from a high-grade ore lens,

was not representative of run-of-mine ore. Nevertheless, a large number of tests

were made which resulted in the development of bulk and selective flotation techniques that served as a basis for the current investigation.

The sample tested at Salt Lake City assayed 9.43 percent Hg and 16.6 percent

Sb. Cinnabar and stibnite occurred in a gangue consisting chiefly of quartz and

kaolin, with a small amount of dolomite. The gangue rock was silicified and impregnated with fine needles of pyrite. Grinding to minus-100-mesh liberated the cinnabar and most of the stibnite.

Gravity methods of concentration yielded only partial separation of cinnabar

and stibnite.

A sample of the ore was ground to approximately 65-mesh with sodium cyanide to

depress pyrite; a cinnabar-stibnite bulk concentrate was floated with a higher xanthate collector. Lead acetate was used as an activator, and an alcohol frother was

employed. The rougher concentrate was cleaned once using small additional amounts

of collector and activator. The cleaned bulk concentrate contained about one-third

of the weight of the original ore, 97.8 percent of the total mercury, and 94.7 percent of the total antimony; it assayed 27.1 percent Hg and 47.4 percent Sb.

A selective flotation process for making separate cinnabar and stibnite products was developed after considerable investigation. A dichromate salt and a promoter (Minerec No. 194) were added to the grinding circuit. Stibnite remained depressed, but the cinnabar was floated readily. Sulfuric acid and lead acetate were

added to reactivate the stibnite, and an antimony product was floated using a higher

xanthate and an alcohol frother. Each concentrate was cleaned once. Applying this

method to ore ground to approximately 80-mesh, 96.0 percent of the mercury in the

ore was recovered in a cinnabar concentrate that assayed 80.3 percent Hg and 1.3

percent Sb. The antimony product contained 80.7 percent of the antimony; it assayed

0.82 percent Hg and 62.3 percent Sb. The mercury product contained 11.6 percent of

the original weight; the antimony concentrate contained 20.8 percent of the original

weight.
U

5

CURRENT INVESTIGATION

Scope

This investigation was directed toward a general survey of possible methods of

selective recovery of mercury and antimony rather than toward detailed study of any

one method. To this end, preliminary studies were made of mineral dressing, furnace

reduction, and hydrometallurgical treatments. The presentation of data and the discussion of results from each type of treatment are reported separately.

Mineral-Dressing Studies

Character of the Ore

The ore for this investigation was selected carefully to obtain a sample as

nearly as possible representative of the known ore body, both as to mineralogical

and chemical content.

The material as received at the laboratory was crushed and sampled to obtain

representative fractions for petrographic, spectrographic, and chemical analyses.

Physical

The sample contains essentially quartz and sericite, with some altered feldspar, carbonaceous matter, and stibnite, and less cinnabar and limonite. Small

amounts of chlorite, realgar, calcite, and kaolinite are present, as well as very

small amounts of gypsum, amphibole, stibiconite, and pyrite.

Cinnabar and stibnite are associated closely with the realgar and with each

other. Some of the cinnabar occurs as inclusions in stibnite, and some of the stibnite is included in cinnabar. The stibnite and cinnabar are essentially freed from

gangue minerals in the plus-100-mesh fraction; some of the stibnite, however, is so

fine grained that it remains locked, even in the minus-200-mesh fraction.

Chemical

The chemical analysis is given in table 1.

TABLE 1. - Chemical analysis of Bed Devil ore

Assay, percent

Hg Sb As Fe SiO2

1.6 l 2.7 -0.73 - 3.5 - 45.6

A semiquantitative spectrographic analysis showed the presence and approximate

amounts of the metals listed in table 2. Any other elements, if present, are in

quantities less than the minimum detectable by the routine method employed.

TABLE 2. - Spectrographic analysis

Ba A1 Sb ~As Be I Ca Cu I Mg Cr I e I Mn Ni i Si Ti V B

F C/ C C- F E E D F C/ D E A E E

Legend: A - over 10 percent E - 0.01 to 0.1 percent

B - 5 to 10 percent F - 0.001 to 0.01 percent

C - 1 to 5 percent G - less than 0.001 percent

D - 0.1 to 1 percent
6

Preliminary Tests

Screen Analysis

A representative portion of the ore was roll-crushed to minus-10-mesh and subjected to wet-screen analysis to determine the distribution of the mineral values in

relation to particle size using standard Tyler sieves. Results of the screen analysis are shown in table 3.

TABLE 3. - Screen analysis

Weight- Assay, ercent Distribution, percent

Product percent Hg Sb Hg Sb

Plus-20-mesh ........................ 25.82 1.52 2.42 22.7 24.2

20/35 .............................. 19.34 1.80 2.60 20.1 20.0

35/48 ................................. 7.32 1.98 2.79 8.4 7.9

48/65 ............................ 6.89 1.96 2.99 7.8 8.0

65/100 ............................ 5.46 1.90 2.74 6.0 5.8

100/200 .......................... 8.09 2.20 2.94 10.3 9.2

Minus-200-mesh sand ................ .10.21 3.40 2.98 20.0 11.8

Slime ............................ 16.87 .48 2.01 4.7 13.1

Calculated head ............. 00.00 100.0 100.0

The data in table 4 show that mineral liberation is incomplete in the particle

sizes for which gravity processes are applicable. The distribution of antimony in

each size fraction parallels closely the weight of the fraction; this demonstrates

intimate association of stibnite with other minerals throughout the entire size

range. A similar relationship exists between fraction weight and mercury content

in the material coarser than 100-mesh; this indicates that cinnabar is liberated

only in the finer sizes. The difference in liberation of cinnabar and stibnite is

further shown by the fact that the minus-200-mesh slime contained 13.1 percent of

the total antimony but only 4.7 percent of the total mercury.

Heavy-Liquid Test

To explore further the amenability of the ore to specific-gravity separation

processes, a heavy-liquid test was made. A representative portion of the ore was

roll-crushed to 10-mesh and screened at 35-mesh. The minus-10-, plus-35-mesh material was separated successively at specific gravities of 2.67 and 2.54 in tetrabromoethane. The minus-35-mesh portion was not treated.

The results of this test corroborated the findings of the petrographic examination and the screen analysis of the ore.

Methods of Concentration

Gravity Concentration

The possibility remained that gravity separation could be employed as a preliminary concentration step to remove a coarse concentrate. The gravity tailings

would represent a reduced tonnage for fine grinding and subsequent beneficiation

processes.

A representative portion of the ore was roll-crushed to minus-1/4-inch and

screened at 10-, 20-, 48-, 100-, and 200-mesh. The 1/4-inch to 10-mesh portion was

jigged in a laboratory Harz-type jig. The remaining size fractions were tabled
7

I

successively on a laboratory Wilfley table. Results of the combined jig-table tests

are shown in table 4. The composite concentrate contained 16.3 percent Hg and 18.0

percent Sb, with a recovery of 70.1 percent of the total mercury and 53.0 percent of

the total antimony. This product represented only 7.5 percent of the original sample weight; no tailing product obtained was low enough in mercury and antimony content to reject without further treatment. Therefore, a negligible weight rejection

was achieved by gravity concentration.

TABLE 4. - Jig and table test

Weight- Assay, ercent Distribution, percent

Product percent HSb _ _Sb

Jig concentrate .................... 5.06 9.68 15 28.2 31.1

Table concentrate .................. .41 30.2 23.0 41.9 21.9

Jig middling ....................... 12.61 .89 1.54 6.4 7.7

Table middling ..................... 3.30 1.15 4.45 2.2 5.8

Jig tailing ........................ 38.46 .43 .61 9.5 9.3

Table tailing ...................... 29.40 .45 1.12 7.6 13.0

Slime .............................. 8.76 .84 3.25 4.2 11.2

Calculated head . ................. 100.00 . 2.-53 100.0 100.0

Composite concentrate .............. . 747 18.0 70.1 53.0

Composite middling ................. 15.91 .94 2.14 8.6 13.5

Comosite tailing .............. 67.86 .44 .8 17.1 22.3

Bulk Flotation

The necessity for finer grinding to unlock cinnabar and stibnite from gangue

minerals dictated the use of flotation techniques on Red Devil ore. Batch tests

were made on samples ground to minus- 6 5- and minus-100-mesh, using various reagent

combinations to float a bulk sulfide concentrate. The best procedure was determined

to be grinding with sodium cyanide to essentially minus-65-mesh, conditioning with

lead acetate, and rougher floating with American Cyanamid Reagent 301 and Frother 63.

The rougher concentrate was cleaned once. Employing this procedure, a locked test

was made on 15 one-kilogram samples; cleaner tailings were added to the next successive rougher pulp. Results are shown in table 5. The bulk cinnabar-stibnite concentrate contained 23.4 percent Hg and 30.7 percent Sb, with a recovery of 96.9 percent

of the total mercury and 76.1 percent of the total antimony in the ore. Even with a

65-mesh grind, approximately 20 percent of the antimony remains locked with gangue

minerals.

More than 93 percent of the total weight of the ore was rejected as tailing;

only 6.32 percent would require furnacing or retorting for the production of mercury

metal.

TABLE 5. - Locked flotation test

Metallurgical data

Weight- Assay, ercent Distribution) percent

Productpercent Sb Hg Sb

Cleaner concentrate .................. 6.32 23. 30.7 96.9 76.1

Rougher tailing .................. 93.68 0.565 3.1 23.9

Calculated head ................. 100.00 1.53 2.55 100.0 100.0
8

TABLE 5. - Locked flotation test (Con.)

Operat inrg dlata

Time, Pb Reagent Frother

Circuit min. pH NaCN acetate 301 63

Grind ....................................... 1.0 - -

Conditioner ........................... 5 - 1.5 0.2 0.08

Rougher ............................... 7 6.0 - - .2

Cleaner ................................. 3 -- - -

Total .......................................... 1.0 1.5 .4 -08

Selective Flotation

A long series of tests was run in the attempt to produce separate mercury and

antimony flotation concentrates. No results approached those obtained by the Salt

Lake laboratory from the high-grade sample tested there. The dichromate depressant

technique developed by the Salt Lake staff, however, appeared to be superior to any

other method investigated, although the required consumption of sodium dichromate

was excessively high for treatment of the lower grade material.

The best selective flotation results were obtained by grinding minus-10-mesh

ore with sodium dichromate and promoter; a rougher mercury concentrate was floated

and was cleaned twice. The mercury rougher tailing was reground, and an antimony

concentrate was floated in an acid circuit. The antimony product was cleaned twice.

Like cleaner tailings were combined for assay. Results are summarized in table 6.

TABLE 6. - Selective flotation, Red Devil ore

Metallurgical data

Weight- Assay, percent Distribution, percent

Product percent Hg Sb Hg Sb

Hg cleaner conc. ...................... 1.10 7 .72 2.3 62.2 1.0

Hg cleaner tailings ................ 5.72 6.09 13.4 25.0 30.0

Sb cleaner conc. ........ 1.72 5.54 49.4 6.8 33.3

Sb cleaner tailings ................ 5.02 1.24 4.5 4.4 8.9

Rougher tailings ....... ....... 86.44 o _. -0.79 1.6 26.8

Calculated heads ................ 1000 1.4 2.6 100.0 100.0

Calculated Hg rougher conce... ...... [ '6.82r - 87.28 31.0

O erating data

Time, Pb Colletor Collector

Circuit min. pH N2Cr20 H2S04 acetate 1' 2/ Frother4/

Grind ....... 10 6.0 0.12

Hg rougher . 4 8.5 6.0 0.08

Hg cleaner 1. 4 3.0

Hg cleaner 2. 3 6.o0 04

Regrind 5 .

Sb rougher 5 4.9 2.0 0.3 0.15 .04

Sb cleaner 1. 3 .

Sb cleaner 2. 3 _ .4_ 

Total ________ 21.0 2.0 0.3 0.12 0.15 0.-24

I/l1 Less than.

Collector 1 - Minerec 27.

Collector 2 - American Cyanamid

Frother - Dowfroth 250.

reagent 301.
9

Discussion

By selective flotation, 87.2 percent of the total mercury was recovered in a

rougher concentrate that assayed 17.8 percent Hg and 11.6 percent Sb. Cleaning improved the grade to 78.72 percent Hg and 2.3 percent Sb, but with a sharp decrease

in mercury recovery; only 62.2 percent of the total mercury was recovered in the

cleaner concentrate. A recovery of 33.3 percent of the total antimony was made with

a grade of 49.4 percent Sb and 5.54 percent Hg.

In no test was selectivity sharp enough to allow production of an antimony concentrate low in mercury content; 10 to 20 percent of the mercury reported in the

antimony products. Thus, as both the mercury and antimony fractions would require

further treatment to yield end products, it was concluded that application of selective flotation to Red Devil ore is not promising.

Furnace-Reduction Studies

Experimental Procedure

Experiments were made on crude ore and on bulk flotation concentrate to volatilize the cinnabar and to retain the stibnite in the furnace calcine. A laboratory

horizontal-tube furnace was employed for this work. Charges were heated without

rabbling in 10-gram alundum boats. Air was regulated by applying suction to one

end of the furnace tube. Extraction calculations were based solely on weight and

analysis of the residue after calcining; no attempt was made to collect the furnace

fume. The limitation of charge size coupled with accuracy limitation of analyses,

yielded erratic results. The trend of recoveries, however, was indicative enough

to make general conclusions.

Results

Excess Air

For the purpose of studying the results that would be obtained by standard

mercury-recovery methods, a series of tests was run with a plentiful supply of air

to oxidize the cinnabar and to remove the mercury as volatilized metal. Results on

both crude ore and concentrate indicated mercury extractions of 98 to nearly 100

percent, but the results also demonstrated that this method of roasting converted

25 to 50 percent of the antimony to antimony oxide. These laboratory results agreed

with the results obtained at the Red Devil plant, where considerable trouble was experienced with large amounts of antimony oxides entering the condensing system with

the volatilized mercury.

Limited Air

A series of tests was run to determine the effect of limited air admission to

the furnace. It was found that oxidation of antimony could be minimized by regulating the air supply. The bulk of the mercury was removed as volatilized cinnabar,

which would require treatment in an oxidizing atmosphere to yield mercury metal.

Results of numerous tests, conducted on bulk concentrate between 450° and

650° C., showed consistent mercury extractions of 99.9 percent, with a retention

of 83 to 93 percent of the antimony in the calcine. A similar series of tests on

minus-10-mesh crude ore showed extractions of more than 98.5 percent of the mercury

and a retention of 80 to 90 percent of the antimony. Within the 450° to 650° C.

range the temperature of reduction did not appear to be a significant factor.
10

The results of the furnacing tests indicate that, with more precise control of

air supply than was possible with the equipment used, a nearly complete separation

of mercury and antimony can be made.

Discussion

The results of the laboratory-scale furnacing tests on bulk flotation concentrate suggest the possibility of applying a recovery method described by H. B.

Menardi (15). Menardi's small plant at El Segundo, Calif., was used to produce

separate mercury and antimony products from livingstonite (HgS.2Sb2S 3) flotation

concentrates from the Huitzuco deposit, Guerrero, Mexico (26). The concentrate was

heated in a concurrently fired rotary kiln with careful regulation of air to produce

volatilized cinnabar and to prevent excessive oxidation of stibnite. The volatilized cinnabar was oxidized in a combustion chamber to produce mercury vapor, which

was condensed to obtain metallic mercury. The stibnite-rich calcine was smelted and

refined to produce antimony metal. Roasting of cinnabar ores in concurrently fired

furnaces has been investigated further and has been employed successfully on a large

scale by the Bradley Mining Co. at its Reed and Sulfur Bank properties (11).

Assuming that controlled air-furnace treatment would give selective extractions

equal to those indicated by the laboratory tests and assuming that plant flotation

would yield recoveries equivalent to laboratory flotation tests, an overall recovery

of 97 percent of the mercury and 68 percent of the antimony could be expected from

combined flotation-furnacing treatment of Red Devil ore.

Further investigation and detailed study, preferably on a small, continuous

scale, would be required, however, to prove the practicability and economic feasibility of applying Menardi's process, or modification thereof, to Red Devil

concentrate.

Hydrometallurgical Studies

Library Research

A survey of the literature revealed no mention of previous application of the

hydrometallurgical methods to ores containing both mercury and antimony. Leaching

methods have been employed, however, on ores of each metal.

In one instance, mercury was recovered from amalgamation tailings by leaching

in a solution of sodium sulfide and sodium hydroxide. This process was described by

Thornhill (34). Experimental work using the process on mercury ores was done by

Bradley (1). Concerning antimony, Van Arsdale (36, p, _2) states that:

Stibnite is soluble in alkali sulfides or hydroxides, and a pure

antimony sulfide can be recovered from the resulting solutions by precipitation with carbon dioxide or sulfur dioxide. The method has yet

to be tested on a plant scale.

Mercuric sulfide is reported to be insoluble in caustic soda and potash solutions but readily soluble in sodium or potassium sulfide to form a complex thiocompound that is completely hydrolyzed into mercuric sulfide by dilution with water

(35, P. 119).
11

Leaching

For this investigation the initial laboratory leaching tests were made on synthetic 1-to-1 mixtures of pure cinnabar and stibnite. A bulk flotation concentrate

that contained 23.4 percent Hg and 30.7 percent Sb was used for a subsequent series

of tests. Lixiviants were sodium sulfide and sodium hydroxide, used separately and

in mixtures of both; one cursory test was performed in which concentrated hydrochloric acid was the lixiviant.

Ten-gram samples were leached at atmospheric pressure with various concentrations of leaching agent. Other variables were temperature of the pulp, weight ratio

of lixiviant to charge, and duration of leach. The pulps were agitated continuously

during the leaching cycles by an electric stirrer. After digestion the slurries

were filtered on a suction filter. The residues were washed and dried; residues and

filtrates were submitted for chemical analysis. All leaching tests were singlestage.

Results

Data obtained from

table 7.

leaching tests on cinnabar-stibnite mixtures are shown in

TABLE 7. - Leaching tests on 1:1 cinnabar-stibnite mixtures

Extraction,

Time, Strength, Quantity, Temp., percent of total

min. Lixiviant percent ml. C. . H Sb

10 NaOH 5 600P 20 1.8 91.7

10 Na2S 5H20 3 400 20 1.0 97.3

10 (NaOH'2S5 0 ( 2 400 2 1.0 97.1

NaOH 1

10 NaOH 3 400 20 1.0 86.5

10 (N2S'5 20 ( 400 20 1.0 96.7

15 NaOH 5 600 20 1.8 88.4

20 NaOH 5 600 20 1.8 89.6

10 NaOH 10 600 20 15.2 75.9

15 NaOH 10 600 20 7.8 86.1

20 NaOH 10 600 20 11.7 94.6

15 Na2S-5H20 5 600 20 4.5 97.4

20 Na2S'5H2 0 5 600 20 5.3 93.5

10 Na2S.5H20 10 600 20 99.9 87.3

20 Na2S'5H20 10 600 20 99.9 86.8

35 Na2S-5H20 10 600 20 99.6 89.7

180 HC1 Concentrated 600 20 1.0 99.4

Sodium hydroxide and sodium sulfide, in concentrations of 5 percent or less,

both dissolve antimony selectively from mixtures of pure cinnabar and stibnite;

sodium sulfide was slightly more effective than sodium hydroxide. At 3-percent

concentrations mixed sodium sulfide-sodium hydroxide leach solutions gave results

similar to those obtained with sodium sulfide alone under similar conditions. Tenpercent solutions of sodium hydroxide dissolved 7 to 15 percent of the mercury;

similar concentration of sodium sulfide effected nearly complete solution of the

mercury.

I

J
12

One cursory test was made in which a mixture of the pure sulfides was leached

in concentrated hydrochloric acid. The concentrated acid dissolved approximately 99

percent of the total antimony and only 1 percent of the total mercury. This method

was not pursued further because concentrated hydrochloric acid would be difficult to

handle in a leaching plant. Continued improvements in corrosion-resistant equipment,

however, might make this of more than academic interest in the future.

The facts that elevated temperatures were not required for dissolution of the

stibnite and that only weak solutions and a relatively short contact time were necessary indicated that the process warranted further study. Therefore tests were initiated on the bulk flotation concentrate. Table 8 shows the results obtained.

TABLE 8. - Leach tests on bulk flotation concentrate

Extraction,

Time, Strength, Quantity, Temp., percent of total

min. Lixiviant percent ml. °C. Hg Sb

5 Na2S.5H20 3 400 20 5.8 4t8.

10 Na2S-5H20 3 400 20 3.9 59.4

15 Na2S.5HO2 3 400 20 7.2 56.5

60 Na2S'S520 3 400 20 4.2 81.0

5 Na2S-5H20 5 400 20 4.1 63.6

10 Na2S-5H 20 5 400 20 6.4 81.5

15 Na2S'5H20 5 400 20 6.2 77.1

60 Na2S-5H20 5 400 20 5.9 94.3

5 Na2S'5H20 5 400 60 7.1 86.9

10 Na2S-5E20 5 400 60 5.9 87.8

15 Na2S'5120 5 400 60 9.8 83.1

60 Na2S'5H2 0 5 400 60 12.3 95.5

5 Na2S-5H20 10 400 20 16.8 92.7

10 Na2S5gH20 10 400 20 16.2 90.9

15 Na2S-5H20 10 400 20 18.0 81.9

60 Na2S.5H20 10 400 20 30.5 95.9

165 NaOH 6 200 20 38.6 49.9

165 NaOH 8 200 20 36.9 61.0

165 NaOH 10 200 20 58.2 80.7

360 NaOH 4 100 20 .4 31.3

1,140 NaOH 4 200 20 35.5 30.9

165 NaOH 6 200 70 27.7 4o.5

165 NaOH 10 200 70 45.7 73.9

165 NaOH 6 200 80 59.0 60.1

15 NaOH 4 200 99 8.3 93.4

15 NaOH 6 200 99 13.5 96.3

15 NaOH 8 200 99 28.1 96.9

15 NaOH 10 200 99 36. 95.1

Table 8 shows that greater difficulty was experienced with both sodium sulfide

and sodium hydroxide in obtaining selective extractions of the stibnite from the

bulk flotation concentrate than from synthetic mixtures. Sodium sulfide was definitely superior to sodium hydroxide as a differential leaching agent for the flotation concentrate. With a 60-minute contact time in a 5-percent sodium sulfide

solution at room temperature, approximately 6 percent of the total mercury and 94

percent of the total antimony were dissolved. The most effective leaching conditions, using sodium hydroxide, were a 15-minute contact time in a 4-percent solution

at the boiling point. Under these conditions approximately 8 percent of the total

mercury and 93 percent of the total antimony were dissolved.
13

Precipitation

The following chemical reactions are believed to be involved in leaching cinnabar and stibnite with sodium sulfide:

Sb2S3+3Na 2S -- SNa3 SbS 3

HgS+Na 2 S -- Na2 HgS2

Na2HgS2+H120 ~ NaOH+NaSH+HgS4

In a dilute solution of sodium sulfide both the cinnabar and stibnite are dissolved to form soluble thio-compounds; the mercury thio-compound, however, is hydrolyzed immediately to a mercuric sulfide precipitate. The mercury sulfide is removed

by filtration; the antimony-pregnant solution reportedly can be treated with carbon

dioxide or sulfur dioxide to recover a purified antimony sulfide (35). Zinc, lead,

tin, and other metals are reported to precipitate metallic antimony from solutions

of antimony compounds (35). This presents an alternate method of recovery if metallic antimony, rather than the sulfide, should be an economically preferable

commodity.

Discussion

This investigation has shown on a laboratory scale that leaching of cinnabarstibnite ores with sodium sulfide is a potential means of producing separate mercury

and antimony concentrates.

The limited scope of this investigation did not permit study of the precipitation of antimony from the leach solutions, although several possibilities have been

suggested. No attempt was made to investigate regeneration of the lixiviant, although regeneration would be desirable economically. The pregnant leaching solutions contain sodium hydrosulfide and sodium hydroxide in addition to the antimony

thio-compound. It appears within the realm of possibility that sodium sulfide for

recycling in the process could be obtained from the hydrosulfide and hydroxide.

Further work would be necessary to test the validity of these assumptions.

BIBLIOGRAPHY

1. BRADLEY, W. W. Quicksilver Resources of California. California State Mining

Bureau Bull. 78, 1918, 389 pp.

2. BRADLEY, WORTH. Method and Cost of Recovering Quicksilver From Low-Grade Ore

at the Reduction Plant of the Sulfur Bank Syndicate, Clearlake, Calif. Bureau

of Mines Inf. Circ. 6429, 1931, 17 PP.

3. BRAY, J. L. Nonferrous Production Metallurgy. John Wiley & Sons, Inc., 1947,

PP. 376-386.

4. CADY, W. M., WALLACE, R. E., HOARE, J. M. and WEBBER, E. J. The Central

Kuskokwin Region, Alaska. Geol. Survey Prof. Paper 268, 1955, 132 pp.

5. DENVER EQUIPMENT CO. Compania Mercurio en Sain Alto and Beneficiadora de

Mercurio. S A. Bull. M4-B24, July 1942.

6. DUSCHAK, L. H., and ACHUETTE, C. N. Fume and Other Losses in Condensing Quicksilver From Furnace Gases. Bureau of Mines Tech. Paper 96, 1918, 29 pp.
14

7. DUSCHAK, L. H. The Metallurgy of Quicksilver. Bureau of Mines Bull. 222,

1925, 173 Pp.

8. . Improvements in the Metallurgy of Quicksilver. AIME Tech. Pub. 264,

190, 37 PP.

9. FROLLI, A. W. Mining and Reduction Methods and Costs at the Oceanic Quicksilver

Mine, Cambria, San Luis Obispo County, Calif. Bureau of Mines Inf. Circ. 6950,

1937, 13 PP.

10. GAUDIN, A. M. Flotation. McGraw-Hill Book Co., Inc., 1932, pp. 338-342.

11. HALL, R. G., and BRADLEY, W. Concurrent Firing at Sulfur Bank and Reed Quicksilver Plants. AIME Tech. Pub. 1889, Metals Technol., December 1945, 11 pp.

12. HAYWARD, C. R. An Outline of Metallurgical Practice. D. Van Nostrand Co.,

3d ed., 1952, pp. 333-349.

13. KNAPP, L E. Report of the Governor of Alaska for the Fiscal Year 1881-1882,

P. 31.

14. LIDDELL, D. M. Handbook of Nonferrous Metallurgy. McGraw-Hill Book Co., Inc.,

1945, PP. 125-138, 473-516.

15. MENARDI, H. B. Reduction of Livingstonite Concentrate. AIME Tech. Pub. 1042,

Metals Technol., February 1939, 8 pp.

16. MERTIE J. B., Jr. Mineral Deposits of the Ruby-Kuskokwim Region, Alaska.

Geol. Survey Bull. 864-C, 1936, pp. 115-245.

17. MERTIE, J. B., Jr., and HARRINGTON, G. L. The Ruby-Kuskokwim Region, Alaska.

Geol. Survey Bull. 754, 1924, 129 pp.

18. MOOREHEAD, W. R. Methods and Costs of Mining Quicksilver Ore at the New Idria

Mine, San Benito County, Calif. Bureau of Mines Inf. Circ. 6462, 1931, 14 pp.

19. PETROF, IVAN. Report on Population, Industry and Resources of Alaska, 1884,

189 pp.

20. RANDALL, M., and HUMPREY, H. B. New Process for Controlling Mercury Vapor.

Bureau of Mines Inf. Circ. 7206, 1942, 10 pp.

21. RUTLEDGE, F. A. Investigation of the Rainy Creek Mercury Prospect, Bethel District, Kuskokwim Region, Southwestern Alaska. Bureau of Mines Rept. of Investigations 4361, 1948, 7 PP.

22. _ . Investigation of Mercury Deposits, Cinnabar Creek Area, Georgetown and

Akiak Districts, Kuskokwim Region, Southwestern Alaska. Bureau of Mines Rept.

of Investigations 4719, 1950, 9 pp.

23. REY, M. and BREVERS, H. Differential Flotation of an Arsenical Quicksilver

Ore. AIME Tech. Pub. 1264, Min. Technol., July 1941, 3 pp.

24. SCHEUTTE, C. N. Quicksilver. Bureau of Mines Bull. 335, 1931, 168 pp.

25. . Wuicksilver. Modern Uses of Nonferrous Metals. AIME, 2d ed., 1953,

pp. 318-332.
15

26. SEGURA, D. AIME Tech. Pub. 896, Metals Technol., February 1938, 7 pp.

27. SHAFFER, L. H., NEWTON, J., and FAHRENWALD, A. W. Flotation of Mercury Ores.

Eng. and Min. Jour., May 1940, p. 40.

28. SMITH, P. S., and MADDREN, A. G. Quicksilver Deposits of the Kuskokwim Region.

Geol. Survey Bull. 622-H, 1915, pp. 272-291.

29. SPURR, J. E. A Reconnaissance in Southwestern Alaska in 1898. Geol. Survey

20th Ann. Rept., Pt. 7, 1900, p. 261.

30. SPURR, J. E., and POST, W. S. Report of the Kuskokwim Expedition-Maps and

Descriptions of Routes of Exploration in Alaska in 1898. Geol. Survey, 1899.

31. SWINEFORD, A. P. Report of the Governor of Alaska for the Fiscal Year 1886-

1886, p. 949.

32. _ . Report of the Governor of Alaska for the Fiscal Year 1888-1888, p. 18.

33. TAGGART, A. F. Handbook of Mineral Dressing. John Wiley & Sons, Inc., 1945,

pp. 2-206 - 2-209.

34. THORNHILL, E. B. Recovery of Mercury From Amalgamation Tailing. AIME Bull.

104, 1915.

35. TREADWELL, F. P., and HALL, W. T. Analytical Chemistry. John Wiley & Sons,

Inc., vol. I, 9th ed., 1949, p. 119.

36. VAN ARSDALE, G. D. Hydrometallurgy of Base Metals. McGraw-Hill Book Co., Inc.,

1953, p. 29.

37. VON BERNEWITZ, M. W. Occurrence and Treatment of Mercury Ore at Small Mines.

Bureau of Mines Inf. Circ. 6966, 1937, 40 pp.

38. WALLACE, R E. Preliminary Report on the Quicksilver-Antimony Properties in

the Vicinity of Sleitmute, Georgetown District, Southwestern Alaska. Geol.

Survey, 1943.

39. WEBBER, E. J., and HOARE, J. M. Quicksilver-Antimony Deposits Near DeCoursey

Mountain, Iditarod District, Southwestern Alaska. Geol. Survey unpublished

report, December 1943.

40. WEBBEt, B. S., BJORKLUND, S. C., RUTLEDGE, F. A., THOMAS, B. I., and WRIGHT,

W. S. Mercury Deposits of Southwestern Alaska. Bureau of Mines Rept. of

Investigations 4065, 1947, 57 pp.

41. WELLS, R. R. Laboratory Concentration of Mercury Ores From Oregon, California,

Idaho, and Nevada. Bureau of Mines Rept. of Investigations 4620, 1950, 19 PP.

42. FEDERAL GEOLOGICAL SURVEY. Mineral Resources of the United States, 1882-1923.

43. FEDERAL BUREAU OF MINES. Mineral Resources of the United States, 1924-31;

Minerals Yearbook, 1932-53.

INT. -6U.OFMlINES.PGH. ,PA. 7575