MINERALOGICALVOLUME 56

MAGAZINE

NUMBER 384

SEPTEMBER 1992

The distribution of trace precious metals inminerals and mineral products
THE 23rd HALLIMOND LECTURE (1991)
L o u i s J. CABRI
Canada Centre for Mineral and Energy Technology, Ottawa, Ontario, Canada KIA 0G1
Abstract
DATA on the distribution of the precious metals (Au, A g and the platinum=group elements Ru, Rh, Pd,
Os, Ir, Pt) are rapidly accumulating as a consequence of the application of microbeam analyticaltechniques such as the electron microprobe, the proton microprobe, and the ion microprobe. The newdata obtained with these techniques build on knowledge accumulated over the last 140 years using bulk
analytical methods (e.g. spectrographic). The nature of the occurrence, the concentrations, and thevariation in amount s of precious metals in minerals and mineral products are reviewed with comments
as to the significance of the results with respect to industrial and environmental concerns. Future directions in mineral-specific analysis of trace precious-metal concentrations and applications outside ofprocess mineralogy are discussed.
KEYWORDS: precious metals, Au, Ag, PGE, microbeam techniques, electron-, proton- and ionmicroprobes, process mineralogy, mineral balances, mineral processing, extractive metallurgy,
environment.
Introduction
Ix is appropriate that this Hallimond lecture isconcerned with the distribution of trace precious
metals in minerals because the approach used
depends upon a good understanding of ore
mineralogy. Hallimond himself made important
contributions to ore mineralogy through studiesof the reflectance properties of minerals (Hallimond, 1956). It is interesting, moreover, to notethat some exploration geologists are beginning torealize the importance of ore mineralogy. The
statement by Canadian exploration geologists in amonograph on Archaean gold deposits is appropriate: ' . . . it became obvious in the preparationof this paper that the available data on ore
mineralogy and associated minerals were corn-

pletely inadequate to assess their significance.
While modern technology has rendered thisresearch "old fashioned", it may provide criticalinformation on ore paragenesis and have practicalapplications. This deficiency will be addressed by
the authors, and hopefully by other researchers',
Colvine et al. (1988). I hope to be able todemonstrate that effective application of modern
technology depends on a good understanding ofore mineralogy.
In an important review of minor element
concentrations in some sulphide minerals,
Fleischer (1955) pointed out that investigation of
the minor elements in sulphide ores in theprevious 100 years was aimed at finding newelements and new sources of some of the rarer
elements. He pointed out that the quality of data

Mineralogical Magazine, September 1992, Vol. 56, pp. 289-308
9 Copyright the Mineralogical Society

290

L . J . CABRI

obtained from these bulk analyses had steadily
improved over those 100 years, ranging from
qualitative to semi-quantitative, to quantitative
for most analyses in the previous 15 years. He
concluded that the emphasis had also shifted, thus
the aim of many of the recent investigations' . . . has been to obtain information about theprocesses of ore formation'. Fleischer also quoted
from Gavelin and Gabrielson (1947) 'From a
metallurgical point of view it is obviously of great
importance to know the amount and distribution
even of small concentrations of certain chemicalelements that may be of importance for therefining processes and which influence the character of the substances to be extracted from theore'. How perceptive of Gavelin and Gabrielson,
especially because commercial ores in those days
were usually coarser grained and richer in tenor
when compared to current ores; in addition,
environmental concerns did not receive the attention they do today.
What about the distribution of trace quantitiesof the precious metals, in particular, among
different minerals? The precious metals are
generally considered to include gold, silver andthe six platinum-group elements (ruthenium,
rhodium, palladium, osmium, iridium and platinum). These are well-known to occur as discrete
metallic alloys a n d as minerals, in which theprecious metal is an essential component. The
author has been involved, for a number of years,
in the application of several microbeam analyticaltechniques to determine trace quantities ofprecious metals in sulphides and oxides (Cabri
and Laflamme, 1976, with the electron microprobe; McIntyre et al., 1984, Chryssoulis et al.,1987 and 1989b with the ion microprobe; Cabri
et al., 1984 and 1985 with the proton microprobe). These trace quantities constitute the socalled 'invisible precious metals'.
It was just five years ago, for example, that theauthor pointed out the importance of several new
developments for in situ trace-element analysis
for the precious metals (Cabri, 1987). However,is is only recently that the significance of submicroscopic concentrations of precious metals is
being demonstrated through the application ofthese powerful new microbeam analytical methods to produce mineralogical balances for ores
containing different precious metals (Cabri and
Laflamme, 1984; Chryssoulis and Surges, 1988;
Chryssoulis and Cabri, 1990; Cabri, 1991).
Our understanding of the nature of the distribution of sub-microscopic precious metals, on theother hand, is not as far advanced as is our ability
to measure their concentration at very low levels.
For example, we do not always know whether

microbeam analyses represent finely divided'invisible' precious metal mineral grains less than
0.1 ~tm in diameter, colloidal-size particulates,
clusters of less than about 100 atoms, or true solid
solutions. Some information is available, in thecase of gold, through use of bulk analyticaltechniques such as t9VAu M6ssbauer spectroscopy(Wagner et aL, 1986) and use of relatively broadbeam spectroscopic methods such as X-Ray
Photoelectron Spectroscopy (Chryssoulis, pers.
commun., 1991). Being able to distinguishbetween, for example, gold in solid solution andcolloidal-size gold is important when preconcentration by sulphide flotation is under consideration as part of a metallurgical flow-sheet,because solid solution will be restricted to one or
few sulphides (e.g. arsenopyrite, pyrite), whereas
colloidal-size gold can occur in several minerals,
including gangue minerals (Chryssoulis, 1990).
Thus, the theme developed in this paper
follows my interest in the precious metals, beginning with the first analyses of platinum-groupminerals (PGM) by electron microprobe in
Canada (Cabri and Traill, 1966). This first
exposure to the PGM was accidental and aroseout of my interests in phase relations in the C u Fe-S system (MacLean et al., 1972; Cabri, 1973)
and Cu-Fe sulphide mineralogy (Cabri and Hall,
1972; Cabri et al. 1973). I had received a sample
of so-called 'cubic chalcopyrite' from Noril'sk(Cabri, 1967) which was eventually known as
talnakhite (Cabri and Harris, 1971). The electron
microprobe used in 1965-66, the first to be
manufactured commercially in North America(1961), was a relatively primitive instrument incomparison to today's models (Harris, 1990a).
That first introduction to the electron microprobe
and to P G M made me realize that: (a) little was
known about the PGM, and, (b) microbeam
methods would play an increasingly importantrole in the analysis of precious metals.

Microbeam techniques
The author's principal interest in microbeam
techniques in the early 1970s was in obtaining
quantitative analyses of PGM in order to understand the platinum-group-element (PGE) distribution in terms of discrete PGM. Indeed, the useof the electron microprobe has contributed to an
explosion in the number of new PGM species in
the last 20 years (Cabri, 1980) and to a better
understanding of PGM and P G E distributions(Cabri, 1981). The increase in the number of fully
characterized PGM from 1748, the year attributed to the discovery of platinum, to 1990 isshown in Fig. 1. Quantitative analyses of PGM in

TRACE PRECIOUS METALS

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FIG. 1. Histogram, updated from Cabri (1980), showing the increase in PGM species, from the discovery of platinumin Columbia in 1748, to 1990. An additional estimated 200 unnamed PGM are not considered. EPMA refers to thefirst application of EPMA to analysis of PGM by Genkin (1959). 1965 Review is that of Wright and Fleischer (1965),
1972 Review is that of Cabri (1972) and 1976 Review is that of Cabri (1976).

the author's laboratory were done using synthetic
standards and corrections to the microprobe data
were applied mainly using E M P A D R VII (Rucklidge and Gasparrini, 1969), but also using
M A G I C IV of J.W. Colby.
Though quantitative analysis of PGM was theprincipal goal, it became apparent that some of
the PGE must occur as trace constituents in the
more common sulphides and sulpharsenides.
Results were limited by the relatively high detection levels of the electron microprobes available;in those days a reasonable balance had to beachieved between stability and sensitivity for an
acceptable counting period. Minimum detectionlevels were of the order of 300 ppm for Pd, 400
ppm for Pt, and 500 ppm for Rh. The early
results, however, documented that sulpharsenides were carriers of P G E at concentrations
above these detection levels in some Cu-Ni
sulphide deposits in which PGE were a byproduct, as for example, the Kanichee deposit,
Ontario (Cabri, 1974) and the South Range
deposits of the Sudbury irruptive (Cabri and
Laflamme, 1976). On the other hand, the socalled sulphide-poor, PGE-dominant deposits
such as those at Stiltwater (Montana), Lac des Iles
(Ontario), and Merensky reef (Transvaal) con-tain measurable concentrations of P G E in supphides, particularly pentlandite (Cabri, 1981).
Unfortunately, satisfactory P G E mineralogicalbalances could not be achieved for many deposits
because of the relatively high detection levels byelectron microprobe (Cabri and Laflamme,
1984). An invitation by Ahmed El Goresy tospend several weeks in 1982 at the Max-PlanckInstitut fiir Kernphysik, Heidelberg, introduced
me to what was then the world's only mineralogically dedicated proton microprobe.
Operating conditions on the Max-Planck-Institut's proton microprobe at that time were far from
ideal, and the results obtained were a source ofboth great excitement, as well as frustration. It
was very exciting to be able to easily achieve
detection levels for many elements in the regionof 10 ppm, and with a bit more effort, in the rangeof 2-10 ppm. Particularly interesting were thevarious tests done to verify the results for bothaccuracy and precision (Cabri et al., 1984; Harris
et al., 1984; Remond et al., 1987). Eventually,
comparisons were made among results obtainedwith different proton microprobes and different
computer programs (Campbell et al., 1987). It
was frustrating, on the other hand, not to be able
to detect any P G E in Sudbury sulphides!

292

L . J . CABRIAt about the same time, I was exploring otheranalytical methods for the determination of trace
quantities of precious metals such as Secondary
Ion Mass Spectrometry (SIMS), also known as
the ion microprobe. The first studies relied onprior experience with the electron and proton
microprobes, using synthetic standards for reference (e.g. Cabri et al., 1985). Because silver was
an important commodity in the early 1980s, I
investigated silver distribution in galena andsphalerite by SIMS. The first published ionmicroprobe study of ore minerals was based onsynthetic standards: a series of Pb and Z n - F esulphides with varying silver contents (McIntyre
et al., 1984). The study demonstrated that theintensity ratios of the secondary ions of a major
element were proportional, for the element ofinterest, to the concentration of a trace element ina specific matrix. It became apparent, however,that this approach to quantification was notsuitable for ore minerals, primarily because of theeffects of substitutions between major elements(e.g. Fe for Zn) on ion yields. Fortunately, thisproblem was successfully addressed by Chryssoulis et al. (1986) using ion implantation (after Letaand Morrison, 1980) to determine silver insphalerite. An interesting development that arosefrom the investigations of the trace silver content
of major sulphide minerals at Brunswick Mining
and Smelting Corporation Limited concerned thedifferent results from electron microprobe analyses done on ore mined in 1977 and that mined in
1985. The 1985 results (involving 700 spot analyses) showed clearly a single galena population
with a median value of 0.074 wt.% Ag, (Laflamme and Cabri, 1986 a,b), which is about 10%
less than the earlier result of 0.084 wt.% Ag,
obtained with an older electron microprobe on
240 galena grains. The difference in results issignificant when doing a mineralogical balance(e.g. Chryssoulis and Surges, 1988), but in thiscase one cannot be certain whether the variation
is attributable to different instrumentation or to achange in ore.
The ion implantation technique has since beensuccessfully applied to trace gold analyses(Chryssoulis et al., 1987, 1989b), and is being
tested on the PGE (Chryssoulis et al., 1989a,
1991; Chryssoulis and Cabri, 1991). I look forward with a great deal of interest to talks byseveral speakers at this symposium because they
will be presenting new data on trace preciousmetal concentrations obtained by proton and ion
microprobes [See Abstracts o f the WinterConference].

Trace precious metal database

The precious-metal database that is presented
is derived largely from microbeam analytical datain the literature. Each precious metal is treatedseparately, with reference to analyses documented in the literature for identified deposits.
Though there are, unfortunately, many analysesof material from unidentified sources recorded in
the literature, it is felt that there are sufficient
data on samples from identified sources toprovide a realistic assessment of the concentration and range of values that may be expectedin different minerals for different deposit-types.
The trace-element data for different host'common' minerals are presented in the form oftables for each precious metal.
Minor- to trace-element substitutions occur
frequently in rarer minerals, which often are
precious-metal minerals, but these minerals are
not listed. Also, the deposits are not tabulated interms of type, but are grouped only by country;
further details for each locality may be found in
the references. It should be noted that data aretabulated in ppmw (~tg/g) as well as in wt.%.
Abbreviations for minerals and analytical methods are given in Table 1.
Gold. According to Harris (1990b), there are
only 26 valid gold-mineral species, 3 new gold
minerals whose descriptions await publication, 2
doubtful species and 13 incompletely cbaracterized, unnamed gold-containing compounds. The
distribution of gold in ores had been somewhat
controversial since the recognition of 'invisible
gold' (e.g. BOrg, 1930), and reliable mineralogicalbalances for gold in many ores were not possible
until recently (Chryssoulis and Cabri, 1990). The
data presented in Table 2 indicate that gold may
occur invisibly in many sulphide, sulpharsenide,
and arsenide minerals. The greatest concentration is attained in arsenopyrite, followed bypyrite, 1611ingite, tetrahedrite, and chalcopyrite.
The highest concentrations in pyrite werereported for a Columbian sample (up to3000 ppmw) and from Villeranges, France, (up to
600 ppmw) by Marion et al. (1991b) using SIMS
and EPMA, but with few analytical details.
Additionally, 0.05q3.46% Au, occurring as Aurich rims in pyrite from the Carlin-type Banqi
deposit (Guizhou, China) was determined by
EPMA; the gold occurs as submicroscopic native
gold inclusions, as deduced from chemical dissolution tests (Mao, 1991). High concentrations
obtained by E P M A of pyrite are 550-800 ppmw
(Picot and Marcoux, 1987) and 488ppmw(Swash, 1988), but few analytical details areprovided; hence these high values should beTRACE
Table

asp
bn
br
cp
cob
cb
ga

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=
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=
=

i. A b b r e v i a t i o n sarsenopyrite
bornitebreithauptite
chalcopyrite
cobaltite
cubanitegalena

AMS
BSE
EPMA
ICP-MS
HRTEM
LIMS
PIXE
SEM
STM
SIMS
SXRF
XPS
XRD

PRECIOUS
used

293

METALSin s u b s e q u e n t

gf= gersdorffite
gd= godlevskiteio= lollingitema= marcasite
mau = maucheritemh= mooihoekite
mk= mackinawite

tables

mt
nc
pn
po
py
tet
tk
tr

and text

=
=
=
=
=
=
=

magnetite
nickeline
pentlandite
pyrrhotite
pyritetetrahedrite
talnakhite
troilite

Accelerator Mass Spectrometry
Back Scattered Electron
Electron Probe Micro-Analysis
Inductively Coupled Plasma Mass Spectrometry
High Resolution Transmission Electron Microscopy
Laser Ionization Mass Spectrometry
Particle-Induced-X-Ray
Emission
Scanning Electron Microscopy
Scanning Tunneling Microscopy
Secondary Ion Mass Spectrometry
Synchrotron X-Ray Fluorescence
X-Ray Photoelectron Spectroscopy
X-Ray Diffraction

treated with caution. On the other hand, the highproportion of pyrite in many ore deposits often
makes even low concentrations of gold in pyrite
significant (e.g. <2 ppmw). Gold has also beenreported in magnetite from a skarn deposit at
Rossland, British Columbia (Wilson et al., 1990)
but more information is required for an evaluation of this type of trace-gold occurrence.
Whether the gold is structurally bound in every
example listed in Table 2 is not known. Many of
the analyses represent invisible colloidal-size
gold, possibly smaller than 20 nm in diameter(Chryssoulis and Cabri, 1990; Mao, 1991).
Detailed investigations using MOssbauer spectroscopy (e.g. Wagner et al., 1986), EPMA and
XRD (e.g. Johan et al., 1989), S I M S - E P M A H R T E M (Cabri et al., 1989), hydrothermal
synthesis (Wu and Delbove, 1989) and XPS
analyses (Chryssoulis, in prep.) collectively indicate that in many cases the gold is bound in thearsenopyrite structure, probably as Au 3+. A
MOssbauer spectrum of pyrite from Columbia, inwhich a large proportion of the gold is structurallybound was illustrated by Marion et al. (1991b).
Understanding the nature and distribution ofinvisible gold has been more of a concern to theprocess mineralogist and the metallurgist than tothe exploration geologist. The exsolution of gold
globules from pyrite (Birg, 1935) and arsenopyrite (Dunn et al., 1989) by heating at temperaturesof 500-700~ was indirect evidence of thepresence of invisible gold. It has long been
postulated, however, before confirmation of theoccurrence of invisible gold in sulphides, thatcertain epigenetic deposits contained gold-bearing arsenopyrite and pyrite during their formationin greenstone belts (e.g. Boyle, 1979). Subsequent recrystallization of sulphides during deformation and continued hydrothermal activity
caused much of the gold, silver, antimony andother elements to be exsolved from the earlysulphides, and redeposited in local dilatant zones
such as fractures in quartz lenses, local tension
fractures, and small faults in shear zones. It cantherefore be expected that auriferous deposits (orparts of deposits) which had been subjected to lessrecrystallization may contain a greater proportionof invisible gold.
It is becoming apparent from the database thatinvisible gold occurs in a great variety of geological environments and throughout geological time,depending on the extent and nature of metamorphic events. For example, the polymetallic Olympias deposit (Greece), of Late Cretaceous to
Tertiary age (Kalogeropoulos et at., 1989) contains 5.5 g/t Au, most of which is invisible. A
pyrite-rich concentrate containing 31.5 g/t Au
had 43.8% of the Au in arsenopyrite (av.
49 ppmw), 31.5% in As-rich pyrite (av.
49 ppmw), and 15.7% in As-poor pyrite (av.

294

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TRACE PRECIOUS METALS
3.4 ppmw) (Cabri and Chryssoulis, 1990). The
Cambro-Ordovician gold-beating As-Sb vein
deposits of the Massif Central (France), such as
Le Ch~telet, Viges, and Villeranges, are known
for their high contents of invisible gold in manysulphides, especially arsenopyrite (Picot andMarcoux, 1987). These authors suggested thatgold was originally concentrated in early pyrrhorite (at concentrations up to 30 ppm) from which
native gold later exsolved during alteration of thepyrrhotite to pyrite-marcasite. Picot and Marcoux also related the type of final gold mineralization to the type of fractures: in tight shear zones(e.g. Le Chfitelet) the transformation of pyrrhotite to marcasite-pyrite occurred early, coevallywith the deposition of arsenopyrite that attracted
and retained the gold released from the pyrrhotite. Deposits which had large structural openings, (e.g. Le Bournieux) contain no arsenopyrite, and hence the gold is released from thepyrrhotite in the native form.
Archaean gold deposits and volcanogenic massive sulphide (VMS) deposits in the Canadian
Shield generally contain smaller quantities ofinvisible gold than deposits in younger rocks. For
example, at the Trout Lake VMS deposit (Manitoba), it was estimated that 93% of the gold
occurs as A u - A g - H g alloys and that the balanceoccurs as invisible gold in pyrite (6%) andarsenopyrite (1%) (Healy and Petruk, 1990) Partof this study involved the tailings and showed that
26.7% of the gold was in pyrite, 8.9% in
arsenopyrite, and 64.4% as minute inclusions<1.5 ~tm (Chryssoulis, 1989).
Silver. In contrast to the other precious metals,
silver forms a large number of minerals: currently
131 species are recognised (Nickel and Nichols,
1991). There are, in addition, about 55 unnamed
or incompletely characterized silver-bearing
minerals, some of which may be new species.
Besides native silver, acanthite, and othercommon silver minerals, it is well-known thatsilver occurs in solid solution, substituting forcopper in many sulphide and sulphosalt minerals,
and for gold or lead in telluride and selenide
minerals (Thorpe et al., 1976; Gasparrini, 1984).
The significance of dilute solid solution of silver incommon base metal minerals (e.g. bornite, chalcopyrite, galena, pyrite, pyrrhotite, sphalerite) isquestioned by some. Gasparrini (1984), for example, opined that silver in solid solution in theseminerals is not common (at concentrations within
the detection limits of EPMA, which are given as
50-100 ppmw) and that ' . . . a large proportion,
if not all of the silver, occurs in discrete fineparticles which are not visible at average magnifications of 100x to 200x'. There is no doubt that295

there are published analyses of common minerals
in which the silver is erroneously ascribed to solid
solution. The silver contents may be partly or
principally accounted for by very fine grains ofsilver minerals, such as silver sulphosalts in galena(as discussed by Hall and Czamanske, 1972,
Laflamme and Cabri, 1986b, Knights and Patterson, 1988, and Foord and Shawe, 1989).
The complexity of having coexisting solidsolution silver as well as very small inclusions of asilver sulphosalt is described by Sharp et al.,(1990) in a study using scanning tunnelling
microscopy. They observed surface pits, 10300 nm in diameter, on cleavage faces, interpreted as due to plucking of diaphorite inclusions
during cleavage of the galena. Rod-shaped inclusions of diaphorite had previously been observed
to occur in these samples by H R T E M images(Sharp and Buseck, 1989). Sha W et al. (1990)
were also able to discern that the surface structureof galena has local distortions, which wereinterpreted to result from substitution of Ag + andSb 3+ for Pb 2+. In a related study (using EPMA,
BSE imaging and HRTEM), Sharp and Buseck(1992) have shown that in galena from La Paz andZacatecas (Mexico) there is coupled substitutionof Ag and Sb for Pb, as well as rod-shaped
inclusions of diaphorite (20 nm to 50~tm long) anddisk-shaped inclusions of frankeite. In othersilver-rich galenas, however, there seems to be abetter correlation between Ag and Bi thanbetween Ag and Sb (Jambor and Laflamme,
1978; Chen and Petruk, 1980; Laflamme andCabri, 1986a) as shown in Figs. 2 and 3. The
highest reported silver value for galena (9.4 wt.%Ag) is part of an extensive solid solution between
galena and ourayite from the Round MountainManhattan area (Nevada), the apparent solid
solution involves a coupling 2Pb 2§ ~ Bi 3+ +Ag +, with minor substitution of Sb for Bi (Foord
et al., 1988). Exchange between trace elements
and Pb in galena is probably more complicatedthan simple coupled substitution, and moreelements may be involved (e.g. Johan, 1988). For
example, Micro-PIXE analyses of carefully characterised galena from the Blagodatskoye deposit,
Transbaikal (Cabri et al., in prep.) showed avariety of trace elements (in ppmw): 6542 Sb,
6050 Ag, 582 Sn and 52 Cd (bismuth could not be
detected by EPMA, MDL 700 ppmw).
It is also apparent, however, that some careful
studies document silver solid solution as a significant contribution to the silver mineralogicaldistribution of base-metal: ores, whereas othercareful studies indicate that the silver contribution from the common base-metal minerals is
minor. For example, Chryssoulis and Surges

296

L . J . CABRIBMS GALENA 1986
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SILVER ATOMIC PROPORTIONS
FIG. 2. Silver versus antimony by EPMA in galena from Brunswick No. 12 mine, New Brunswick (Laflamme andCabri, 1986b).
(1988) concluded that 30% of the silver in the mill
circuits at Brunswick Mining and Smelting Corporation, New Brunswick, occurs as solid solution
m galena. This work was based on an average Ag
concentration in galena of 740 ppmw, obtainedby EPMA, with high silver values checked forminute inclusions at magnifications of over 3000 •(Laflamme and Cabri, 1986b). A lesser contribution to the silver mineralogical balance wasascribed to sphalerite, chalcopyrite, and pyrite(totalling 7.5%) by ion-microprobe analyses, andthe bulk of the silver was attributed to tetrahedrite, which contains from 0.1 to 36.5 wt.% Ag. In
another example, approximately 60% of thesilver in the 'A' ore at the Kidd Creek mine
(2800L), Ontario, occurs as a solid solution inchalcopyrite, with an average of 183 ppmw
obtained by Micro-PIXE analyses (Cabri, 1988b).
Two separate studies of the silver distribution at
Zacatecas (Mexico) arrived at exactly the same
figures for the contribution of silver due to solid
solution versus that due to silver mineral inclusions. Pearson et al. (1988) determined that thebulk of the silver was in solid solution in galena
and calculated that 29% of the silver recoveredfrom lead concentrates was due to silver mineralinclusions. Sharp and Buseck (1992) independently determined that 29% of the silver in
Zacatecas galena is due to silver sulphosalt
inclusions.
B6rub6 et al. (1985) determined by EPMA that
averages of 880 and 850 ppmw Ag occur in solid
solution in chalcopyrite and galena, respectively,
in tailings from the New Calumet mine (Quebec).
By using these values, they were able to calculatethat the bulk of the silver Was lost as Ag-bearing
tetrahedrite, rather than from the chalcopyrite
and galena, which contained only 6% and 9% of
the total silver in the tailings, respectively.
Thus, the author believes that the silver content
of base-metal minerals, though usually well below
1 wt.% for individual minerals, is frequently asignificant contributor to the silver mineralogicalbalance. This is especially true where the silver
occurs in the more common base metal minerals.
It is therefore essential to determine the silver
mineralogical balance in order to understand thesilver distribution, without any preconceived
notions. Naturally, care must be taken with traceelement analyses, and selection of appropriate
analytical techniques and methodologies has to beemphasized.

TRACE PRECIOUS METALS

297

BMS GALENA 1986
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SILVER ATOMIC PROPORTIONS

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FIG. 3. Silver versus bismuth by EPMA in galena from Brunswick No. 12 mine, New Brunswick (Laflamme andCabri, 1986a).
Palladium. Palladium is a component of mostPGM, and it also attains the highest traceelement concentrations of all the PGE in common
minerals (Cabri, 1991). The best-known concentrators of palladium are pentlandite (up to
4.6 wt.%), breithauptite (up to 18.9%), and theCo- and Ni-sulpharsenides (up to 5.5 wt.%), as
shown in Table 3. Also listed in the Table is a'pentlandite' from Stillwater that contains 8.612.1 wt.% Pd. Because this mineral contains
approximately one atom of Pd, it can be thoughtof as a new mineral, analogous to argentopentlandite. Other common sulphides, such as chalcopyrite and pyrrhotite, seem to carry trace quantitiesof palladium, but the database is insufficient forquantitative assessment at this time. Palladium
concentrations as high as 2920 ppmw in mackinawite are reported by Czamanske et al. (1992).
Maximum values reported for Pd in pentlandite(in ppmw) are: J-M reef (46000), Platreef(20000), Lac des Iles (6500), Medvezhy Creek(2540), Great Dyke (1990), Penikat (1800),
Merensky reef (1164). It should be noted, however, that some of these values represent anomalously high concentrations, e.g. the value for

Platreef pentlandite (Kinloch, 1982). In marked
contrast, some sulphide-rich Ni-Cu deposits containing PGE as by-products such as Sudbury andKambalda, probably contain less than 2ppmw Pd
in pentlandite. These sulphide-rich deposits,
however, have high concentrations of palladiumin arsenides and sulpharsenides (up to severalwt.%).
The palladium concentration in sulphideminerals seems to be related to, or controlled by
the geological environment. It is interesting tonote that the sulphide-poor PGE-dominantdeposits as well as some of the sulphide-rich
Noril'sk Cu-Ni ores contain variable, but significant quantities of Pd in pentlandite. Kinloch andPeyerl (1990) suggest that the variations in PGE
solid solution in base-metal sulphides in the
Bushveld deposits are due to genetic factors.
They postulate that so-called normal volatilebearing Merensky reef has low PGE in solid
solution in pentlandite and pyrrhotite because
there was sufficient time between liquidus andsolidus for crystallization of the PGM. In contrast, in the abnormal volatile-enriched Merensky
reef the liquidus temperature was lowered, result-

298

L.J.

CABRI

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299

TRACE PRECIOUS METALS

maximum of 14 ppmw Pd is present in pentlandite. These authors postulate that differences incrystallization history account for the variation in
the palladium content of pentlandite.
The precious-metal-bearing Kupferschieffer
deposits, which contain a concentration of heavy
metals at the base of a black shale unit, have beenstudied by EPMA and Micro-PIXE (Pryzbylowicz et al., 1990; Kucha, 1991). It is reported that
optically inclusion-free areas of thucolite containfrom <110 to 1000 ppmw Pd, apparently associated with As concentrations. The exact siting of
the Pd in this unusual geochemical setting
requires further investigation.
R h o d i u m and ruthenium. The databases forrhodium (Table 4) and for ruthenium (Table 5)show a distribution similar to that of palladium,
except for the expectedly smaller concentrations.
High concentrations occur in Co- and Ni-sulpharsenides (up to 10.5 and 3.7 wt%, respectively,
for rhodium; up to 3.3. and 2.1 wt.%, respectively, for ruthenium) and in pentlandite (up to
866 ppmw for rhodium and 79 ppmw for ruthenium). Whereas palladium concentrations inpentlandite from pothole Merensky reef ore arehigher than those for normal Merensky reef ore,
the opposite is found for the highest rhodium andruthenium values. Troilite in the Merensky reefcontains up to 65 ppmw Ru, but the database islimited for this occurrence, and is also limited foring in a compressed time-flame for crystallizationof PGM, thus allowing much more of the PGE toremain trapped as solid solutions in the basemetal sulphides (except for Pt which formed
eutectic textures of Pt-Fe alloys with the sulphides). PIXE data (Table 3) for pentlandite
from pothole reef compared to that for normal
reef appear to support the suggestion of Kinloch
and Peyerl (1990).
The ores of the Noril'sk area, which vary frommassive sulphide ore bodies to veins to disseminations, may shown high concentrations of numerous trace elements and a wide diversity of PGM(Genkin, 1968, 1991). The extent of solid solutions of PGE in Noril'sk ores is being reinvestigated (e.g. Chryssoulis et al., 1991a; Czamanske
et al., 1991, 1992), and it had been estimated thatthe majority of the PGE accumulated as solid
solutions in the main ore-forming iron-nickel
sulphides (Genkin, 1991). Czamanske et al.
(1992) show that correlation between palladiumin pentlandite and the Ni/Fe ratio in pentlandite
may be largely an artifact of considering mixed
sample populations; there is also no correlation
between palladium concentration in pentlandite
and the bulk palladium content of the ore. For the
Oktyabr'sky orebody, they show that pyrrhotitebearing ores containing 10-39 ppmw Pd have 68284 ppmw Pd in pentlandite, whereas in pyrrhorite-free ores containing 77-149 ppmw Pd, a
Table

4,

DepositClarahelle mill
Copper Cliff mill
Crean Hill
Creighton
Falconbridge
Frood
Inco concentrates
Namew Lake
Pipe
gottenstone
NellgreenJ-M reef
J-M reef (b)Merensky reef(c)Merensky reef(d)Merensky reef(e)Merensky reef(f)Merensky reef
Volspruit pyroxenite
Bushveld anorthosite
UG-2
Pechenga
Komsomolsky
Medvezhy Creek
Oktyabr'sky
Lower Tagil'sk
Yubdo
Overall range:

Concentrations
cp

of rhodium

determined

cob

gf

av.225<0.05%-2.4%<0.05%-1.5%<0.05%-3,3%<0.05%-1.53%

in c o m m o nmineralspnpo(ppmw and wt.%)tr

Method

Ref.

<9

EPMA
EPMA
EPMA
EPMA
EPMA
EPMA
PIXE
PIXE
EPMA
PIXE
PIXE/EPMA
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
EPMA
PIXE
PIXE
PIXE
EPMA
EPMA

5
4
3
3
3
3
8
12
4
i
6
8
i
2
I
I
i
13
12
12
12
I0
9
9
9
ii
7

<0.05%-0.58%
<1.9

<1.6
<5-8

<3.0

<0.05%-7.2%<13-17(a)<5-10

up to 2.7%

<4-12
<14-86351-685
250-266
21-361
7-55<6-86619-708
27-41
<5-6
<5-12

<5
<6
<6-7(a)<7-24(a)0.45%-10.5%<5-10(a)<5-30<54,0%(h)13.2%-14.1%(h)<0.05%-10.5%<6-13(a)

3.7%

<5
<5
<5

<2-24

<6
<5-10

<0.05%-3,7%<2-866<4<4-12(g)
<4

<3-13

<9

Notes: (a) Single analysis >MDL; (b) Minneapolis adit; (c) regional pothole reef, Spitzkop; (d) pothole reef, NW;(e) normal reef, NW; (f) normal reef, NNW; (g) monoclinie pyrrhotite; (h) see text for discussion.
References: i. Cahri (unpublished); 2. Cabri (1988a); 3. Cabri and Laflamme (1976); 4. Cabri and Laflamme (1981);
5. Cabri and Lafla~me (1984); 6. Cabri and Laflamme (1991); 7. Cabri et al. (1981); 8. Cabri et al.
(1984); 9. Czamanske et al, (1992); i0. Distler and Laputina (1979); ii. Genkin et al. (1974);
12, Paktunc et al, (1990); 13. S i e e t al. (1990).

300

L . J . CABRI

Table

5.

Concentrations

of

cp

cob

Deposit
Rottenstone
WellgreenJ-M reef (b)J-M reef (e)Merensky reef(d)Merensky reef(e)Merensky reef(f)M e r e n s k y reef(g)Merensky reef
Pechenga
Lower Tagil'sk
Komsomolsky
Medvezhy Creek
Oktyabr'sky
Overall range:
Notes:

ruthenium

determined

gf

in

pn

commonmineralspotr<9-15(a)
<5-10

<5-13
<14-47
50-79

<5

<6
<i0-65<5<6<6-19(a)

<7
<5-22
7-28
-5
1.65%-3.3%

1.65%-3.3%

6-7
8-12

2.1%

<5
<5
<5
<5-19

<5-25

to 2.1%

9.6%(b)
<5
<5-14
<5

<4
<4
<4-21(h)

<5-79

<4-25

(ppmw

and

wt.%)

Method

Ref.

PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
PIXE
EPMA
EPMA
PIXE
PIXE
PIXE

i
3
4
i
2
I
i
i
9
6
7
5
5
5

<10-65(a) Single analysis >MDL; (b) see text for discussion; (c) Minneapolis adit; (d) regional
pothole reef, Spitzkop; (e) pothole reef, NW; (f) normal reef, NW; (g) normal reef, NNW;(h) Hexagonal pyrrhotite

References:

I. Cabri (unpublished); 2, Cabri (1988a); 3. Cabri and Laflamme (1991); 4. Cabri et al.
(1984); 5. Czamanske e t a l . (1992); 6, Distler and Laputina (1979); 7. Genkin e t a l .
(1974); 8. Paktunc et el. (1990); 9. S i e e t el. (1990).

rhodium and ruthenium analyses in pentlandite.
Although it is, therefore, premature to draw
conclusions from these relationships, these data
are in agreement with the general conclusions ofmineralogists studying the Bushveld ores (e.g.
Peyerl, 1983) that virtually all the rhodium in the
Merensky reef occurs in solid solution withinbase-metal sulphides.
Czamanske et al. (1992) found that droplet ore
in picritic-taxitic gabbrodolerite at the Medvezhy
Creek deposit in the Noril'sk I instrusion contains
Rh (<5-28 ppmw) and Ru (11-12 ppmw) inpentlandite, and 11-12 ppmw Rh in monoclinic
pyrrhotite. They ascribed this contrast to other
Noril'sk-area ores as due to more rapid quenchingof this ore type. Chryssoulis et al. (1991a)reported orally that certain Noril'sk pyrrhotite
grains from massive and disseminated ores contain 2 ppmw rhodium. Genkin (1991) concluded
that nearly all of the rhodium and ruthenium in
the Noril'sk deposits occurs in solid solution inpentlandite and pyrrhotite. Cabri and Laflamme
Table 6.
DepositClarabelle mill
Copper Cliff mill
Copper Cliff South
Crean Hill
Freed
Pipe
Pechenga
Noril'sk
Yubdo
Duluth Complex (b)Overall r a n g e :(1984) concluded that 60-80% of the rhodium in asample from the Clarabelle primary mill (Sudbury) occurs in solid solution in sulpharsenides,
20-30% in hollingworthite, and 5% as a solid
solution in sperrylite. There is no doubt that
future SIMS analyses of Sudbury pentlandite andpyrrhotite will enable better mineralogicalbalances to be calculated for rhodium.
High rhodium and ruthenium concentrations
found in rare 'pentlandites' from Alaskan-type
deposits such as Lower Tagil'sk (Genkin et al.,1974) and Yubdo (Cabri e t a L , 1981) probably
represent a rhodium or rhodium-bearing ruthenian analogue of argentopentlandite, respectively. Genkin et al. (1974) proposed that these(Ru, Rh)-rich 'pentlandites' were formed duringserpentinization, with the P G E derived fromprimary PGM which were subjected to hydrothermal metamorphism. Interestingly, these minerals, which are known in association only with
Alaskan-type ultramafics, contain no palladium.
This contrasts with the common occurrence ofConcentrations of platinum determined in common minerals (ppmw and wt.%)

cob

~f

av. 330

av. 330
to 0.15%

mau

mt

ne

<0,04%-0.08%<0.04%-0,19%<0,04%-0,25%
max. 0.20%
0.91%
to 0.34%
0.06%-0,13%<0.04%-0.91% <0.04~-0.25%

0.06%-0.13%

0.ii%

pn

po

Method Ref,

EPMA
EPMA
EPMA
EPMA
EPMA
EPMA
EPMA
<0.05-2.0<0.1-2.0 SIMS
<0,04%-1.9%(a)
EPMA
EPMA

to 0.34% <0,04%-0,Ii% <0.05ppmw-l.9% <0.1-2.0

Notes: (a) see text for details; (b) massive sulphide samples.
References: I, Cabri and Laflamme (1976); 2. Cabri and Laflamme (1981); 3. Cabri and Laflamme (1984); 4. Cabri et al. (1981);
5, Chryssoulfs etal. (1991a): 6. Distler and Laputina (1979); 7. lwasaki et al. (1983).

3
2
I
I
I
2
6
5
4
7

TRACE
Table

7.

Concentrations

PRECIOUSMETALSof i r i d i u m a n d o s m i u m d e t e r m i n e d

301
in c o m m o n m i n e r a l s(wt.%)

Depositcob(It)cob(Os)gf(Ir)

gf(Os)

MethodRef.

Copper Cliff mill
Pipe
Peehenga

0.26%
0.19%
5.8%

0,15%-3.2%

0.23%

9.6%

EPMA
EPMA
EPMA

i
i
2

0.19%-5.8%

0,15%-3.25%

0.23%

9.6%

Overall range:

References: i. Cabri and Laflamme (1981); 2. Distler and Laputina (1979)

palladium, with subordinate quantities of rhodium and ruthenium, in pentlandite associatedwith sulphide-poor, PGE-dominant deposits andsome Noril'sk ores (Table 3).
Platinum. The database for trace platinum is
not well-developed (Table 6). The Co- and Nisulpharsenides and arsenides in sulphide-richCu-Ni deposits such as Sudbury and Pechenga, as
well as in the Duluth c o m p l e x , contain someplatinum but in much lower concentrations thanpalladium or rhodium. The relatively high platinum content listed under pentlandite (Table 6) isfor the possibly new rhodium-bearing, ruthenian
equivalent of argentopentlandite. Platinum wasreported to be present in magnetite (Cabri et al.,1981), but more work is necessary to verify thistype of occurrence because the magnetite isclosely associated with P t - F e alloys. New data on
platinum in pyrrhotite and pentlandite for some
Noril'sk samples were reported orally by Chryssoulis et al. (1991a). Pyrrhotite in disseminatedores contains from about 0.3 to 2 ppmw Pt,
whereas pyrrhotite from massive ores has lessthan 0.1 ppmw Pt. Similarly, pentlandite indisseminated ores has 0.2 to 2 ppmw Pt, whereas
pentlandite in massive ores contains less than
0.05 ppmw Pt.
Iridium and osmium. Very little is known about
the mineralogy of iridium and osmium. They
form well-known alloys with each other and withother PGE (Harris and Cabri, 1991), but only a
few other minerals, such as irarsite and erlichmanite, are known. Iridium and osmium havebeen reported to occur in sulpharsenides from
Pechenga and Sudbury (Table 7), but the distribution and concentration of these elements incommon minerals can only be inferred from bulk
assays.
Industrial products and environmental
considerations

The distribution of trace precious metals inindustrial products may be discussed in terms of
three principal types of products, categorizedaccording to the treatment that they receive, i.e.
mineral dressing, extractive metallurgy, or manufacturing. A brief survey of the first two categories follows.
Mineral dressing products. The products ofmineral dressing range widely in type, comingfrom anywhere in a mill circuit, but it is often the
trailings that are the focus of study. The distribution of trace precious metals in mineral-dressing
products is currently undergoing a revival, andquantitative mineral balances are possible
because of the application of sensitive microbeam
analytical techniques. For many gold-bearingores, the concern is with sulphide refractory ores;
i.e. ores which are not readily treated by conventional milling and cyanide extraction techniques.
Examples of the trace gold distribution in tailings
are those from the Trout Lake deposit (Manitoba) and the Olympias deposit (Greece), both
mentioned previously. For Olympias, knowledgeof the gold distribution and contrasts in concentrations of gold in As-rich and As-poor pyrite may
be relevant to a grinding strategy of maximum
exposure of the As- and Au-rich areas forselective dissolution or oxidation.
As discussed above, it is not always possible todistinguish between precious metals that are insolid solution or, say, those that occur as minutecolloidal-size particles. It is interesting to notethat Haycock (1937) concluded that, 'As regards
actual milling practice, it may be of little significance whether sub-microscopic gold is merely
very finely divided, colloidal, or in solid solution,b e c a u s e . . , the lower limit of visible gold falls
below the lower limit o f . . .
cyanidation'.
Though this statement is true, and in principle
also applies to silver and the PGE, there arecircumstances where this information is important. For example, preconcentrated gold-bearing
concentrates may have the gold restricted to one
or more sulphides such as arsenopyrite andpyrite, whereas colloidal-size gold can occur inseveral minerals, including gangue minerals(Chryssoulis, 1990). The positive correlation
observed between gold and arsenic in gold-

302

L. J. CABRI

bearing pyrite (Chryssoulis and Cabri, 1990) may
be an important consideration both for environmental and economic reasons.
A summary of research into the silver carriers
at the Brunswick massive sulphide deposit (New
Brunswick) is given in Cabri (1988b) and Chryssoulis and Surges (1988). Though some galena is
an important silver carrier, great care must be
taken to discern between areas containing finesulphosalt inclusions (e.g. diaphorite) and inclusion-free areas of galena (Laflamme and Cabri,
1986a, b; Sharp and Buseck, 1992).
Data for the PGE in sulpharsenides andarsenides from mills in the Sudbury area are given
in Tables 3, 4, 6 and 7. The distribution ofrhodium is perhaps less-well understood in ores
such as those at Sudbury (Kerr and Kipkie, 1985).
There is no doubt, however, that the occurrenceof rhodium in sulpharsenides and the close
association of these sulpharsenides with rarerhodium minerals, together with the high frequency of these minerals as small inclusions inpyrrhotite, is a matter of concern (Cabri and
Laflamme, 1984). On the one hand, sulphur
abatement requires reduction of pyrrhotite inconcentrates, but on the other hand, the highprice of rhodium makes maximum recovery
highly desirable.
Extractive metallurgy products. Refractory sulphide-rich gold-bearing ores have traditionallybeen treated by oxidative roasting of flotation
concentrates (as a single- or two-stage process,depending on arsenic content) to produce a
porous calcine that is more amenable to cyanidation (Demopoulos and Papangelakis, 1989). The
process involves the topochemical breakdown ofarsenopyrite and pyrite by gaseous diffusion of As
and S to produce pyrrhotite (Swash and Ellis,
1986). The pyrrhotite then recrystallizes, diffusion becomes grain-boundary controlled, andfurther oxidation of the pyrrhotite may produce
magnetite and hematite. Though oxidative roasting is a well established technology, it is considered less attractive than aqueous oxidation
processes due to environmental concerns anddemand for increasingly higher recoveries. The
latter may be affected adversely by changes inroaster temperature and feed composition, which
may result in incomplete gold liberation. Eachmetallurgical product, however, requires separate study. For example, most of the unleachable
gold in the cyanide-leach residue from a particular calcine was determined by SIMS analyses to becontained within roasted pyrrhotite (67 ppmw),
with only part of the gold occurring locked within
iron-oxide particles (Chryssoulis, 1990). The
location of As- and Au-rich zones on pyrite grain

margins in some deposits is also of interest to
metallurgical engineers. In a study done mainly
on refractory ore from the Kalgoorlie region, it
was found by SEM examination that only thesmallest gold grains were trapped within ironoxide minerals, representing an insignificant loss
(Dunn et al., 1989). On the other hand, theseauthors concluded that the coarser gold grains,
especially those derived from the alteration ofprimary tellurides, may be the source of gold
losses because they require longer leaching times
for total dissolution.
In yet a different product, a matte (140 g/t Au)
produced by smelting of a gold ore was examinedby 197Au MOssbauer spectroscopy. The spectra
were interpreted as indicating that most of thegold was chemically bound in troilite (Wagner
et al., 1988). The only SIMS data on smelting
products is a qualitative SIMS line scan showing
gold distribution across phases in a PGE-bearing
matte (Lindsay and Sellschop, 1988).
There is still research being done in testing new
approaches to roasting to make the process lessenvironmentally damaging, but most of the current work on refractory gold-extraction technologies involves hydrometallurgical methods(Demopoulos and Papangelakis, 1989). These
new approaches are chiefly aqueous-oxidationroutes: chemical oxidation, biochemical oxidation, and pressure oxidation. Little quantitative data are currently available on the trace
distribution in products from such processes.
Smelting of copper-sulphide concentrates andpyrometallurgical treatment of copper scrap
produces the bulk of the world's copper. In bothcases, many impurities report to the copper
anodes. PIXE and SIMS analyses of several
copper anodes indicated low levels of gold whichranged between 61ppmw (PIXE) and 716 ppmw (SIMS), demonstrating that the bulk of
the gold is in solid solution in the copper (Chenand Dutrizac, 1991).
Besides the classic vein and disseminated silver
deposits, much of the world's silver is recoveredfrom base-metal ores that, typically, yield copper,lead, and zinc concentrates, or combinations
thereof. Silver recovery from copper concentrates
is usually very good, probably because most of thesilver collects in the smelter matte and hence inthe copper anodes. Chen and Dutrizac (1990)
concluded that >95% of the silver occurs in asolid solution in copper anodes, based on EPMA
with values ranging from 0.03-%0.44wt.%.
Good silver recovery is also possible from leadconcentrate treated by conventional pyrometallurgical technology. However, in-plant hygiene
problems associated with lead-bearing dusts andTRACE PRECIOUS METALS
fines produced during conventional pyrometallurgical processing have provided an impetus forresearch into hydrometallurgical methods fortreating lead concentrates. The research has met
with mixed success, in part because silver recovery is very erratic (0-100%).
Silver losses from zinc concentrates are also
very erratic and can be 100%, depending on theextractive process used. Over 80% of the world's
primary zinc production is based on the electrolytic process, and this technology continues toexpand at the expense of older, less efficient zincprocesses (Temple, 1985). As zinc concentrates
treated by the roast-leach-electrolysis processcommonly contain 0--300 g/t Ag, the silver can bea valuable by-product of zinc processing (Dutrizac and Chen, 1990). A study of roaster products
from the Kidd Creek Division of Falconbridge(Ontario) showed a complex silver mineralogy
and distribution (Dutrizac and Chen, 1990). The
bulk of the silver occurs as <0.5 ~tm grains ofsilver metal, but silver is also an important
constituent of an abundant lead silicate phase (upto 1.37 wt.% Ag).
The recovery of PGE after milling of nickelcopper ores is largely done through pyrometallurgical-hydrometallurgical refining technologies,
usually by first concentrating the PGE in magnetic alloys. The recovery of PGE in Canadiannickel/copper smelters is generally greater than
95% (Cabri et al., 1988), and similarly highrecoveries are typical world-wide. The newly
operating process employed for Stillwater (Montana) Pd-Pt ore, electric furnace smelting
followed by upgrading with oxygen-blown converting, gives recoveries of better than 99%(Hodges et at., 1991). South African practice on
treatment of PGE-dominant ores is to produce a
matte containing all of the metals and subject it topressure leaching and refining, again with highrecoveries. There may, therefore, be little incentive to study the distribution of trace Pt and Pd insuch operations, especially as the bulk of the PGE
losses (which may be substantial) are thought tooccur during mineral-dressing operations (Newman, 1973). Neverthelesss, tracking the moreelusive PGE such as Rh and Ir in smelting,converting, or treatment of leach residues may
prove profitable. Little is known about SIMS
analyses of matte samples, except that Lindsay(1988) states that the Ni-poor matrix of Ni-Fe
alloys inhibits the ionization of most of the PGE.
Conclusions and future directions

The trace-precious-metal database is growing
rapidly, and it is now possible to predetermine the303

most likely carriers of several of these metals.
Much more data must be gathered, as moreminerals from more deposit types are studied
quantitatively, possibly changing our understanding of the relative importance of gold- and silvercarrier minerals for different geological environments. Carriers for platinum and rhodium will bebetter understood and mineralogical balances forthese elements will be achieved. All this accumulation of data will have implications for our
understanding of the mineral chemistry (e.g. by
HRTEM and STM) with possible spin-offs formineral processing and environmental sciencethrough advances in surface chemistry. By providing better mineralogical balances for precious
metals, the process mineralogist will be able toplay a greater and earlier role in the design andimplementation of processing flow-sheets.
Microbeam techniques capable of providing
quantitative analyses of all the precious metals attrace levels are attracting wide interest, especiallyin process mineralogy, because of the practicalapplications. It seems likely that greater interest
will be shown by the exploration and economic
geology fraternity as the potential of reliable
mineral-specific analyses at the ppmw and ppbw
levels becomes appreciated. As was noted by
Colvine et al. (1988) 'The absence of systematicstudies of these mineralogical associations of gold
is an obvious deficiency in the documenation ofgold deposits. Much of the problem . . . is thelack of precise, detailed mineralogical studies
completed within mine environments. Without anadequate data base, it is impossible to evaluate if
such mineral associations are of fundamental
significance to understanding how Archean lodegold deposits formed'. The application of the
proton microprobe to diamond exploration is one
example of the application of microbeam analysesof trace elements to mineral exploration (Griffin
et al., 1990) as is the non-destructive aspect ofMicro-PlXE analyses of fluid inclusions during
the study of hydrothermal deposits (Anderson
et al., 1989; Halden et al., 1991). Likewise, thepotential for studies of magmatic ore deposits
through large mineral-specific, trace-element
data bases, including the precious metals, has its
attractions (Czamanske et al., 1992). It can also
be envisaged that the combination of trace
precious-metal analyses by SIMS may be done on
the same mineral grains that provide stableisotope data, with obvious possibilities for new
insight in understanding hydrothermal deposits.
The complementarity of these methods must
always be kept in mind. Competition between
different research groups will continue and it isthose who apply the data who will make the final

304

L . J . CABRI

decisions on analytical methods. Price and availability are important considerations, but equallyimportant is the reliability of the analyses coupledwith an understanding of the ore mineralogy.
Data will increase from research groups withaccess to more unique trace-element techniquessuch as AMS and SXRF. Some newer methods
will be explored to solve particular problems
involving trace elements that cannot be done withmethods mentioned above, or that can be done
more cheaply. Some of these methods will
depend on laser applications such as LIMS (e.g.
Chryssoulis et al. 1991b) or other commerciallyavailable configurations such as laser ablation
ICP-MS (e.g. Broadhead et al., 1990; Jackson etal., 1990) which has attracted considerable interest due to the rapidity of analyses.
The use of scanning techniques for imaging
trace elements is also bound to have a dramatic
impact on future directions. Visual representation, combined with the quantitative documentation of the trace-precious-metal distribution, isvery attractive. For SIMS, imaging with a resistive-anode encoder shows substantial promise(Elphick et al., 1991; Marion et al., 1991a).
Similarly, the scanning-proton microprobe offers
much (e.g. Campbell et al., 1991). Each competing method will have to be carefully evaluated,mineral by mineral, with attention given toaccuracy, precision and a clear understanding as
to what each analysis represents. The future
potential is indeed great.

Acknowledgements
I am fortunate to have had over the years
numerous collaborators in my attempts to arrive
at a better understanding of precious metals andtheir distribution through ore microscopy, X-ray
diffraction, and the development and applicationof new microbeam analytical approaches, f wish
to mention, in particular, present and former
colleagues (at was then the Mines Branch) such as
T. T. Chen, John Dutrizac, Syd Hall, Don Harris,
John Jambor, Gilles Laflamme, Rolando Lastra,
Ernie Nickel, Bill Petruk, John Rowland, Jan
Szymanski, and the late John Stewart. In addition, I wish to record my appreciation of theinteraction with colleague investigators Ahmed
E1 Goresy, Rainer Nobiling, and Kurt Traxel(Max-Planck4nstitut ffir Kernphysik, Heidelberg), Tony Naldrett and John Rucklidge(University of Toronto), Dick Alcock (Inco),
Doug Scott (Kidd Creek, Falconbridge), David
Carson (Noranda), Zdenek Johan and Guy
Remond (BRGM, Orl6ans), Stephen Chryssoulis
and Stuart McIntyre (Surface Science Western),Andrew Clark, Alan Criddle and the late Max
Hey (The Natural History Museum), Iain Campbell, John Maxwell, and Bill Teesdale (Universityof Guelph), Soey Sie and Chris Ryan (CSIRO,
North Ryde), Roger Eckstrand, Larry Hulbert,
Dogan Paktunc, John Scoates, and Bob Traill(Geological Survey of Canada), Keiko Hattori(University of Ottawa), Graham Nixon (British
Columbia Geological Survey), Gerry Czamanske, Mike Fleischer, and Mike Zientek (USGeological Survey), J o h a n de Villiers (MINTEK), Clive Feather (Anglo American ResearchLaboratory), Euan Kinloch and Wolf Peyerl (JCI
Mineral Processing Research Laboratory), Peter
Buseck and Tom Sharp (Arizona State University), Alexandr Genkin ((IGEM, Moscow), Tim
Bejamin and Pam Rogers (Los Alamos National
Laboratory), and Mark Rivers and Joe Smith(Brookhaven National Laboratory/University of
Chicago). All of those mentioned above havegenerously contributed to my understanding andto my continued interest in the subject. Most, but
regrettably not all of our collaborations resulted
in joint publications.

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[Manuscript received 14 January 1992]