CHAPTER 3: LITERATURE REVIEW

3.1 INTRODUCTION

This section contains a review of the platinum group minerals and BMs associated

with the UG2 Reef of the BC. This includes an appraisal of the mineralogy of

chromium to include spinel structures, a review of the concept of solid solution

and details how PGEs and BMs values are concentrated from UG2 ore. In

addition, a literature review of available analytical techniques and instrumentation

for the determination of the physical and chemical properties of UG2 Reef

minerals is included.

3.2 GEOLOGY OF THE BUSHVELD COMPLEX

The Bushveld Complex is the largest known igneous system ranging in

composition from ultra-mafic to felsic [18]. The BC is unique both in terms of its

size, covering a total estimated area of 66 000 km2 and of its mineral values. The

mafic rocks of the BC host layers rich in PGEs, chromium, nickel, titanium and

vanadium. In practice, significant quantities of nickel, copper, cobalt and gold are

extracted as by-products of platinum mining, which contribute both towards

meeting Operation costs and enhancing mining revenues [19].

The Rustenburg Layered Suite of the BC is divided vertically into five zones,

being the Marginal Zone, the Lower Zone, the Critical Zone, the Main Zone and

the Upper Zone [18]. The chromitite layers of the BC occur within the lower and

upper Critical Zone in three distinct groups, namely [20]: the lower group (LG),

middle group (MG) and upper group (UG). The LG6 layer is the most important

of those mined in the lower group for chrome. The upper group which includes

the UG1 and UG2 chromitite layers, contained in norite, anorthosite and

pyroxenite, hosts chromite deposits plus the PGM-bearing Merensky and UG2

Reefs. The differing pyroxenite and chromitite layers as well as zones of the

western BC are shown in Figure 3.1 and Figure 3.2.  

 

12  

 

Figure 3.1 A schematic of the western Bushveld Complex showing the lay-out

of the Merensky and UG2 Reefs’ horizontal layers plus the commercially

important chromitite layers [18]. 

13  

UG2 Chromitite layer

Figure 3.2  Pyroxenite and chromitite layers of the UG2 Reef (Photographed by

Anel McDowell, 2007).

3.3 MINERALOGY OF THE UG2 REEF

3.3.1 General mineralogy

According to McLaren and De Villiers [21] the UG2 Reef consists primarily of

chromite (60 -90% by volume), orthopyroxene (5-25%) and plagioclase (5-15%).

Secondary or accessory minerals also form part of the UG2 Reef and are listed in

Table 3.1.

Table 3.1 Secondary or accessory minerals of the UG2 Reef

Minerals Composition

Clinopyroxene Ca, Mg, Fe, Na, Li silicates Base metal sulphides and other sulphides

Chalcopyrite (CuFeS2), Pyrrhotite (FeS), Pyrite (FeS2), Pentlandite ((FeNi)S)

PGMs Refer to Table 3.2. Oxides Ilmenite (FeTiO3), Magnetite (Fe3O4),

Rutile (TiO2) Micas K, Li substitution

in K2O.Al2O3.6SiO2.2H2O Secondary quartz SiO2Chlorite (MgFe)5Al(SiAl)O10(OH)8 Serpentine Mg3Si2O5(OH)4Adapted from various sources [21 – 24]

14  

3.3.2 Mineralogy of platinum group minerals

There are 109 PGM species recognised by the International Mineralogical

Association (IMA), which include sulphides, tellurides, antimonides, bismuthides,

arsenides and alloys to native species [22]. The most common base metal sulphide

minerals associated with PGMs which are mainly found in the Merensky Reef are

pyrrhotite (FeS), chalcopyrite (CuFeS2), pentlandite ((FeNi)S), cobaltian (CoNiS)

and pyrite (FeS2 ) [22, 23].

According to a mineralogical investigation completed by Penberthy, Oosthysen

and Merkle, [24] the most commonly found platinum group sulphide minerals

found in the UG2 chromitites are laurite, cooperite, malanite and braggite as

highlighted in Table 3.2. Other resources revealed non-sulphidic minerals, such as

antimonides, arsenides, bismuthides, tellurides and alloys, [18,21,24,25] which are

also listed in Table 3.2. Common substitutions are identified for some of the

minerals by Vermaak [25].

Table 3.2 Mineral PGM speciation of the UG2 chromitites

PGM Mineral Species Mineral Composition Common

substitution Antimonides Geversite PtSb2 -

Stibiopaladinite -

Arsenides Sperrylite PtAs2 Rh,Ir,Sb,S

Bismuthides Insizwaite PtBi2 -

Sulphides Cooperite PtS Pd,Ni

Laurite Ru,S2 Ir,Os

Braggite (Pt,Pd)S Ni Malanite (Pt,Rh,Ir)2CuS4 - Rh sulphides - Vysotskite PdS Pt,Ni

Tellurides Merenskyite PdTe2 - Monocheite PtTe2 -

Alloy Isoferroplatinum Pt3Fe Ru,Rh,Ir,Pd,Os,Cu,Ni

McLaren and De Villiers [21] reported that accessory minerals, such as PGMs were

observed as discrete grains which are associated with base metal sulphides (33-

15  

85%), along grain boundaries (6-57%), in silicates (2-29%) and in chromite

(<12%). The grains identified were small and were between 1 and 4 μm in size.

McLaren and De Villiers [21] referred to chromite as a solid solution of 98% spinel

and 2% ulvöspinel.

Others suggested that there are also non-disclosing platinum group components

present as sub-microscopic particles or dilute solid solution, within minerals. [18]

Unfortunately, since these sub-microscopic particles are generally below the

detection limit of even the most sophisticated electron microscopic

instrumentation, there is limited information available on these particles and dilute

solid solution.

3.3.3 Mineralogy of chromium

Although some fifteen chromium minerals are known, only one is of commercial

importance as a source of chromium. The mineral is chromite, which has the

theoretical composition Fe.Cr2O4 containing 68 percent chromic oxide (Cr2O3). [27] The term chromitite is used to indicate the layer in which chromite

predominates. Figure 3.3 shows a sample which was obtained from the Critical

Zone, Upper Group 2 (UG2) chromitite layer which clearly exhibits chromitite

crystals.

Figure 3.3 Chromitite crystals in the UG2 Reef. (Photographed by Anel

McDowell, 2007)

16  

Figure 3.3 would suggest that the mineral chromite is iron-black in colour,

however in nature, this varies even to dark-brown. Chromite has a sub-metallic to

metallic lustre, a hardness of 5.5 to 6.5 and a relative density of 4.5 to 4.8.

Depending on its purity, the melting point may range from 1545 - 1730 °C. [25,26]

These physical properties all contribute towards the challenge associated with the

smelting of UG2 Concentrate as referred in Chapter 1.

3.3.4 Chromium spinel structures

Chromite belongs to the spinel group of minerals which have the common

formula X2+ Y23+ O4, in which X2+ represents the divalent cations of magnesium,

zinc, manganese and ferrous iron and Y3+ represents the trivalent cations of

chromium, manganese, aluminium and ferric iron [31].

The spinel group may be divided into three series, according to whether the

trivalent ion is aluminium, iron or chromium [27, 28] and which are listed in Table

3.1.

Table 3.3 Spinel series [27, 28]

Spinel series (Al) Magnetite series (Fe3+) Chromite series (Cr)

Spinel (Mg.Al2O4) Magnesioferrite (Mg.Fe2O4)

Magnesiochromite (Mg.Cr2O4)

Galaxite (M n.Al2O4)

Jacobsite (Mn.Fe2O4) Ferrochromite (Fe.Cr2O4)

Hercynite (Fe.Al2O4) Hausmannite(MnMn2O4)

Gahnite (Zn.Al2O4) Franklinite (Zn.Fe2O4) Magnetite (Fe.Fe2O4)

Trevorite (Ni.Fe2O4)

Ulvöspinel (FeFeTiO4)

The simplest mineral of the series is MgAl2O4 (or MgO.Al2O3), which is more

commonly known as spinel. Natural chromite has the ideal composition FeCr2O4

(or FeO.Cr2O3), but other forms of chromium containing spinels also exist, such

as picochromite (MgCr2O4).

17  

Chromite, crystallizes with a spinel structure. Large oxygen ions are closely

packed in a cubic framework of 32 oxygen anions with 24 cations in the

interstitial sites between them. This structure was first investigated by Bragg and

Nishikawa who were of the opinion that there are 8 tetrahedral A sites in four-fold

coordination and 16 octahedral B sites in six-fold coordination. [26, 27]

The difference in distribution of the cations between the A and B sites results in

two structural types of spinel. With the general formula X2+8 Y2 3+

16 O32 the two

structures are:

Normal 8X2+ in A, 16Y2 3+ in B

Inverse 8X3+ in A, 8X2+ + Y2 3+ in B

Normal spinels have divalent cations in the tetrahedral A sites and trivalent

cations in octahedral B sites e.g. Fe2+(Cr23+)O4. Inverse spinels have 8 of the 16

trivalent cations in the tetrahedral A sites, and the octahedral B sites are occupied

by eight divalent and eight trivalent cations e.g. Fe3+(Fe2+Fe3+)O4.

There are many spinels in which the distribution of cations lies between the two

extremes of normal and inverse structures. Figure 3.4 is an example of an unit cell

of a spinel structure.

Figure 3.4 Unit cell of spinel [33]

18  

3.3.5 Compositional variation in minerals and the concept of solid solution

Chromite is given the general formula (Mg, Fe)O.(Cr, Al, Fe)2O3 in which the

proportions of Mg2+, Fe2+ and Cr3+, Al3+, Fe3+ may vary considerably. [17,32] This

means that chromite can vary in composition between different deposits and

within any particular deposit, due to partial substitution of iron by magnesium and

chromium, to a varying extent by aluminium and ferric iron.

Crystal theory explains composition variation in a mineral as a result of

substitution in a given structure of an ion or ionic group, for another ion or ionic

group. This process is known as ionic substitution or solid solution and occurs

among minerals that are iso-structural. [29] The term solid solution can be defined

as follows: It is a mineral structure in which specific atomic site(s) are occupied in

variable proportion by two or more different chemical elements (or groups).

The main factors which determine the amount of solid solution taking place in a

crystal structure are: [30, 31, 34]

• The comparative size of the ions, atoms, or ionic groups substituting for

each other. Substitution between ions can only happen when the difference

in ion size is less than 15%. Substitution becomes limited when the

difference in ion size is between 15 – 30% and is unlikely when the

difference is greater than 30%.

• The charges of the ions involved in the substitution. Charge neutrality

must be maintained in any substitution mechanism. If the ion charges are

the same, then the structure shall be electrically neutral. If the charges are

not the same, additional ionic substitution must take place elsewhere in the

structure to maintain charge neutrality.

• The temperature at which the substitution takes place. The extent of ionic

substitution of dissimilar-sized cations is greatly encouraged at higher

temperatures and decreases at lower temperatures. As a result a greater

variability in the composition of a structure is found at higher

temperatures.

19  

3.3.6 Association of PGEs in chromite as solid solution or inclusions

There are a series of published reviews regarding the occurrence of PGEs within

the UG2 Reef or more specifically within chromitite crystals. On initial review

some of the more contradictory statements found were: [35]

• PGEs may be accommodated to some extent in solid solution in the

chromite spinel lattice, which may be a collector of PGEs at high

temperature.

• Chromites in the Merensky and UG2 Reefs contain no PGEs in solid

solution and all PGEs can be accounted for by the presence of minute

inclusions.

• Solid solution of platinum group elements such as osmium , ruthenium

and iridium in chromite is a distinct possibility.

Chromite, is an example of fractional crystallisation, formed through slow

cooling, convection, crystal sorting, diffusion, compaction and inputs of new

batches of magma. [36]

Further literature reviews supported the following regarding PGM mineralisation

in chromitite crystals: [35 – 38]

• PGEs are contained in discrete platinum group minerals or in solid

solution in sulphides.

• Minor sulphide minerals, predominantly laurite (RuS2), are enclosed

within or at the grain boundary of base metal sulphide grains.

• A small percentage (5-12 %), of PGMs are found, predominantly laurite

(RuS2), enclosed, within chromite.

• Pt-(Pd-Au), occasionally occurs as discrete metal-alloy inclusions.

20  

An unpublished article found on solid solution theory which attempts to explain

the accommodation of PGEs in the chromitite crystal lattice was written by Thom. [39]

According to Thom, the ionic radius of the chromium ion in a chromite spinel is

0.62 ° A, which indicates that any trivalent cation with a co-ordination number of

6 and ionic radius of up to 0.74 ° A can replace chromium. This would include

Rh, Ru and Ir and possibly Pd. He further postulates that divalent Pd may replace

divalent magnesium in a spinel as its ionic radius is 19% bigger than that of Mg.

Although divalent Pt with ionic radius of 0.80 ° A can also replace Mg in spinel, it

cannot replace Cr or Al in spinel as Pt is not trivalent.

Thus, according to crystal theory and postulated by Thom, [39] the platinum group

elements Rh, Ru and Ir may form part of the crystal lattice of chromite found in

the Merensky and in particular the UG2 Reef.

3.4 MINERAL PROCESSING OF PLATINUM ORE

3.4.1 The recovery process for precious and base metals from Merensky and

UG2 ores

The PGM and base metals in Merensky and UG2 ores are extracted and separated

from the gangue by mineral processing of the ore to produce a Concentrate

containing typically 100 – 300 g t-1 PGM. There are two fundamental operations

in the mineral processing of ores. The valuable minerals are released or liberated

from the waste gangue minerals and then they are separated by the process of

concentration.

Liberation of the valuable minerals is accomplished by comminution, which

involves crushing and if necessary grinding to such a particle size that the product

is a mixture of relatively discrete particles of mineral and gangue. Over grinding

is however not favourable as it can lead to the production of very fine untreatable

“slime” particles which may be lost with gangue to tailings. Grinding is therefore

frequently a compromise between the grade of Concentrate achievable and the

potential loss of fine minerals [40].

21  

Concentration of the PGE values in the ore involves discrete sulphide froth

flotation of the crushed or ground ore exploiting the differing surface properties of

the minerals. The extent to which froth flotation is successful is influenced by the

degree of affinity of the minerals for air bubbles within an agitated pulp. The

conditions of the pulp are adjusted by adding reagents which make the valuable

minerals aerophilic and the gangue minerals aerophobic. Thereby the valuable

minerals are joined to the air bubbles and form a froth which floats on the surface

of a pulp which contains the non-floating gangue minerals.

In practise, complete liberation is seldom achieved even if the ore is ground down

to the grain size of the desired mineral particle as the particle containing the

mineral may also contain a portion of the gangue. Particles of “locked” minerals

and gangue are known as middlings and further liberation from this fraction may

only be achieved by further comminution. The “degree of liberation” refers to that

percentage of the mineral which occurs as free particles in the ore relative to the

total content. Quantification of liberation is now possible using dedicated

instrumentation like MLA and QEMSCAN and Concentrators are increasingly

using such systems to continuously monitor degrees of liberation [41].

The mineralogy of the UG2 ore is significantly different to that of Merensky ore

and as such, Concentrator design requires customisation to accommodate the

unique challenges of floating the PGE minerals contained within UG2 ore. The

more significant differences between UG2 and Merensky ores include:

• The much higher chrome content of the UG2 ore where the challenge is to

produce Concentrates with low enough chrome content so as to not present

undue problems for the downstream smelting process as referred in

Chapter 1.

• A grain size of PGMs of 2-5 μm within UG2 ore when compared to the

much larger grain size of the PGMs associated with base metal sulphides

in the Merensky ore. The challenge is to achieve commercially acceptable

PGMs recoveries from UG2 ores when it is generally accepted that

recovery by flotation is reduced as the grain size is reduced.

22  

These characteristics resulted in the development of mill-float-mill-float circuits

(MF2) for the processing of UG2 ores with screens as classifiers rather than

cyclones. In this approach, the initial grind of the UG2 ore is coarse in order to

maximise chrome recovery during primary flotation and the secondary grind is

much finer to liberate the PGMs grains occurring interstitially within the chromite

and thereby enhance PGMs recovery.

The aim of mineral processing, regardless of the process used, such as that in

Figure 3.5 is to, separate the minerals into two or more products with the value in

the Concentrate, the gangue in the tailings and “locked” particles in the middlings.

Such separations are not perfect however, fine liberated valuable particles may

report to the middlings and tailings, as technology for treating such fine-sized

minerals requires further development. Concentrator tailings still contain valuable

minerals and companies like Sylvania as indicated in Chapter 1 have begun to

custom design Concentrators for the retreatment of platinum mine tailings dams.

Ore

Primary grind

Pre-concentration

Middlings Tailings

Re-grind

Seperation

Concentrate Middlings Tailings

Figure 3.5 A mineral processing layout utilising two-stage separation [40].

Recovery as referred above, in the case of a metallic ore is that percentage of the

desired metal which is contained in the ore that is recovered in the Concentrate.

By way of example, a recovery of 90% may be interpreted as 90% of the metal in

23  

the ore being recovered in the Concentrate with 10% lost to the tailings. The grade

or assay refers to that content of the Concentrate which shall be processed to a

marketable end product.

3.5 ASSAY METHODS USED FOR PLATINUM GROUP METALS

3.5.1 Commonly used assay techniques

Literature refers to many analytical techniques for the determination of precious

metal concentration in ore. The determination of these metals is specialised and

complex because of the close similarity of their chemistry and the low

concentrations at which they occur [42]. Since PGEs occur in ores at very low

concentrations, typically in micrograms per gram (μg g-1) or nanograms per gram

(ng g-1) range, a pre-concentration step is necessary to facilitate their quantitative

determination. The most commonly known analytical techniques in the platinum

industry for collecting and pre-concentrating PGEs are that of lead fire assay and

nickel sulphide collection.

3.5.1.1 Lead fire assay collection technique (FA-Pb)

Lead fire assay collection dates back many centuries and is still one of the most

reliable methods for analysing ores which contain precious metals [1]. An

advantage of the fire assay technique is that it allows the use of large quantities of

sample, from which trace amounts of precious metals may be concentrated which

in turn minimizes any sampling error which may arise from the occurrence of the

precious metals as discrete particles [42].

During the Pb-FA method, PGEs are extracted from the sample by a reductive

fusion and collected in a lead button [43]. Heat and suitable fluxes are used to

separate the precious metals from gangue (waste material) in the ore. Typically

the sample is blended with a flux mixture comprising litharge (lead monoxide),

silica, sodium carbonate, borax and carbon as a reducing agent (usually in the

form of flour or charcoal). The flux mixture converts compounds which are

infusible at a certain temperature into compounds which melt at desired

temperatures [44]. The composition of the flux is critical as it is customised

24  

according to the sample and as such some knowledge of the ore type and an

understanding of the principles of pyrochemistry is necessary. Principally, an ore

with an acidic component will require a basic flux, whereas an ore with a basic

component will require an acidic flux [1,44].

The blended sample with flux is transferred to crucibles which are loaded into

preheated furnaces and fused at about 1100 °C for one hour at which point the

slag becomes fluid and the litharge in the flux is reduced to fine droplets of lead.

This encourages the collection of the precious metal particles in the lead button

which then settles at the bottom of the crucible. The molten slag and lead are then

poured into an iron mould and cooled. After cooling, the molten charge settles

into two distinct layers, a top layer of slag and a bottom layer consisting of a lead

alloy button. The slag is removed from the lead button by tapping and the lead

from the lead button is removed by a process called cupellation. The lead button

is placed in a shallow porous dish called a cupel, which is made of bone ash or

magnesium oxide. The cupel is placed in a muffle with an oxidizing atmosphere,

which is heated to between 900 - 1000 °C. Most of the lead is absorbed as lead

monoxide during the heating stage and leaves a silver bead also known as a prill

behind, which contains the precious metals.

These prills are exposed to high temperature cupellation at approximately

1300 °C for several hours to remove impurities through the addition of small

quantities of pure lead. Most analytical methods proceed from the prill stage,

either by obtaining the prill mass or analysing the solution after dissolution of the

prill for the elements platinum, palladium, rhodium and gold by AAS, ICP-AES

or ICP-MS [42].

Although this method is rapid and sensitive for low concentrations, the precious

metals are subject to losses which vary such that 98-100%(m/m) of the osmium

and ruthenium are volatilised as tetroxides, and 50-60%(m/m) of iridium and 0 to

10%(m/m) of platinum, palladium, rhodium and gold are lost [42]. The Pb-FA

collection process is shown in Figure 3.6.

25  

a)  b)

c) d)

e) f)  Figure 3.6 The Pb-FA process consists of the following steps: a and b) pouring,

c and d) de-slagging e) cupellation and f) hammering of prills. (Photographed by,

Impala laboratory management, 2006)

26  

3.5.1.2 Nickel sulphide fire assay collection technique (NiS-FA)

An alternate method which compliments the Pb-FA method for the analysis of

precious metals is that of nickel sulphide collection. It is known that the precious

metals are frequently associated with base metal sulphides, therefore making NiS-

FA an effective collector of precious metals. Robert et.al [45] developed successful

methods of fusing sample and flux mixtures containing nickel carbonate, sulphur,

borax and sodium carbonate to maximise collection of the individual PGEs.

The sample is fused at temperatures between 1100 °C to 1300 °C for about an

hour at which point the nickel sulphide melt shall collect the PGEs and gold. The

melt is then poured into iron crucibles and allowed to cool down. The nickel

sulphide button is then removed from the slag and the slag is either discarded or

retained for a second fusion. Any losses which may occur during the initial fusion

are recovered during a second fusion. Nickel sulphide buttons are then crushed or

ground and treated with a mixture of hydrochloric acid and ammonium chloride.

This dissolves the base metal matrix and the insoluble platinum group metal

sulphides are collected by filtration. The black precipitant is dissolved in acid and

the final solution analysed by spectroscopic techniques. The NiS-FA collection

process is shown in Figure 3.7.

a) b)

27  

c) d)

e) f) Figure 3.7 The NiS-FA process consists of the following steps: a and b) de-

slagging of NiS buttons, c) digestion d) filtering, e and f) microwave digestion.

(Photographed by Impala laboratory management, 2006)

3.5.1.3 Tellurium (Te) co-precipitation in conjunction with NiS-FA

Te co-precipitation is used to pre-concentrate trace amounts of PGEs from acidic

solutions and in doing so to separate the PGEs from the matrices which would

interfere seriously with determination by ICP-OES and ICP-MS. This pre-

concentration procedure is not new and was investigated as far back as 1971 by

Palmer et al. [46] at the National Institute for Metallurgy (MINTEK). Tellurium is

added to the final solution to form a tellurite which is reduced to the metal by the

28  

addition of stannous chloride. The tellurium precipitate containing the precious

metals is then dissolved by aqua regia. The platinum industry uses Tellurium co-

precipitation in conjunction with the NiS-FA technique [47-50]. Other sample

preparation techniques involving Te co-precipitation are also published and are

referred to in section 3.6.

3.6 APPLICATIONS OTHER THAN Pb-FA USED IN THE

DETERMINATION OF TRACE AMOUNTS OF PRECIOUS METALS

Applications other than fire assay techniques include acid digestion, ion exchange

and isotope dilution. The measurement of trace amounts of PGEs can be obtained

in the final solution using various spectroscopic techniques that have been well

documented and include inductively coupled plasma mass spectroscopy (ICP-

MS) and graphite atomic absorption spectroscopy (GFAAS).

3.6.1 Applications using Inductively Coupled Plasma Mass Spectroscopy

(ICP-MS)

ICP-MS has been widely employed in the determination of precious metals due to

its low detection capability. Several procedures involving its use in conjunction

with various sample preparation techniques have been published..

Meisel et al. [51] developed an analytical procedure for the determination of Ru,

Pd, Re, Os, Ir and Pt in chromitites and other geological materials by isotope

dilution inductively coupled plasma mass spectroscopy (ID-ICP-MS) after

digestion in a high pressure asher (HPA-S). The digestion technique is similar to

that using Carius tubes but may require higher temperature. The isotope selected

was that of 190Os. Osmium was determined as OsO4 by ICP-MS directly after

digestion through a sparging technique. The rest of the elements were pre-

concentrated by means of anion column chromatography, the resin was digested

directly without elution and determined by ICP-MS. The results obtained for

reference material chromitite CHR-Bkg for the elements Ru, Os and Ir compared

well to the certified values. Although the mean values for Pt and Pd were slightly

lower than the reference values, they were still within the confidence intervals

(CI) of 22 % for Pt and 16 % for Pd. The method demonstrated its ability to

29  

determine both Os for isotopic measurements and Ru, Pd, Rh, Ir and Pt for

geochemical interpretation in chromitites. One of the limitations of this method is

that HPA-S can only digest small sample portions.

Totland et. al [52] studied microwave digestion and alkali fusion procedures for

the determination of precious metals in geological materials by ICP-MS. One

method developed involved the digestion of sulphide rich samples with an acid

mixture of HNO3-HCl-HF-HClO4 in closed low pressure microwave vessels,

followed by evaporation to incipient dryness and dissolution in 1 M HCl. A

second method involved the digestion of silicates, sulphides and chromite

samples with aqua regia – HF in closed low pressure microwave vessels, followed

by evaporation to incipient dryness and dissolution in 0.5 M HCl. The insoluble

residue was separated and fused with 1:1 sodium peroxide (Na2O2) and dissolved

in 0.5 M HCl. The detection limit was found to be between 0.2 and 1 ug g-1 for

both methods restricting their application to mineralised samples. Good

reproducible results were obtained for the reference materials studied. However,

the resulting solution required dilution to reduce the total dissolved solids (TDS)

which meant that the PGEs were below the detection limits of an ICP-MS for an

in-house chromite sample. Another limitation was the small sample size which

could be accommodated by microwave systems.

Microwave digestion is also well documented in the automotive catalytic industry

for the determination of Pt, Pd and Rh by ICP-MS. Brown et al. [53] developed

two methods involving microwave digestion for the digestion of a 5g autocat

sample which produced constant results with good precision, however there

appeared to be problems with the elements Ce, Ni, Fe and Ba.

Simpson et al. [54] developed a microwave digestion method which provided a

quick total dissolution of precious metals within 0.2g of sample. For greatest

accuracy in the analysis of Pt and Pd, a matching isotope dilution calibration

procedure was used. For Rh a matching bracketing calibration procedure was

used with Ru as an internal standard. The results showed excellent agreement

with the reference values for Pt, Pd and Rh with expanded uncertainties of 0.9%,

1.1% and 0.9%, respectively (NIST 2556).

30  

3.6.2 Applications using graphite atomic absorption spectroscopy (GFAAS)

GFAAS was widely employed for the determination of trace elements, before the

arrival of ICP-MS. Although several publications have been found on the

determination of ultra trace amounts of precious metals by GFAAS, it is not a

favoured technique for trace analysis in the platinum industry. Sample

throughputs are slow and fall short of production requirements which require

shorter turnaround times. Several procedures involving its use, in conjunction

with various sample preparation techniques, have been published.

Gupta [55] developed a procedure for the determination of trace amounts of

precious metals in geological material by GFAAS after separation by ion-

exchange and co-precipitation with tellurium. Ore, Concentrate, mattes, silicate

and iron-forming rock reference materials were dissolved in hydrochloric (HCl)

acid and aqua regia in a Teflon beaker. These samples were further boiled for

several hours, brought to dryness, wetted by addition of HCl acid, brought to

dryness and repeated until all the metal salts were converted into chlorides. The

residue was filtered and washed with 1M HCl and filtrate retained. The residue

was fused with Na2O2 and the melt dissolved in HCl and transferred to the

retained filtrate. For the separation of the precious metals a portion of the sample

was prepared and passed through an ion-exchange column and the eluate

collected. The eluate was prepared for analysis by GFAAS. Another portion of

the retained filtrate was treated with 1mg per ml Te solution and brought to

boiling point. 40% stannous chloride was added until a copious black precipitant

was formed. The solution was filtered and the black precipitant containing the

precious metal tellurites was dissolved in aqua regia. The final solution was

treated repetitively with numerous acid portions that were brought to dryness until

all the salt had been converted to the chloride. The final solution was analysed by

GFAAS. Good recoveries were obtained for reference materials of ores,

concentrate and mattes containing trace levels of ug g-1. The Canadian iron-

forming reference materials after co-precipitation with Te also produced good

agreement containing ng g -1.The silicate rock however showed high results for Pt

and Ir because of incomplete separation of elements such as titanium by the ion-

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exchange columns used. Although this procedure produced successful results it is

a very tedious procedure and would be prone to handling error in a routine set up.

3.7 CHARACTERIZATION OF MINERALS

3.7.1 Introduction

Although mineralogical technologies have improved dramatically over the last

few decades, the characterization of PGEs in ore remains problematic. Some

mineralogists have even compared it to finding a needle in a haystack.

Penberthy, Oosthysen and Merkle [24] mentioned the following contributing

factors towards inconsistent data obtained from mineralogy studies of the UG2

Reef:

• Low concentration levels of PGEs < 1 g t -1.

• Inconsistent recovery of PGEs from the UG2 Reef.

• The occurrence of PGEs both as discrete platinum group mineral phases

and less commonly, sub-microscopically in other phases, mostly sulphide

minerals and to a lesser extent, chromite and silicate.

• The small grain size of the discrete platinum group minerals, < 10 μm.

Cuyper et al. [56] emphasized that the occurrence of sub-microscopic mineral sizes

below the detection limit of an electron probe micro analyser would result in

inconclusive data analysis. However, the most important factor is that there is no

cost-effective universal standard pre-concentration method for samples prior to

mineralogy studies. This presents an industrial problem, as mineralogy studies

are invaluable in customising Mineral processing operations with a view to

improving final PGE recoveries and final Concentrate grades.

3.7.2 Mineralogical instrumentation and applications

Many techniques for the studying of mineralogical information have been

developed and include: [41]

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• Scanning electron microscope equipped with energy dispersive x-ray

analyser (SEM/EDX) or wavelength dispersive x-ray analyser

(SEM/WDX).

• Variable pressure scanning electron microscope (VP-SEM).

• Electron probe micro-analyzer (EPMA).

• X-ray diffraction (XRD).

• Quantitative evaluating material by scanning electron microscope (QEM-

SEM).

• Mineral Liberation Analyzer (MLA).

The mining industry worldwide has benefitted greatly from these technologies, as

mineralogy techniques have improved dramatically over the last few decades. The

platinum industry in South Africa for instance is well known for using QEM-

SEM and MLA extensively to obtain mineralogical information which assists in

identifying those factors which affect plant performance in particular recoveries

and final Concentrate grades. [41]

The occurrence of PGE fractions as solid solution in base metal sulphides is

common in platinum ores. An investigation done by Ballhaus et al. [51], on

Merensky Reef material suggested that pentlandite was the only base metal phase

which accommodated appreciable amounts of PGE, Pd from 114 to 359 ppm, Rh

< 3 to 99 ppm and Ru from < 3 to 45 ppm. The material was analysed by a proton

microprobe using PIXE. The mean detection limits (MDL) for trace analysis at a

99% confidence level (CL) are in ppm: 15 for Pt, 4 for Pd, 3 for Ru and Rh. No

significant quantities of PGEs were found in the base metals of pyrrhotite and

chalcopyrite as they were close to detection limits.

Iwasaki et al. [58] reported PGEs which were found in solid solution in arsenides

and sulfarsenides. The PGE concentration in solid solution in cobaltite (CoAsS)

and gersdorfite (NiAsS) ranged from 56 to 1840 ppm. Platinum and palladium

concentrations in chalcopyrite, pentlandite and pyrrhotite were below the

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detection limits of the proton microprobe analyser, which detection limits were

reported to be 50-60 ppm for Pt, and 1.3-3 ppm for Pd.

Laser ablation microprobe inductively coupled plasma mass spectroscopy (LAM-

ICPMS) is also a much favoured technique for the detection of PGEs within

mineral grains and has a much lower detection limit than the proton microprobe. [59] A study using LAM-ICPMS showed that the sulphide mineral pentlandite may

accommodate levels of Ru, Rh and Pd of up to a few percent. The simultaneous

analysis of Au and Pt indicated discrepancies between the bulk and in situ

analysis suggesting that Au and Pt were either present as minor sulphide phases

which were too small to be analysed by LAM-ICPMS or that they occured as

discrete minerals at the sulphide-silicate boundaries of the grain. Laser ablation

studies were able to confirm that PGEs are concentrated in the mantle sulphides

and occur in the lattice of the sulphide phase with the exception of Pt.