METALS

BACKGROUND TO THE INVENTION

This invention relates to an extraction method for the recovery of Precious Metals including Platinum Group Elements (PGEs), and other Precious Metals (gold and silver); Base Metals (including nickel, cobalt and copper) from ores and concentrates, particularly "Platreef ores and concentrates.

Platreef is a unique stratabound PGE reef type (deposit type) belonging to the suite of PGE deposits found in the Bushveld Igneous Complex (BIC) in Southern Africa, where the other reef types of economical importance are the stratiform Merensky and stratiform Upper Group 2 (UG2) Chromitite reefs, each with their unique petrology and mineralogy. All three reef types belong to the Rustenburg Layered Suite (RLS). Platreef is located in the region of Potgietersrus / Mokopane in the Republic of South Africa and the outcrop has a general North-South strike in the proximity of the 29°00' N-S longitude (from 29°00' up to 25 km West of this longitude) and dips down towards a west-south-westerly direction. It is roughly situated between the E-W latitudes of 23°30' and 24°30'.

Traditional concentrate (using milling and flotation) - smelt - refine processing routes for Platreef are significantly more expensive on a "per Platinum ounce" basis than for Merensky and UG2, due to the mineralogical and process metallurgical particularities associated with this reef type. The mineralisation and "prill split" (the contribution of each PGE to the total PGEs) of PGEs & Au and its association with base metals sulphides and silicate minerals are significantly different from the traditional PGE resources, namely the Merensky & UG2 Chromitite Reef Types, leading to the conventional metallurgical extraction processes (concentrate- smelt-refine) often being uneconomical for Platreef only. In some cases, companies blend this reef with high grade Chromitite reefs to alleviate this problem, where this blending is a feasible process option. However, investing capital into erecting Greenfields plants for the conventional processing of Platreef by itself is marginal. Furthermore, Platreef is situated in an arid and remote area of South Africa, the region being poorly supported by available infrastructure.

Platreef is characterised by a high Pd: Pt ratio which is mineralised mostly as tellurides, arsenides and bismuth-tellurides, whilst Pt/Pd sulphides and alloys are much less predominant than in Merensky or UG2 ores. Platreef contains a significant gold concentration, but contains very low concentrations of rhodium, ruthenium and iridium. Platreef contains significant amounts of base metals (copper, nickel and cobalt) compared to other PGE-reefs and they are mineralised mostly as sulphides. The gangue mineral matrix is rich in magnesium and iron silicates and iron sulphides (mostly pyrrhotite with some pyrite).

The significant mineralogical variation of Platreef in comparison to other South African BIC-based ores is well documented in public archival journals. The uniqueness of Platreef is not only in terms of its mineralisation, geology and stratigraphy, but also in terms of the associations of PGEs with the underlying base metals sulphides, gangue silicate phases and its distribution along grain boundaries of silicates and base metal sulphides (BMS).

Currently, the existing commercial processing and recovery of base metals and PGEs from ores entail the crushing, milling, flotation of the ores, smelting of concentrates and converting the furnace matte or alloy to a final product mixture containing a mixture of sulphide and alloy phases. The product leaving the smelter is either leached to recover the nickel, cobalt and copper, or the smelter product is slow-cooled, crushed and the sulphide and alloy portions are separated by magnetic separation. The base metals refinery (BMR) / base metals removal plant is also used to reject iron, sulphur and waste components such as Se, Te, Bi, As, Pb and other contaminants which are deleterious to subsequent PGE separation. After removal of the bulk of the base metals and other contaminants, a PGE residue with a grade of greater than 50% is typically produced. This PGE concentrate derived from the BMR is typically leached in a chlorine and hydrochloric acid environment to solubilise the PGEs. The PGE- containing solution is then separated into its constituent metals through a number of processes that might include any of a range of repeated precipitation and dissolution processes, solvent extraction, ion exchange or Molecular Recognition Technology (MRT™). The conventional process route to recover and beneficiate PGMs has limitations. It is considered suitable for sulphide containing ores from which relatively low quantities of concentrate are produced with high recoveries of PGEs that can be treated economically through a smelter (implying a low mass pull at the concentrator, in other words a small percentage of the ore mass reporting to flotation concentrate). Ores that have been oxidised or weathered such as reef outcrop material or shallow ores, and ores where much of the PGEs are associated with large amounts of Pyrrhotite/Pyrite, or as ultrafine inclusions in silicaceous waste rock, or ores with much of the PGEs occurring as tellurides, bismuthides, or arsenides or combinations of these, tend to show poor (slow) flotation kinetics, and large amounts of concentrate of low grade are often required to be produced to achieve an acceptable recovery (i.e. large mass pulls combined with ultra fine milling in energy intensive stirred mills). This implies a large investment in smelting infrastructure for a given amount of PGE production and can lead to uneconomic smelting and refining operations. Ultrafine milling and the smelting of low grade concentrates are very energy intensive per mass unit of the contained PGEs. The erection of new smelters to process PGE containing flotation concentrates has been significantly curtailed due to a combination of environmental legislation, very high capital costs, very limited availability of electrical power (on the Southern African subcontinent) since the year 2008 and high associated greenhouse gas emissions (indirectly through electricity provided through coal fire power stations). Furthermore, to supply the skilled labour requirement associated with smelting, the smelter is required to be close to major towns or cities, yet the same towns and cities are often negatively inclined towards having smelters in their close proximity. This all means that any new mining and concentrating operation will have to ship flotation concentrates over excessively large distances to existing smelters, at high and rapidly rising costs, leading to increased metal pipelines. The smelters themselves, due to the number of safety incidents (run-outs and explosions) and the cost of smelting and handling off gas emissions are imposing ever tighter concentrate specifications on concentrators.

Recently, pressure leaching of flotation concentrates has been proposed to circumvent the smelting requirements, however they do require energy intensive ultrafine grinding, significant water consumption and, due to mother liquor associated losses, residues are produced that are extremely fine, difficult to filter and environmentally unstable, so that environmental risks remain high and a number of risk factors makes the long term sustainability questionable.

It is an object of this Invention to provide a low (operating and capital) cost alternative extraction method that can be used to leach Precious Metals, particularly PGEs and gold as well as the Base Metals, particularly nickel, copper and cobalt from heaps such as crushed run-of-mine Platreef ore or Platreef-based flotation concentrates.

SUMMARY OF THE INVENTION

According to the invention there is provided a process for recovering Base Metals and Precious Metals from ore or concentrate, the process including the steps of: i) constructing a heap comprising ore and/or concentrate containing Base Metals and Precious Metals; ii) leaching the heap to remove Base Metals and, preferably, also sulphur, in a leachate solution, in a bioleaching process, preferably by inoculating the heap with microbial inocula of mesophilic and/or thermophilic bacteria and/or archaea and bioleaching the ore and/or concentrate, in an exothermic process and, after rinsing the heap with water; iii) leaching the heap with aqueous cyanide solution at ambient pressure and elevated temperature above 40°C and below 90°C to remove Precious Metals from the heap in a leachate solution.

The Base Metals are typically Ni, Co, Cu and Fe.

The Precious Metals are typically PGMs consisting of Platinum, Palladium, Rhodium, Ruthenium, Iridium, Osmium; as well as Gold and Silver. For the purposes of this invention, the precious metals are PGM-rich and poor in gold and silver, that is, the distribution of precious metals is such that the PGMs make up more than 75% of the total precious metals and the gold and the silver the remaining 25% of the precious metals, or less. The invention therefore relates to ore and concentrates where such ores or concentrates are predominantly classified as PGM ore and concentrates of which gold and silver are by-products and are secondary to PGM production, but occur by association. The ore or concentrate may be from an ore or concentrate that is characterised by a high Pd: Pt ratio (typically larger than 0.8:1 ratio) which is predominantly mineralised as tellurides, arsenides and bismuth- tellurides, such as Platreef ore, but may also occur in smaller proportions as sulphides, selenides, antimonides or alloys mineralised as tellurides, arsenides and bismuth-tellurides.

Typically, the ore has been crushed to a particle size with 100% passing 25 mm.

Concentrate (typically 80 % passing 150 micrometre) may be coated onto rock or gravel (typically 80% passing 25 mm), in which case microbial inocula may be included with the coating.

Preferably, the bioleaching step is carried out under acidic conditions with pH<3, preferably pH<2, and preferably under forced aeration conditions at temperatures between 50 °C and 90 °C, preferably between 50 °C and 70 °C using a mixed biomass culture consisting mostly (but not only) of moderate and extreme thermophiles such as:

» Acidithiobacillus ferrooxidans

• Ferroplasma cupricumulans

• Sulfolobus metallicus

• Metallosphaera sedula

• Sulfobacillus thermosulfidooxidans

« Acidithiobacillus thiooxidans

« Leptospirillum ferriphilum

• Leptospirillum ferrooxidans, or

• or any combination of the above or other microbial species (a genetic changes may occur over time as the biomass adapts to its environment). During the heat-up phase of the heap, mesophiles will be active in the intermediate temperature ranges (20°C to 50°C) and contribute to establishing the overall microbiological activity in the heap. The population of the biomass culture will change with time as the heap heats up through exothermic reactions and will stabilise around between 50°C and 70°C depending on the availability of base metal sulphide minerals, sufficient aeration and irrigation and biomass population density.

The bioleaching step may take 1 to 2 years.

After the removal of Base Metals from the heap, the heap is preferably rinsed and treated with a base such as lime to increase the pH to greater than 8, preferably greater than 9 and up to 12.

In the cyanide leach step, cyanide-containing solution is heated, using solar heating, and augmenting the heating (if required) by inter-heap heat exchange, of the process liquors to a temperature greater than 40°C to 90°C, preferably 50°C to 70°C which is preferably applied to the heap by drip irrigation. Heat and moisture can be more efficiently retained in the heap through the use of transparent, flexible plastic sheeting to cover the outer surface of the heap, but its use is not mandatory.

The cyanide solution may have a concentration of 0.15 gram per litre (g/l) to 15 g/l, preferably 0.5 g/l to 2.0 g/l.

The cyanide leaching step may take 1 to 2 years, depending on the recovery target for platinum, and the use of either concentrate or crushed ore.

Heating of the cyanide-containing solution is preferably done using a renewable energy source such as energy which comes from natural resources such as sunlight, wind, rain, tides, and geothermal heat, most preferably solar energy using solar panels. If required, the heating can be supplemented by heat exchange with the process liquors from a bio- leaching heap.

Preferred aspects of the process of the present invention include:

• utilising a inter-heap heat exchange from liquors derived from the microbial heap leach to heat the return liquor to the cyanide heap leach, once both heap operations are running.

o utilising solar heat via heating panels to heat process liquors, in particular the alkaline cyanide solutions.

® recovering the base metals through precipitation as a mixed sulphide precipitate (for example through NaSH, H ₂ S or Na ₂ S) addition) or solvent extraction and electrowinning to produce refined metals.

o recovering and upgrading the PGMs via adsorption and elution from activated carbon, Molecular Recognition Technology (MRT™), ion exchange, solvent extraction, zinc or iron cementation or specific combinations of these technologies.

o if required, the heaps after the dual stage leaching can be used for back-fill into mines as essentially all of the minerals leading to potential acid-mine- drainage has been converted and is in oxidised form.

DESCRIPTION OF THE DRAWINGS

Figure 1 is a high level process flow of an embodiment of the invention with inter-heap and solar heating of cyanide heap leach liquors;

Figure 2 is a graph showing the extraction of base metals in a bio leaching experiment; Figure 3 is a graph showing the extraction of platinum in a cyanide leaching experiment;

Figure 4 is a graph showing the Fe extraction curves in a bio leaching experiment;

Figure 5 is a graph showing the Cu extraction curves in a bio leaching experiment;

Figure 6 is a graph showing the Ni extraction curves in a bio leaching experiment; and

Figure 7 is a graph showing the extraction of platinum in a cyanide leaching experiment.

DESCRIPTION OF PREFERRED EMBODIMENTS

This invention relates to an extraction method for the recovery of Precious Metals including Platinum Group Elements (PGEs), and other precious metals (gold and silver); and Base Metals (including nickel, cobalt and copper) to aqueous solution from ores and concentrates, particularly Platreef ores and concentrates. The PGEs consist of platinum, palladium, rhodium, ruthenium, iridium, and osmium. For the purposes of this invention, the precious metals are PGM-rich and poor in gold and silver, that is, the distribution is such that the PGMs make up more than 75% of the total and the gold and the silver the remaining 25% or less. The invention therefore relates to ore and concentrates where such ores or concentrates are predominantly classified as PGM ore and concentrates of which gold and silver are by-products and are secondary to PGM production, but occur by association.

The extraction of Base Metals and Precious Metals extraction is achieved through a sequential heap leach operation where high temperature (greater than 50°C) acidic microbial heap leaching is applied to extract most (greater than 70%) of the contained nickel, copper and cobalt, as well as some of the iron and most of the sulphur, to expose the Precious Metals and their minerals to a subsequent stage of cyanide based heap leaching where the cyanide based extraction also occurs at elevated temperatures (greater than 40°C). It is noted that while the bulk extraction of base metals occur above heap temperatures of 45°C, bio-heap leaching is inherently a batch operation and the heap takes a significant period to heat up from ambient temperature (15-30°C, depending on the initial conditions and climatic conditions) to the preferred leach temperatures of greater than 45°C. During this initial heat-up period some incipient leaching will occur. Most (greater than 95%) of the sulphide sulphur is either leached as base metal sulphates, or is converted into a ferri-hydroxy-sulphates residue (such as from the jarosite group of minerals) which does not lend itself to further acid drainage. Jarosite formation can be reduced by manipulating the feed composition, specifically the amount or iron and the pH (by adjusting the amount of acid in the feed).

While the microbial heap bioleaching of base metal sulphides is exothermic and therefore a net producer of heat, the cyanide heap leach is roughly energy-neutral. However, the cyanide heap leach requires heating (to above 50°C) for good PGM extraction. The energy may come from any source, but the source is preferably a renewable energy source such as energy which comes from natural resources such as sunlight, wind, rain, tides, and geothermal heat. The preferred energy source is solar energy which, if required, can be supplemented by further heat exchange between the heaps (heat from the heap bioleach being used to heat up the liquor that feeds the cyanide heap. Due to heap leaching being a low intensity, slow process (compared to smelting and pressure leaching), it lends itself to low intensity heating mechanisms such as solar heating, especially as temperatures are below the boiling point of water. The microbial heap leach is operated under forced aeration and acidic conditions (pH<3), whilst the cyanide heap leach can be performed under basic (pH>9) conditions using either natural or forced aeration. The heaps can be operated using either crushed Platreef ore (<25 mm) or Platreef flotation concentrate (< 50 micrometre) coated onto waste rock (<25 mm).

In an embodiment of the invention, crushed Platreef ore or Platreef concentrate is treated in a process of extracting Base Metals and Precious Metals to aqueous solution, which includes the following steps: a) a heap is stacked using an automatic stacker or a dump-truck to stack size-graded material of a particle size smaller than 25 mm onto the heap. In the case of ore leaching the ore itself is stacked or dumped, whereas for concentrate leaching, size-graded waste rock, gravel or low grade ore, is used onto which concentrate slurry is coated, for example using the Geocoat® process. Microbial inoculums may be included with the slurry coating, or sprayed onto the heap during a later stage.

In both cases the ore or concentrate-on-gravel heaps are loaded onto impervious membranes, such as welded high density polyethylene (HDPE) sheeting which, in turn, is supported on a geo- membrane made of graded gravel rocks and clays to ensure an impervious base and which is geotechnically designed to carry the load of the heap above. Pregnant leach solution derived from the heaps is accumulated in ponds from where it is pumped for base metal reclamation or PGM recovery and upgrading. b) The heap is inoculated using a mixed culture of extreme thermophiles, thermophiles and mesophiles.

The microbial culture used consists typically of, but is not limited to:

• Acidithiobacillus ferrooxidans

« Ferroplasma cupricumulans

• Sulfolobus metallicus

• Metallosphaera sedula

• Sulfobacillus thermosulfidooxidans • Acidithiobacillus thiooxidans

* Leptospirillum ferriphilum

o Leptospirillum ferrooxidans. c) After inoculation, microbial bio-heap leaching proceeds in an exothermic reaction at elevated temperatures of 40°C to 90°C, taking into account incipient leaching during the heap heat-up stage from ambient conditions to 45°C, at atmospheric pressure associated with altitudes from mean sea level up to 2500 meters above mean sea level and under acidic conditions with pH<3, but preferably the pH<2.0, and forced aeration conditions to extract most of the nickel, cobalt, copper, and sulphur and some of the iron. The leach cycle may extend up to 2 years. The acidic and oxidising nature of the leach also solubilises arsenides, tellurides, selenides and bismuthides, thereby making Precious Metals more amenable to subsequent cyanide leaching. These elements (As, Te, Se, Bi) may reprecipitate in the heap, but separate from the Precious Metal minerals of which they were originally part. Most of the sulphidic iron and sulphur in the heap re-precipitates as ferri-hydroxi-sulphates (e.g. jarosites), the rest being converted to ferrous and ferric sulphate solutions, sulphuric acid and base metal sulphates. d) After sufficient removal of the Base Metals and sulphur, the heap is rinsed with water and treated with lime to increase the pH to greater 9, preferably 10.5. e) The heap is irrigated with hot cyanide-containing solution at a temperature greater than 40°C, but lower than 90°C, most preferably 50°C to 60°C and cyanide leaching at elevated temperature and atmospheric pressure and proceeds to extract the Precious Metals including PGEs, Au and Ag. For cyanide heap leaching, either forced (using a fan or blower) or natural draft aeration may be used to provide the required oxygen for leaching. The cyanide solution has a cyanide concentration of 0.15 gram per litre (g/l) to 15 g/l, preferably 0.5 g/l to 2.0 g/l. The leach cycle of the heap can be up to 2 years. The addition of hydrogen peroxide (0.1 M) or a similar oxidant may aid the dissolution of PGMs. f) Heating of the cyanide containing leach solution to temperatures above ambient may be supplied using solar panels and/or via heat exchange using warm water that is heated by the process liquors from a bio-leaching heap. g) The barren heap can be reclaimed for mine backfill of stacked as waste rock. h) The preferred use of the heap leach pad is as a reusable leach pad where ore (or waste rock and concentrate) is loaded/stacked upon, leached and reclaimed for removal elsewhere.

For good Precious Metal recovery in the cyanide heap leach, it is important for there to be high Base Metals extraction in the bio-heap leach, otherwise the Base Metals consume the cyanide. Preferably, more than 70% (by mass) of the Base Metals are removed. This is achieved by sufficiently long leach periods, coupled with the maintenance of the correct pH (pH<2.0).

It is also important that elemental sulphur is removed from the bio-leach heap, otherwise the sulphur will consume cyanide and form undesirable thiocyanates in the cyanide heap leach. The removal of sulphur is achieved by the inoculation of the heap during bioleaching with sulphur oxidising bacteria (part of the microbial population described above).

The following operating ranges as depicted in Table 1 apply for the bio- heap leach and cyanide heap leach: Table 1 : Operating ranges for the two-stage elevated temperature heap leach operation

Operating Bio-heap leach Cyanide Heap Leach

ParaMinimum Maximum Typical Minimum Maximum Typical meter

Aeration 0.0 m ³ 0.6 m ³ 0.4 m ³ 0.0 m ³ 0.6 m ³ 0.4 m ³ Rates per per per per per per

(hr.m ² ) (hr.m ² ) (hr.m ² ) (hr.m ² ) (hr.m ² ) (hr.m ² )

Solution 2 litre 12 litre 8 litre 5 litre 60 litre 15 litre

Irrigation per per per per per per

Rates (hr.m ² ) (hr.m ² ) (hr.m ² ) (hr.m ² ) (hr.m ² ) (hr.m ² ) pH 0.5 2.2 1.5 9 12 10

Sulphuric 5 gram 20 gram 10

Acid per litre per litre gram

Concenper

tration litre

Cyanide 0.15 15.00 2.00 Concengram per gram per gram tration litre litre per litre

Microbial 10 ⁵ per 10 ¹⁰ per 10 ⁸ per

Concenml ml ml

tration solution solution solution

Pregnant Cu: 0 - 10 g/l

Solution Ni: 0 - 20 g/l

Base Co: 0 - 1 g/l

Metals Fe: 0.5 - 5.0 g/l

concentrations

after

biolech

Precious — Pt: 0 - 20 ppm Metals Pd: 0 - 50 ppm

concenAu: 0 - 10 ppm

trations Ag: 0 - 5 ppm

after Rh: 0 - 1 ppm

Cyanide Ru: 0 - 1 ppm

leach Ir: 0 - 1 ppm

Standard methods exist for the recovery and separation of Base Metals from acidic sulphate solution, such as solvent extraction and electrowinning of base metals or production as crystalline sulphates. Similarly, standard methods exist for the recovery of PGEs and other precious metals from cyanide solutions onto activated carbon, powdered zinc, ion exchange resins, Molecular Recognition Technology (MRT™) resins, and into solvent extraction liquors. All of these techniques concentrate the PGEs from the dilute heap leach solutions into/onto a medium for elution from adsorbed media. The eluted PGEs are cemented or precipitated onto Zn, which is subsequently melted to fume off the Zn and to produce a high grade metallic PGE concentrate which can be processed at a conventional PGE precious metals refinery.

A preferred embodiment of the invention is illustrated in Figure V. a microbial heap bioleaching stack is indicated by the numeral 10, a cyanide heap leaching stack is indicated by the numeral 12, a solar heating system is indicated by the numeral 14, a PGM upgrade and recovery process is indicated by the numeral 16, and a BMS separation and recovery process is indicated by the numeral 18. In this embodiment of the invention, once the heap has reached a steady temperature of 60-70°C, leachate stream 20 from the exothermic bio-heat leaching stack 10 which is at a temperature of about 65°C is passed through a heat exchanger 22 and then onto the BMS separation and recovery process 18. A BMS-barren solution 24 from the BMS process 18 and which is at a temperature of approximately 50°C is recycled to the bio-heat leaching stack 10. A leachate 26 from the cyanide leaching stack 12 is passed through the PGM upgrade and recovery process 16 and a PGM-barren solution 28 which is at a temperature of 40°C is passed through the heater exchanger 22 where it is heated to a temperature of approximately 45°C and passed on by a stream 30 to the solar panel heating system 14 which heats the solution to a temperature of approximately 55°C and which is supplied by a stream 32 to the cyanide leaching stack 12. This process provides an energy efficient system for heating the cyanide solution supplied to the cyanide heat leaching stack 12. In a specific option, PGM-barren solution 28 may pass directly for solar heating, without pre-heating in heat exchanger 22. Sufficient solar panels should be used to bring the cyanide solution temperature to about 55 °C. pH adjustment and the removal and control of the ferric-ferrous ion ratio in the BMS-barren solution is performed in the BMS recovery and separation section 8. pH adjustment through lime addition (or alternative alkali) and cyanide addition (e.g. sodium or potassium cyanide) to the appropriate concentration (refer to Table 1 ) is performed in the PGM-recovery process 16.

This invention relates to the hot, sequential stage heap leaching of base and precious metals from a Platreef ore or concentrate into an aqueous solution, and not the subsequent processing, which is well covered in published literature. Transparent plastic sheeting may be used to better retain moisture within the heap, by covering the heap surface.

The leaching from either crushed ore or from flotation concentrate reflects the alternative embodiments of this invention.

The process of the present invention not only provides a low (operating and capital) cost alternative extraction method that can be used to leach Precious Metals, particularly PGEs and gold as well as the Base Metals, particularly nickel, copper and cobalt from heaps of either crushed run-of- mine Platreef ore or Platreef-based flotation concentrates, it also uses much less water (about 8%) than conventional processes and it also uses low energy because energy is supplied from the sun and/or from the exothermic bio-leaching part of the process.

The invention will now be described in more detail with reference to the following non-limiting Examples.

Example 1 - Concentrate Preparation and Characterisation

For the test work, a low-grade flotation concentrate was acquired from a Platreef mine operation in the Mokopane area, South Africa. The concentrate was received in 10 drums each containing approximately 18 kg of the concentrate. The material was damp and was air dried for a few days and thoroughly mixed to homogenise it. A 2-way riffle splitter was then used to split the ore into 20 kg samples and then 10 kg samples. The 10 kg samples were then split, using a 10-way Dickie and Stockier (Pty) rotary splitter, into 1 kg samples. The samples were then packed into 4 kg samples in polyethylene bags. Samples for the leaching tests, size analysis, and PGM and BM solid assays were obtained from the 4 kg samples by using a Fritsch Rotary 10-way sample divider and a Quantachrome Instruments' 8-way Rotary Micro Riffler. The ore concentrate was sized using wet screening showing the following size distribution:

Table 2: Size analysis results

Screen Sizes ( m) % Passing

75 97%

45 97%

38 83%

A fire assay and mineral liberation analysis (MLA) revealed the following grade assay and mineral composition respectively: Table 3: Precious metals grade assay

Total (6E) Pt Pd Au Rh Ru Ir

g/t g/t g/t g/t g/t g/t g/t

=56 21 27 3.7 1.8 1 .5 0.5

Table 4: Major base metals and gangue elements

Cu Ni Fe Co Mg Al Ca Si Cr Total S

% % % % % % % % % %

2.3 3.4 16.4 0.11 10.4 1.43 4.01 17.3 0.15 8

Table 5: Major PGM minerals grouped by relative abundance

Mineral Group % Area

Alloys 20.8

Sulphides 28.6

Arsenides 23.8

Sulpharsenides 7.8

Tellurides 19.1

Among the above PGM minerals, sperrylite (PtAs ₂ ) and cooperite (PtS)

were present in by the far the largest portions of 19.6 % and 16.1 %

respectively. The MLA also revealed that the PGMs were 61 % liberated.

Table 6: Major base metal and gangue minerals

Minerals/Mineral Group Wt % Abundance

Chalcopyrite 4.93

Pentlandite 7.70

Pyrite 1.93 Pyrrhotite 5.27

Silicates 66.5

Others 13.67

Example 2 - Heap bioleaching on concentrate: Experiment 1

Aim

The aim of this experiment was to determine the effectiveness of using a thermophilic bioleach proces to extract copper and nickel from a platreef ore prior to a cyanide leach for precious metals. A low-grade flotation concentrate was used in this first test as a proxy, to obtain results in a shorter period of time than would be obtained with whole ore leaching. This would allow for repeat experiments to determine the best operating conditions before commencing test work on the whole ore which would proceed for a significantly longer period of time.

Methods

Four samples of flotation concentrate weighing 650 g each were made into slurry using deionised water in a solid to liquid ratio of 5:3. The slurry was coated onto granite pebbles, packed into plastic columns and left overnight to air dry. A solution containing 30 g/l H ₂ S0 ₄ as pumped into the columns from the top at a rate of 1 l/day (translated from an industrial flow rate of 5 l/m ² /h), for 5 days in order to dissolve as much of the acid soluble BM minerals as possible, and the effluent was collected from the bottom. It was thereafter replaced with the main leaching solution containing 2 g/l Fe (1 g Fe ³⁺ as ferric sulphate and 1 g Fe ²⁺ as ferrous sulphate) and 10 g/l H ₂ S0 ₄ , which was pumped into the column at the same rate and in the same manner. A mixed culture of thermophilic microorganisms in which Metallosphaera sedula was the dominant species (99 %), was inoculated into the 4 columns which were operated at temperatures of 65, 70, 75 and 80°C. Real time quantitative qPCR was used to identify and quantify Metallosphaera sedula in the culture. The columns were aerated at a rate of 130 ml/min. Samples of solution were collected from the effluent at various intervals for AAS analysis of Cu, Ni and Fe. The pH and redox (vs Ag/AgCI) were also measured and recorded each time a sample was taken using a standard pH meter and Redox meter respectively. Additionally samples were also collected for microscopic inspection, to determine the well being of the microorganisms. The temperatures were also monitored and recorded when samples were taken. After 88 days of leaching, the extractions achieved were calculated by the amounts present in solution and by fire assay of the residual concentrate material.

On completion of the experiment the columns were emptied, and the concentrate washed off the granite with water and recovered by sieving and pressure filtration. The concentrate was further washed with caustic water to remove residual acid and dried. From the dry samples, sub-samples were obtained for fire assays, XRD and MLA analyses, while the bulk of the samples were used in subsequent cyanide leaching experiments to extract the precious metals.

Results and Discussion

Amongst the four columns (Table 7), the 65°C column was the overall best performer, contrary to the expectation that high temperatures would produce better results. The lowest co-extraction of the Pt and Pd were also experienced in this column. However across all columns Rh and Ru was co-extracted in excess of 50 %. This is comparable to leaching currently experienced in a typical base metal refinery amongst platinum producers in South Africa (Dorfling et. al. 2010). Numerous ion-exchange technologies exist to recover these metals from solution. It was observed that after 43-46 days, iron precipitation occurred in all the columns following the trend in Figure 2. It is believed that at this point the leaching of the iron sulphides was mostly complete and the source of the iron in the precipitate was mostly from the feed solution. An MLA analysis on a sample of concentrate residue confirmed that the BM sulphides had been all but leached and there was hardly any elemental sulphur present. XRD and MLA analyses also confirmed that a considerable amount of the concentrate residue (35 %) was an iron jarosite. The MLA analysis also indicated that the bioleaching had increased the PGM mineral liberation from 61 % to 81 %, but there was no change in the PGM mineralogy.

As expected in an acid leach, gangue element dissolution was high in all columns, but of particular interest are Mg and Al. At levels of 10-12 000 mg/l these cations can inhibit ferrous oxidation (Ojumu 2008). The average concentrations of Mg and Al in this experiment were 243.6 mg/l and 38.9 mg/l respectively. However this was a once through operation with no recycle of leach solution; a full scale heap operation with recycle may have to incorporate a method to prevent build-up of these cations after several cycles.

Table 7: Cumulative extractions of base metals over 88 days

Columns Temperature Cu Ni Co Fe

°C % % % %

1 65 91.08 98.53 83.54 38.39

2 70 65.45 96.94 86.06 30.41

3 75 85.41 97.95 82.71 31.22

4 80 56.76 93.00 76.83 46.47

Example 3 - Cyanide heap leaching of bio-heap leached concentrate Aim

To determine the effectiveness of a cyanide solution in leaching PGMs from the residual material of a thermophilic bioleach process, following high extraction of the BMs.

Methods Two samples of concentrate residue, one a combination of residue from columns 1 and 3 of the bioleach experiment (Column A), and the other from column 4 of the bioleach experiment (Column B) were leached with 0.1 M of cyanide solution in packed columns. The combined sample in Column A weighed 500 g, whereas the one in Column B, which only had concentrate from column 4 of the bioleach experiment, weighed 290 g. As a result this column only, ran two-thirds full. The rest of the concentrate was used to perform the fire assays, XRD and MLA analyses referred to in Example 2. The assay grades of the samples were as follows:

Table 8: PGM head grade of concentrate samples

Columns Pt Pd Au Rh Ru

g/t g/t g/t g/t g/t

A 15.5 21.5 1.68 1.05 0.23

B 18 25.3 1.86 1.7 0.74

Table 9: Base metal and gangue element head grade

Total

Columns Cu Ni Co Fe Mg Al Ca Si Cr

S

% % % % % % % % % %

A 0.26 0.08 <0.05 15.8 6.32 0.79 1.72 19.0 0.10 4.76

B 0.89 0.32 <0.05 14.0 6.58 1.14 1.52 20.3 0.12 4.86

The samples were made into slurry, coated onto granite pebbles and packed in columns as before. They were leached at a flow rate of 1 l/day. The columns were operated at a temperature of 50°C and aerated at a rate of 150 ml/min. The cyanide solution was recycled for 7 days after which it was exchanged with fresh solution. Samples were withdrawn at various intervals for ICP analysis of precious metals, BMs and gangue elements. After 45 days of leaching the extractions achieved were determined by the metals in solution. Results and Discussion

Figure 3 shows that although severely lagging behind the complete and near complete extraction levels of Pd and Au (Table 10), the Pt extraction curves were continuing to increase. This suggests that perhaps over a longer period of time a higher extraction could be achieved. It was observed that in the first instance it took 7 days of recycling the solution to reach maximum Pt extraction, but after that it only took 4 days of recycling the solution. In the second week it appeared that back precipitation of the Pt occurred after 4 days and in subsequent weeks the Pt extraction level stayed constant after 4 days. This may also be a case of cyanide depletion at this stage (after 4 days).

Table 10: Precious metals extractions after 45 days

Columns Pt Pd Au

% % %

A 34.25 96.53 63.42

B 32.21 92.52 97.32

The biggest problem encountered was the considerable presence of sulphur, which complexed to form thiocyanate complex(SCN ^' ), at levels of up to 5000 ppm. The thiocyanate was identified by HPLC and represents a considerable amount of the cyanide consumed. Based on the MLA analysis results in Example 2 it was concluded that the source of the sulphur was the jarosite/sulphates as most of the sulphides had been leached and there was little or no elemental sulphur.

Table 11 : Extraction of major BMs

Columns Cu Ni Fe Co

% % % % 18.32 32.27 0.00

16.94 16.05 0.00

Although the percentage extractions appear low, especially considering how much copper and nickel was extracted in the bioleach; the concentration levels exceeded 100 ppm in the first 14 days. Marsden and House (2006) report that this amount of copper usually has a negative influence on gold recovery via adsorption to carbon, and it may influence PGM recovery similarly. Nickel on the other hand displays less ability to adsorb to carbon than copper so this amount for nickel may not cause problems during the precious metal recovery stage (Marsden and House). It appears that aerating the columns has aided the dissolution of the copper and nickel, but this was only for the first 14 days; after that the levels dropped to the 20-50 ppm range it has no negative impact thereafter. It is postulated that perhaps the initial high levels could have been attributed to more soluble minerals, such as oxides, leaving the less soluble sulphides thereafter (Habashi 1999). This can perhaps be remediated with a longer acid wash before commencing the bioleach process. However, it is encouraging that the iron extraction remained low and, as before, this is attributed to the formation of jarosite via the bioleach process.

Table 12: extraction of major gangue elements

Columns g Al Si Ca Cr

% % % % %

A 0.01 0.10 0.00 0.05 0.78

B 0.00 0.15 0.01 0.05 1.38

As expected from the literature (Drew 1972) the dissolution of gangue elements was relatively insignificant, and did not impact the leaching of precious metals. The low Pt extractions have been attributed to their occurrence as sperrylite. Separate test work will determine their impact on recovery of the precious metals. Example 4 - Cyanide leach in a stirred tank reactor

Aim (Test 1)

As Pt is the priority element for extraction in this ore body, further proof was needed to determine that it can be extracted via the cyanide leach over a longer period of time.

Methods

A sub-sample of 100 g of residue from Column B was leached in 0.5 M of sodium cyanide solution in an Erlenmeyer flask at a pH of 11 for 21 days. The temperature was kept constant at 75°C using a water bath. Slightly more aggressive conditions were used than in the previous cyanide leach test to obtain results in a shorter period of time and hopefully obtain maximum extraction without resorting to high pressure and temperature leaching.

Results and Discussion

Of the 68 % Pt left in the concentrate the new test only achieved a further 17 % extraction. Proportionally and on a cumulative basis this would count for less than an additional 1 % on the 32 % extracted in the heap leach. An MLA analysis on a sample of the residual material from the cyanide heap leach indicated that 78 % of the remaining Pt was sperrylite (PtAs ₂ ). It also showed that at this stage the PGMs were 91 % liberated. Looking back at Figure 3 it is clear that the more soluble Pt minerals leached out early leaving sperrylite which has proven to be refractory to cyanide leaching under mild conditions of temperature.

Aim (Test 2) To understand further the impact of BMs on a precious metals cyanide leach on this ore, in comparison to the literature on the same subject matter. There is popular consensus on the need to extract BM minerals before a cyanide leach; the test will quantify the value of this pre-cyanide leach stage on the Platreef ore.

Methods

Three samples of concentrate were leached in cyanide solution for four days in identical manner to the test conducted in Test 1. The concentration of sodium cyanide used in this case was 0.1 M and the three samples differed as follows:

Sample 1 : A sample of untreated concentrate

Sample 2: A sample from the bioleach process in Example 2

Sample 3: A sample pre-treated with an acidic mixture of 40 g/l H ₂ S0 ₄ and

30 g/l HN0 ₃ in which a considerable amount of the Cu and Ni had been leached

The percentage extractions achieved from the leach test are presented below:

Table 13: Percentage extractions

Pt Pd Au Rh Ru

% % % % %

Sample 1 10.30 35.06 94.98 26.41 39.61

Sample 2 20.35 36.88 100 44.62 54.39

Sample 3 51.10 94.17 100 35.24 18.07

The extraction levels of gold are not deterred by the presence of BMs. However removal of these elements has had a positive influence on the PGMs specifically Pt and Rh. At first glance it may seem that the pre- treatment used on sample 3 is the superior method to the bioleach process resulting in higher Pt and Pd extractions; but the fact is that the concentrations of the elements in solution were identical; Pt at 700 ppm and Pd at 1500 ppm. The concentrations of Pt and Pd in sample 2 were twice as high as those in sample 3 (due to the aggressive pre-leach) hence it reported lower percentage extractions. It is also noted that the type of pre-treatment has heavily impacted the Pd extraction. The oxidative acidic leach appears to have had a better result than the bio-treatment in this instance. However in the heap leach (Table 10) the Pd extractions were identical to Sample 3. It is likely that the refreshing of cyanide solution overcame an equilibrium constraint that was experienced in Sample 2.

Conclusion

This first set of experiments has shown that a two stage heap leach process, consisting of a first stage of bioleaching using thermophiles at 65°C to extract BMs, followed by a cyanide heap leach to extract precious metals, is a potential method to accompany the standard process route for extracting PGMs from a platreef ore. This is on condition that two obstacles are overcome; the first being the formation of large amounts of jarosite during the bioleach stage, which is indirectly the source of sulphur which consumes cyanide by formation of thiocyanate. The second being the introduction of a hydrometallurgical pre-treatment stage before cyanide leaching to convert the sperrylite to a cyanide soluble compound.

Example 5 - Heap bioleaching on concentrate: Experiment 2

Based on the results of experiment 1 , this experiment was conducted with the aim of achieving similar BM extraction levels but with less jarosite formation by manipulating feed composition; specifically the amount of iron and the pH (by adjusting he amount of acid in the feed).

Methods The procedure, process conditions, culture of microorganisms and sampling procedures used were identical to the ones detailed in Example 3, with the exception that a temperature of 65°C was used in all columns and the feed composition across the columns varied as follows:

Column 1 : Was a repeat of column 1 from experiment 1 , the feed solution consisted of 2 g/l Fe (1 g ferrous and 1 g ferric) and 10 g/L H ₂ S0 ₄ . After 43- 46 days the experiment would be stopped to see, by XRD analysis, if any jarosite had formed. Also important to note is that in the previous experiment most of the Cu, Ni and Co had been leached at this point. The sample would then be leached with cyanide in a stirred tank reactor to observe the levels of sulphur as well as the precious metal extraction levels. The results would be compared with the first set of stirred tank tests conducted on the residue from the bioleach column run under similar conditions. The hypothesis in this case is that if all oxides and significant copper is extracted; the remaining sulphides may not interfere significantly in the cyanide leach.

Column 2 and 3: Were run with 0.5 g/l ferrous, no ferric; with column 2 having 10 g/l H ₂ S0 and column 3 having 20 g/l H ₂ S0 ₄ . These tests reduce the amount of iron right at the beginning in case jarosite formation starts earlier in the experiment. The higher acidity in column 3 is a contingency against iron precipitation.

Column 4: Was run with the standard feed solution of 2 g/l Fe (1 g ferrous and 1 g ferric) and 10 g/L H ₂ S0 ₄ for 43-46 days after which the solution was changed to 0.5 g/l ferrous, no ferric and 20 g/l H ₂ S0 ₄ . This was to determine if the remaining base metals (Cu and Co) can be completely extracted without forming jarosite after 40 days.

Results and discussion ICP analysis of leachate samples and fire assays of the residual concentrate material were used to calculate percentage extractions which were as follows:

Table 14: Percentage extractions achieved

Columns Duration Cu Ni Co Fe days % % % %

1 (experiment 1 ) 88 91.08 98.53 83.54 38.39

1 40 79.98 92.84 93.67 46.00

2 89 90.06 98.57 97.46 70.99

3 89 93.13 99.07 98.56 90.90

4 89 93.71 98.85 98.67 86.28

As per the plan, Table 14 shows the extractions in columns 2-4 match those achieved in experiment 1. However the main aim was the reduction of jarosite and sulphur compounds. An XRD analysis and LECO combustion test showed the following results:

Table 15: XRD Results

Duration Total Elemental

Columns Jarosite

of leach Sulphur Sulphur days % % %

1 (experiment 1 ) 88 35.44 4.40 0.0

1 40 1 1.89 5.57 3.26

2 89 16.23 2.57 0.4

3 89 21.03 4.05 1.59

4 89 18.08 3.07 0.83

Table 15 shows that overall Column 2 has the least amount of sulphur compounds. Additionally Column 1 shows that formation of jarosite is unavoidable. The data shows that Column 1 of this experiment was a successful repeat of Column 1 from experiment 1. The extractions for the BMs and the extraction trends were similar at the 40 day mark. It is therefore concluded that if allowed to go beyond 40 days, the jarosite formation in Column 1 would have eventually reached the levels that were attained in the first experiment. Figure 4 shows the Fe extraction curves calculated and drawn using data from the analysis of the effluent leachate samples from the columns. These amounts were calculated based on the Fe in solution minus the initial amounts in the feed. Even though the curves for columns 2-4 show near complete leaching of the Fe, the XRD analysis and fire assays confirm the presence of Fe in the residual material (mainly in the form of the jarosite) showing that jarosite formation is unavoidable.

Conclusion

Jarosite formation in this heap bioleach process cannot be avoided entirely but can be minimised by using less iron in the feed. In a full scale operation perhaps no Fe would be required in the feed as at a large scale the Fe leached from the sulphides would be enough fuel for the process.

Example 6 - Accelerated chemical leaching of ore

Aim

The purpose of this experimental program was to demonstrate the viability of using a two stage leach process to extract BMs and precious metals from the Platreef ore in the form of coarse particles. Equally important was to achieve this in a shorter period of time than would be possible using a bioleach process as the first stage leach.

It was decided that this would be achieved using a chemical leach at elevated temperature as the first stage leach to extract BMs, followed by cyanidation to extract the precious metals. The leaching would be conducted in a packed column. Ore: receipt and preparation

The crushed ore was received in 8 bags of 20 kg each. It was prepared

from a drill core sample crushed with a high pressure grinding roller

(HPGR) at a pressure of 2.5 N/mm ² . The ore was thoroughly mixed and

split into 4 kg samples using a 2-way riffle splitter. A fire assay of the ore

revealed it had the following head grades:

Table 16: Precious metals grade assay

Total (6E) Pt Pd Au Rh Ru Ir

g/t g/t g/t g/t g/t g/t g/t

4.93 1.94 2.24 0.22 0.20 0.16 0.11

Table 17: Major base metals and gangue elements

Cu Ni Fe Co Mg Al Ca Si Cr Total S

% % % g/t % % % % % %

0.16 0.30 8.70 131.6 14.2 2.5 4.39 22.3 0.32 0.94

Base metal leaching Methods

Two samples of 4 kg each of coarse ore crushed using an HPGR at

pressure 2.5 N/mm ² , were screened to size fraction -12.5mm+106pm, acid

agglomerated and packed in columns. One column (A1 ) was leached using

40 g/L H ₂ S0 ₄ and 30 g/L HN0 ₃ while the other (A2) was leached using 40

g/L HN0 ₃ and 15 g/L H ₂ S0 ₄ . The columns were operated at a temperature

of 85°C. The solution was pumped into the columns at a rate of 2 L/day and

recycled for 7 days at which point samples obtained over various intervals

over that period were analysed using AAS for Cu, Ni and Fe. After 7 days of leaching the extractions obtained were very poor, less than 1 % of Cu in each column, although A2 produced a significantly better Ni extraction. As a result the solution combination in column A2 was doubled to 80 g/L HN0 ₃ and 30 g/L H ₂ S0 ₄ and used in both columns. Additionally, it was noted that some back precipitation of both Ni and Cu occurred after 4 days (Figures 5 and 6), hence the solution was recycled for 3 days and then replaced with fresh solution.

On completion of the experiment, the columns were emptied, the ore was thoroughly washed with water to remove residual acid and sub-samples of the ore from each column were obtained for fire assays to determine the extractions of Cu and Ni achieved. The bulk of the remaining ore would be leached with cyanide to extract precious metals.

Results and Discussion

After 46 days of leaching the results achieved are reported in Table 18. It must be noted that the results reported in Table 18 were calculated from fire assays on sub-samples from the ore, before and after leaching; while the leach curves in Figures 5 and 6 are from the AAS analysis on leachate solution samples hence the small difference in final extractions reported in Table 18 and Figures 5 and 6.

Table 18: Percentage extractions after 45 days

Columns Cu Ni

% %

A1 83.71 84.69

A2 81.37 81.83

The combination of HN0 ₃ and H ₂ S0 ₄ has shown to be an effective method of leaching Cu and Ni from this ore. This combination is especially effective when the HN0 ₃ to H ₂ S0 ₄ ratio is around 2.3:1 by wt concentration. This success can also be attributed to other factors apart from the aggressive nature of the reagents, such as the high temperature and the various micro cracks caused by crushing via an HPGR. This likely lead to better contact between lixiviant and the minerals considering that most industrial and laboratory heap leach operations on crushed ore achieve only up to 60 % extraction (Gupta and Mukherjee 1990). Apart from accelerating the process, it is not known if the temperature contributed to chalcopyrite leaching by way of preventing passivation as this combination of acids is known to leach sulphide minerals of copper and nickel even at ambient temperatures.

Example 7 - Cyanide leaching of residual ore Methods

For the cyanide leaching in columns, the samples of ore from the accelerated leach for BMs were each agglomerated using ordinary Portland cement at a ratio of 5g/kg of ore and water at 8 wt %. The agglomerated samples were packed in two columns (A1 and A2 corresponding to the base metal leach) and left over night to dry. The columns were operated at 60°C, at an aeration rate of 150 ml/min and solution feed rate of 1 L/day. The sodium cyanide solution was recycled for 7 days before being replaced with fresh solution and samples are withdrawn on 3-4 day intervals for ICP analysis of precious metals, base metals and gangue elements.

After 77 days of leaching poor extractions of platinum were achieved so the concentration of cyanide was increased in both columns from 0.1 M to 0.3 M. Additionally hydrogen peroxide in a concentration of 0.1 M was added to the solution fed to column A2. The head grade of the leached ore was as follows:

Table 19: Precious metals grade assay

Columns Pt Pd Au Rh Ru Ir g/t g/t g/t g/t g/t g/t

A1 1.54 1.60 0.21 0.12 <0.1 0.51 A2 1.53 1.53 0.17 0.15 <0.1 0.48

Table 20 : Major base metals and gangue elements

Columns Cu Ni Fe Co Mg Al Ca Si Cr Total

S

g/t g/t % g/t % % % % % %

A1 349 598 6.13 <10 10.7 1.18 4.31 24.3 0.29 1.16

A2 392 697 6.41 <10 1 1.4 1.25 4.27 23.7 0.30 1.03

After 155 days of leaching the following results were achieved: Results and discussion

The platinum leach curves (Figure 7) were drawn and percentage metals leached (Table 21 ) were calculated using the results from the ICP analysis of samples from the effluent leachate solutions from the columns.

Table 21 : Precious Metals leached in 155 days

Columns Pt Pd Au Rh Ru

% % % % %

A1 19.09 46.74 86.37 28.01 34.41

A2 15.21 34.40 88.09 20.01 32.68

A familiar trend is immediately apparent with regards to the leaching of platinum. Comparing Figures 3 and 7, it is clear that a certain portion of the platinum leaches out early and then the rate of extraction drops though it continues to increase, bit by bit. Given that the Pd and Au have leached significantly better, it is unlikely to be a liberation problem, but rather, as concluded from the leaching on concentrate samples, a matter of the mineralogy of the remaining platinum. It is apparent that the remaining Pt is in the form of sperrylite. Similarly the Pd is likely to be in the form of an arsenide given that it has not leached as well as it did in the cyanide heap leach on concentrate (Table 10).

Table 22: Base Metals leached in 155 days

Columns Cu Ni Fe Co S

% % % % %

A1 10.61 4.50 0.18 23.81 100.00

A2 4.64 1.86 0.11 12.25 100.00

Table 23: Major Gangue elements leached in 155 days

Columns Mg Al Si Ca Cr

% % % % %

A1 0.00 0.00 0.01 0.82 0.00

A2 0.00 0.00 0.00 0.36 0.00

Both the BMs and the gangue elements have not shown any form of significant leaching during cyanidation, especially considering the starting grades of the BMs (Table 20). Additionally, at no time did the BM concentrations exceed the potentially problematic concentration of 100 ppm.

Final Conclusion

A two stage heap leach process in which a bioleach first extracts BMs followed by a cyanide leach to extract precious metals is a potential method to accompany the conventional method of processing PGMs, to achieve full value of the Platreef ore. An initial acid wash and bioleach to extract high levels of BMs has shown to be beneficial to Pt and Rh leaching.

Abbreviations and acronyms

AAS Atomic Absorption Spectroscopy BM Base Metals (Cu, Ni, Co)

BMR Base Metals Refinery

HPGR High Pressure Grinding Roll

HPLC High Pressure Liquid Chromatography

ICP Inductively Coupled Plasma spectroscopy

ICP-MS Inductively Coupled Plasma with Mass Spectrometry

ICP-OES Inductively Coupled Plasma with Optical Emission

Spectroscopy

MLA Mineral Liberation Analysis

PGM Platinum Group Metals, including Au and Ag by association

PGE Platinum Group Elements, including Au and Ag by association

pH negative logarithm (base 10) of the molar hydrogen concentration

UG2 Upper Group Two (a PGM mineralised reef found in the

Bushveld Igneous Complex)

XRD X-Ray Diffraction

XRF X-Ray Fluorescence

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