PGMs usually occur as discrete minerals, as dilute solid solutions in sulphide minerals, or are associated with silicates and/or chromitites. The processing of these materials consists of a combination of several unit operations, which ultimately produce individual PGMs-platinum, palladium, rhodium, ruthenium, iridium and osmium-plus gold and silver.

Currently PGM ores containing varying amounts of base metals, such as copper, nickel and cobalt, are crushed, milled and floated in flotation cells using reagents typically used for the recovery of base metal sulphides, to produce a low grade PGM concentrate. This flotation concentrate is dried and then smelted in an electric arc furnace where the PGMs are collected in the matte phase.

The molten furnace matte is transferred to a converter where flux is added and air blown into the bath so as to remove the iron as a slag. During the blowing process, the majority of the'sulphur in the furnace matte is converted to sulphur dioxide and emitted as a gas.

The converter matte produced from the converting stage is either granulated or slow cooled. The granulated matte is milled and treated in a whole matte leach base metal removal circuit to separate nickel, copper, cobalt and iron from a PGM rich residue or concentrate. The base metals are sometimes refined to metal and in other plants produced as crude cathode and a semi pure nickel sulphate crystal, which contains iron and cobalt. In cases where the matte is slow cooled, the magnetic fraction is processed as described above, while the non-magnetic fraction, which is the larger mass fraction, is treated to recover nickel and copper as sulphates leaving iron in an insoluble fraction. The iron residue can contain up to 20% of the original PGMs emanating from the converter step.

The high grade PGM concentrate is then treated in a precious metal refinery where the individual PGMs are separated and produced in pure metallic form.

A simplified flow sheet of the above process is shown in Figure 1. This conventional processing route for PGMs has its limitations. It is considered suitable for sulphide containing ores from which relatively low quantities of flotation concentrates are produced with high recoveries of PGMs that can be treated economically through a smelter. However, ores that have been oxidised produce greater amounts of flotation concentrates to achieve the same PGM recoveries and this can lead to uneconomic smelting operations.

Concentrates containing relatively high levels of chromite also adversely affect smelting operations, resulting in high costs and lower recoveries.

The erection of new smetters to process PGM containing f) otation concentrates has been significantly curtailed in recent years and has all but now been stopped due to very strict environmental legislation and prohibitively high capital costs. This means that any new operation will have to ship flotation concentrates excessively large distances to existing smelters, at high cost, leading to increased metal pipelines.

Environmental legislation is also forcing existing smelters to continually reduce the amount of sulphur dioxide being emitted to the atmosphere. This leads to on sulphur dioxide (SO2) fixation plants and often with no revenue return. Significant benefit may therefore accrue were the smelting step to be eliminated.

SUMMARY OF THE INVENTION A process of recovering platinum group metals (PGMs) from a pregnant solution or leachate containing PGM values and base metals, includes the steps of: a) non-selective precipitation of the PGM values and base metals to form an insoluble precipitate comprising the PGM values and base metals ; b) selective leaching of the precipitate to form (i) a leach solution containing the base metals and a residue containing the PGM values, or (ii) a leach solution containing the base metals and PGM values and a depleted residue; and c) recovery of the base metals from the leach solution of b) (i) or b) (ii) and recovery of the PGM values from the residue of b) (i) or the leach solution of b) (ii), depending on the selective leaching in step b).

The pregnant solution or leachate containing the PGM values and base metals is preferably a cyanide solution or leachate.

The non-selective precipitation of the PGM values and base metals in step a) is preferably carried out by controlled reduction of the pH of the pregnant solution or leachate to within the range of about 0 to about 5, preferably about 1 to about 3, in particular about 2.

In one embodiment of the invention, the selective leaching in step b) is a pressure leaching step in which the PGM values remain in the residue. The PGM values are typically recovered from the residue by a fusion step or a further leach step to produce a concentrate rich in PGM values. The base metals are recovered from the leach solution, preferably through precipitation. Alternatively, the pressure leaching step may be carried out by first carrying out a caustic leach followed by an acid leach, in which case the PGM values may be recovered from the residue by, for example, a reduction leach step to produce a high grade PGM concentrate. In an alternative embodiment of the invention, the insoluble precipitate of step a) is fumed with acid, followed by selective leaching to form a leach solution containing both the PGM values as anions and the base metals as cations, and recovery of the PGM values on an ion exchange resin, which is preferably incinerated to form an ash rich in PGM values or eluted to recover the PGM values. The base metals are preferably recovered from the leach solution by precipitation.

In a further alternative embodiment of the invention, the insoluble precipitate of step a) is first calcined and then selectively leached to remove the base metals, producing a PGM rich residue or concentrate. Again the base metals are preferably recovered from the leach solution by precipitation.

Any PGM values remaining in solution after the non-selective precipitation in step a) may be recovered by passing the solution through an ion exchange resin, as discussed above. Any base metals remaining in the solution may also be recovered, preferably by precipitation.

BRIEF DESCRIPTION OF THE DRAWINGS The invention will now be described in more detail, by way of example only, with reference to the accompanying drawings, in which : Figure 1 is a flow sheet showing a prior art method of recovering PGMs ; Figure 2 is a flow sheet of an embodiment of the process of the present invention; Figure 3 is a flow sheet of an alternative embodiment of the process of the present invention; Figure 4 is a flow sheet of a further alternative embodiment of the process of the present invention; Figure 5 is a flow sheet of another alternative embodiment of the process of the present invention ; Figure 6 is a graph that shows the effect of using NaHS over a range of pHs to precipitate PGMs and base metals ; Figure 7 is a graph that shows the effect of using H202 over a range of pHs on recoveries of PGMs and base metals ; Figure 8 is a graph that shows the effect of a combination of H202 and Na2S205 at varying pHs on recoveries of PGMs and base metals ; Figure 9 is a graph that shows the effect of using sulphuric acid over a range of pHs on recoveries of PGMs and base metals from cyanide leachate ; Figure 10 is a graph that shows the loading profile of platinum on the resins in a five-column pilot ion exchange circuit ; Figure 11 is a graph that shows the loading profile of gold on the resins in a five-column pilot ion exchange circuit ; Figure 12 is a graph that shows the effect of temperature of calcination on the leaching extraction of PGMs and base metals ; and Figure 13 is a flow sheet of a pilot campaign for treating a pH2 precipitate.

DETAILED DESCRIPTION OF PREFERRED EMBODIMENTS The present invention provides a process to recover PGMs and base metals from a solution or leachate containing them, in particular a cyanide solution or leachate.

The cyanide leachate is produced through a known cyanide leach process. In this process, a flotation concentrate or an acid pressure leach residue material bearing platinum group metals such as platinum, palladium, rhodium, ruthenium, iridium and osmium, other precious metals such as gold and silver, and various base metals such as copper, nickel and cobalt, is contacted with a cyanide leach solution to solubilise the PGMs and the base metals. The resultant cyanide leachate may, if desired, be concentrated to form a concentrate rich in PGM and base metal complexes.

Referring to Figure 2, in a first stage of the preferred process of the invention, the cyanide leachate is acidified with diluted sulphuric acid in a pH range of 0-5, 0 and conducted at a temperature in the range of ambient to 95°C. The acidification is preferably done using 1: 1 sulphuric acid at around pH 2 and 45°C. The acidification step can take up to 6 hrs.

The precipitation of the PGMs and base metals during acidification may be aided by controlled rates of addition of reagents such as sulphides in the form of sodium sulphide and sodium hydrogen sulphide or oxidants such as hydrogen peroxide and salts of peroxomonosulphuric acid (Caro's acid).

After acidification, the precipitated metals are recovered from solution by solid-liquid separation. Any cyanide released as hydrogen cyanide can be recovered in a scrubber.

The solution from the solid-liquid separation is passed through ion exchange columns where it is contacted with a resin that has functional groups that have stronger affinity for PGM complexes than the base metal complexes e. g. certain ammine groups. The PGM complexes bind to the resin in preference to certain base metal complexes of nickel, copper, cobalt and iron.

The ion exchange columns are typically arranged in a"lead-trail"configuration. As the lead column becomes exhausted by the loading of the PGM and base metal complexes, the PGM complexes displace some of the base metals as additional feed is provided to the lead column. The trail column (s) captures these base metal complexes, as the case may be, and so a partition is achieved between the PGMs and the base metals. Eventually the base metals are displaced from the ion exchange resin and pass out with the raffinate.

To recover the PGM values from the lead PGM column, a number of options are possible. The resin may be removed and ashed to produce a PGM product.

Alternatively the resin may be eluted using an appropriate eluant to remove the PGM values.

To assist with the recovery of gold (Au), an ion exchange column loaded with a resin such as activated carbon is typically installed in the system.

Once the solution has passed through the ion exchange columns, the residual base metal values are preferably precipitated from solution using reagents such as caustic soda and/or sodium carbonate and sodium sulphide. The precipitated base metal products can then be recovered from solution by solid-liquid separation.

Alternative methods such as ion exchange, solvent extraction and electrowinning can also be used to recover the base metal values.

The filtered precipitated residue from the initial step of acidification, also referred to as the acidification residue, may be repu ! ped using fiitrate from a ! ater fixation step.

Sulphuric acid is typically added to the repulping step. Any off-gas is scrubbed to recover hydrogen cyanide (HCN), which is recycled to the cyanide leaching stage.

The slurry from the repulping step is fed to a high-pressure vessel, typically a multi- compartment autoclave, where it is heated to a temperature of 140-220°C, preferably between 175 and 185°C. Oxygen is added counter currently to flush off any HCN into the vent from where it may be scrubbed. The autoclave is steam heated, directly or indirectly, and the pressure in the autoclave is maintained typically at 1500 kPa (g). The pulp density can be such as to yield a base metal concentrate varying from 10 to 180 g/l combined base metals, preferably 40-80 g/l.

The autoclave retention time is between 1 and 6 hours, preferably 3 hours. The discharge from the autoclave is cooled to approx 85°C and flocculants such as Flocculant M351 added to aid settling before solid-liquid separation such as thickening and pressure filtration. After filtration, a solid product stream containing the PGMs and a filtrate stream are recovered.

There may be small quantities of precious metals in the filtrate stream and these can, after conditioning with a small quantity of chloride ions, be recovered on a simple ion exchange resin. The base metal complexes are weakly held by the ion exchange resin and ultimately get displaced The chloride conditioning process employs typically 5 to 10 g/i of chlorides at 90 to 95°C for approximately one hour.

The raffinate from the ion exchange step can be treated by a variety of means to recover the copper. These include : Copper solvent extraction to remove the copper to an electrowin step ; - Subject to the copper nickel ratio in the leachate, metathetically exchanged for nickel-copper sulphide matte at temperatures between 70 and 180°C, preferably 130-140°C. The copper sulphide fraction can be treated to recover copper by a leach-electrowin process ; and -Electrowinning and electrostripping of the copper to concentrations below 0.3 g/i copper.

The remaining nickel and cobalt in the raffinate stream after recovering the copper can be recovered by a variety of means. These include : - Sulphide precipitation either with hydrogen sulphide gas or sodium sulphide/hydro sulphide; - Crystallisation as nickel sulphate hexahydrate; and - Hydroxide/carbonate precipitation with magnesium hydroxide, sodium hydroxide, sodium carbonate.

The solids from the autoclave are dried to approximately 3-10% moisture using typical conventional dryers. The partially dried material is blended with concentrated sulphuric acid, typically in a cone mixer, which is heated indirectly to temperatures up to 350°C, preferably 280-310°C. The off-gas from this high temperature atmospheric leach step can be scrubbed to recover any sulphuric acid vapour. The reaction temperatures can be maintained for periods of up to 1'/hrs.

The slurry from the high temperature atmospheric leach step is drained into a dilution vessel where it is diluted to typically 100g/l free sulphuric acid. The leach is maintained at temperatures up to 95°C for typically 1 hr and then filtered using any suitable solid-liquid separation. The solid is washed before drying to produce a final concentrate rich in PGM values, typically assaying 15-85% PGM plus gold and silver, which may be refined in a conventional precious metal refinery to recover the PGM values.

The filtrate and wash liquor after recovering the PGM concentrate may be recycled to the repulping stage of the acidification residue.

In an alternative embodiment of the invention, pressure leaching of the acidification residue is conducted by first carrying out a caustic leach (sometimes referred to as 'oxydrolysis') followed by an acid leach as shown in Figure 3. This alkali-acid combination leach is conducted to destroy the cyanide complexes and solubilise the base metals for subsequent recovery, typically as described above. The leachate from solid-liquid separation is passed through a lead-trail ion exchange for the recovery of the precious metal values. The residue containing the PGM values may be reduction leached using a reductant such as formic acid to produce a high grade PGM concentrate suitable as a feed material to a conventional precious metal refinery.

A further alternative embodiment, of the invention comprises a sulphuric acid fusion of the acidification residue followed by a chloride leach using reagents such as chlorine, hydrochloric acid and sodium chlorine or a combination of some or all the reagents (Figure 4). The fusion/leach is conducted to produce a solution containing both the PGM values (as anions) and base metals (as cations). The PGM values are recovered by passing the solution through anionic ion exchange resin columns and followed by elution or alternatively ashing the resins. The base metal values are recovered by any suitable known method such as ion exchange, solvent extraction, precipitation or electrowinning.

In yet a further alternative embodiment of the invention, calcining of the acidification residue is conducted at a temperature between 250°C and 800°C, but preferably between 400-600°C, prior to an acid leach (Figure 5). The acid leach is conducted to solubilise the base metal values for subsequent recovery by any suitable known method. The PGM values remain in the residue and may be reductively leached to produce a high grade PGM concentrate.

Illustrative examples based on test work will now be presented to exemplify the present invention and should not be construed to limit the inventive method in any way. The test work is presented below in a series of tests which have been conducted either using cyanide leachate, acidification solution or acidification residue.

Example 1 : Precipitation with Sulphide at Raised Temperature A test was conducted using NaHS to precipitate value metals from cyanide leachate at a raised temperature.

A cyanide leach solution was prepared by calcining underground flotation concentrate at 400°C for 2 hrs in a muffle furnace and then leaching using 0,2% w/w sodium cyanide solution at 60°C for 48 hrs, maintaining the pH at 9,5 with hydrated lime.

115% stoichiometric level of 20% w/v NaHS was added to 200moi cyanide leach solution. H2SO4 was added to adjust the pH to 1 and the temperature maintained at 90°C. The solution was sampled after 120 minutes. The test results are summarised in Table 1. It can be seen that good recoveries were achieved for Pt, Pd, Au, Cu and Co.

TABLE 1: Recoveries into Residue using NaHS Metal Pt Pd Au Cu Ni Co Recovery 97.5% 99. 7% 99.5% 99. 3% 11. 8% 92. 8% Example 2 : Precipitation with Sulphide at Ambient Temperature A second series of tests was conducted on the cyanide leachate using NaHS over a range of pHs at ambient temperature to precipitate the value metals.

A cyanide leach solution was prepared by calcining a finely ground'open cut' flotation concentrate at 4000C for 2 hrs in a muffle furnace and then leaching using 0, 2% NaCN solution at 60°C for 48 hrs, maintaining the pH at 9, 5.

115% stoichiometric level of 20% w/v NaHS solution was added to a series of 200ml cyanide leach solutions, H2SO4 or NaOH was added to adjust the pH in the range 0, 5-11, 5. The reaction time was 5 mins.

Figure 6 shows that good recoveries were obtained for Cu, Co, Pd and Au at low pH values. The test results have been summarised in Table 2.

TABLE 2: Sulphide Precipitation Metal Recoveries Precip Cu Ni Co Fe Pt Pd Au pH (%) (%) (%) (%) (%) (%) (%) 0.5 99.77 11.98 95.45 0.00 47.65 99.46 98.10 1.5 99.97 31.02 92.39 71.42 38.44 31.55 86. 52 3.5 99.75 2.45 4.03 46.77 1.20 0.86 15.87 5. 5 95.66 0.04 1.76 14.15 0.35 19.86 20.42 7.5 0.02 0.00 0.00 0.00 0.00 28.29 23.06 9.5 0.00 0.00 0.00 0.00 0.45 26.36 39.41 11.5 0.00 0.00 0.75 0.00 8.03 15.35 6. 87 Example 3: Precipitation using H202 and Na2S205 A series of tests was conducted using H2O2 and Na2S2O5 to precipitate the value metals from cyanide solution over a range of. pH values.

A cyanide leach solution was prepared by calcining underground flotation concentrate at 400°C for 2 hrs in a muffle furnace and then leaching using 0, 2% NaCN solution at 60°C for 48 hrs maintaining the pH at 9, 5.

Hydrogen peroxide (H202) was added to the cyanide leach solution in the ratio of 1,5mg/mg of CNwAD at ambient temperature and sulphuric acid was added to maintain the pH at set points between the ranges of 1-7. The reaction time was 120 min.

Sodium metabisulphate (Na2S205) was then added in the ratio of 6 mg/mg of CNwAD.

Sulphuric acid was added to maintain the same pH. Again the reaction time was 120 min.

TABLE 3: Metal Recovery with H202 and Na2S205 Treatment Metal removed from solution % Test pH Pt Pd Au Cu Ni Co H2O2 0,9 99,6 99,6 4,2 91,2 81,1 99, 7 Na2S205 1, 0 0 48,3 6,2 29,0 4,6 0 Overall 99,6 99,8 10,2 93,8 82,0 99, 7 H2O2 3,2 97,9 97,9 1,6 98,9 93,8 91,8 Na2S205 3,0 25,0 50,0 3,9 76,4 15, 9 18, 8 Overall 98,4 98,6 5,5 99, 7, 94,8 93,1 H2O2 5,1 42,7 52,1 0 99,5 86,3 44,0 Na2S205 5,6 0 0, 0 0 1,1 13,3 Overall 42,7 52,1 0 99,5 86,4 51,5 H2O2 7,1 21,8 31,5 0 95,3 71,8 15,4 Na2S20s 7,1 2,8 2,5 8,2 0 0 12,3 Overall 24,0 33,2 8,2 95,3 71,8 25,8 Table 3 and Figures 7 and 8 show the effect of H202 only and a combination of H202 and Na2S205, respectively, over a range of pHs on recoveries of PGMs and base metals. It can be seen that with the exception of Au, good recoveries of PGMs and base metals were obtained at pH values at and below 3.

Example 4: Acidification of Cyanide Leachate In a series of tests, a cyanide leachate of a PGM containing concentrate was acidified with 50% H2SO4 over a range of pHs from 0 to 5.

Cyanide leach solution was prepared by calcining underground flotation concentrate as per the previous examples. The alkaline cyanide leachate was warmed to 45°C and while gently agitating, 1: 1 H2SO4 was added slowly. Once the required pH was reached, the solution was allowed to stand while a precipitate formed. The pH was checked and readjusted to the target value. Flocculant M351 was added and allowed to stand for 3 hours without agitation. The concentration trends are shown in Table 4, and Table 5 shows the recovery of value material to the precipitate. PGM recovery to the precipitate was optimised at a pH of approximately 1.0. At pH2, the base metal residuals were at a maximum and then commenced redissolving. A similar trend was observed for the PGMs below a pH of 1.0. Figure 9 trends the precipitated component at various pH's.

TABLE 4: Effect of Variable pH. Test no pH Pt Pd Au Cu, Ni Co Fe ppm Feed 1 4. 21 5. 58 1. 95 1910 3970 17 32 Leachate 4.95 4.56 5.85 0.98 1350 2875 26 27 Feed 2 3.42 2.62 0.75 862 1240 15 1250 Residue 1 3.15 3.04 2.31 0.62 87 102 32 2. 8 Residue 2 2. 41 2. 66 2 0.67 1 1 6. 8 2. 3 Feed 3 3.3 3.53 1.3 1380 1950 50 6.8 Leachate 1 2. 2 2. 37 2. 77 1. 06 1. 56 196 1. 74 0. 22 Leachate 2 1. 95 0. 64 1.35 1 3. 53 132 0. 55 0.17 Leachate 3 1.9 0.47 0.68 1 2.22 207 0.1 0.21 Leachate 4 1.42 0.26 0.37 0.61 1.51 343 0. 15 0.29 Leachate5 1. 01 0. 01 û. 13 0. 56 2. 36 688 0. 09 0. 76 Leachate 6 0 0. 41 0. 25 0. 57 9. 76 872 0. 23 0. 57 TABLE 5: Effect of pH on value material recovery Test pH Pt Pd Au Ni Cu Co 1 0,01 86,7 92,9 56,2 55,3 99,3 99,5 2 1,01 99,7 96,3 56,9 64,7 99,8 99,8 3 1,42 92,1 89,5 53,1 82,4 99,9 99,7 4 1,90 85,8 80,7 23,1 89,4 99,8 99,8 5 1, 95 80, 7 61, 8 23, 1 93, 2 99, 8 98, 9 6 2, 2 28, 2 21, 5 17, 6 89, 9 99, 9 96, 5 7 2,41 22,1 23,7 18,5 99,9 99,9 54,7 8 3,15 11,1 11,8 10,7 91,8 89,9 <1 9 4,95 <1 <1 49,7 27,6 29,3 <1 Example 5: Recovery of PGM values from Filtrate after Acidification using Ion Exchange Resin In this test, filtrates from Example 4 were subjected to ion exchange using a single quaternary ammonium ion exchange resin and a cation ion exchange resin.

A three column circuit was employed with the first two columns charged with IRA- 402 (an anionic resin) and with the final column equipped with an activated carbon.

A total of 190 litres of pregnant liquor was passed through the circuit and the results are given in Table 6.

TABLE 6: Ion exchange on pH2 Filtrate Bed Composition (mg/L) Volumes Au Pt Pd Cu Ni Fe Co Pregnant Liquor - 0.8 1.57 1.62 <1.0 180 <1 <1 Lead Column 760 0.55 <0.01 <0.01 <1 293 1.5 <1 3714 1. 07 0. 32 0. 15 <1 433 4. 0 <1 9820 0. 62 0. 28 0. 67 <1 267 1. 2 <1 Trail Column 9820 0.47 <0.01 0.1 <1 246 2.0 <1 The composition of the resins after ashing is shown in Table 7.

TABLE 7 : Composition of Ion Exchange Resins after ashing Composition % * Au Pt Pd Cu Ni Fe Co Lead Column #1 1. 3 16.5 21.3 23.7 <1.3 <1.3 34.6 Lead Column #2 5. 6 43. 2 43. 1 4.4 <3. 1 3. 8 24.4 Trail Column 21. 7 <13 <13 <3.3 10.0 35. 0 <3.3 *Calculated from mass load net of ash The first bed column had been kept in duty beyond its optimal cut point and was loading strongly held copper and cobalt complexes. Table 7 provides confirmation that a high grade PGM ash can be derived from the resin while Table 6 confirms that low raffinate PGM values can be derived employing the resin suite of IRA-402 and an activated carbon.

Example 6: Recovery of PGM values from Filtrate after Acidification using Ion Exchange Resin In this test, the filtrate from the pH2 acidification step was passed through a series of ion exchange columns with the aim to saturate and analyse the resin in the lead column.

Cyanide leach solution was prepared in a pilot plant run where calcined underground flotation concentrate, produced from a rotary kiln heated to 400-425°C and residence time of 2 hrs, was leached with NaCN in a 5 stage cascading circuit with a retention time of 100 hrs at a pH around 9, 2 and temperature of 60°C. The solution was heated to 45°C before being pumped to a 3 stage cascading circuit with a retention time of 3-5 hours. Sulphuric acid solution of 50Qg/l H2SO4 was added to the first tank to maintain the pH at 2.

The filtered solution from the acidification step was then pumped at a flow rate of approximately 40 I/hr and 20°C into a series of ion exchange columns. The resin charges of the columns were as follows : Column 1 2 litres IRA 402 CL Column 2 2 litres IRA 402 CL Column 3 2 litres activated carbon type CQ 006-12X40 Column 4 2 litres IRA 402 CL The resin was equilibrated with 12 bed volumes of pH1 aqueous sulphuric acid prior to feeding pH2 filtrate.

After 5850 litres (2925 bed volumes) of pH2 filtrate had been processed, platinum and palladium were still loading whilst gold was being displaced and absorbed further down in the fourth column. The raffinate profiles for column 1 are shown in Table 8.

TABLE 8: Raffinate profiles for column 1. Day Pt Pd Au Cu ppb Feed (ave) 530 610 900 5200 1 330 360 680 500 2 3 5 65 100 500 4 5 20 257 500 5 5 5 563 500 6 5 5 1027 500 7 5 5 1097 500 8 5 5 717 500 9 35 68 633 500 10 43 110 653 500 11 83 87 837 500 12 72 85 930 500 13 105 115 1045 500 14 127 140 1073 500 15 147 177 102Q 500 16 17 187 207 1510 500 The loading profiles for Pt, Pd, Au and Cu are shown in Table 9.

No Co or Ni absorption was measured for both the ion exchange resin and the activated carbon.

TABLE 9: Loading profiles for column 1 DAY Pt Pd Au Cu Grams 1 0 0 0 0, 174 2 3 0,128 0,113 0,232 0,174 4 0,320 0,300 0,462 1,481 5 0,608 0,566 0,637 1,481 6 0,968 0,849 0,737 1,582 7 1,288 1,107 0,797 2,342 8 1,487 1,299 0,830 2,860 9 1,641 1,464 0,880 3,665 10 1,779 1,603 0,911 4,653 11 1, 874 1,663 0,928 5,185 12 2,029 1,760 1,070 5,396 13 2,171 1,842 1, 085 5,629 14 2, 311 1, 917 1, 138 5, 860 15 2, 413 1, 973 1, 148 5, 860 16 2, 565 2,073 1,159 5, 860 17 2, 622 2, 092 1, 073 6, 019 The load profiles for platinum and palladium are similar with the blend of IRA-402 and activated carbon employed as exchangers in the multi-column circuit. Figure 10 shows the load profile for platinum from pH2 leachates in l : he five-column lead-trail pilot ion exchange campaign. Figure 11 shows gold is not as strongly bound to the resin functional groups as platinum.

The average concentration of the resin in column 1 after the pilot run is shown in Table 10.

TABLE 10: Average composition of resin in column 1 Ppm Pt 93580 Pd 104821 Au 55480 Example 7: Recovery of PGM values from Acidification Residue by Acid Fuming and Ion Exchange A series of tests were conducted to absorb PGM values onto an ion exchange resin from a solution following an acid fusion on acidification residue.

Wet acidification residues, as prepared in Example 4, were mixed with 1: 1 H2SO4 in the ratio of 1,5 H2SO4 : 1 wet cake, boiled down for 15 mins till white fumes were evolved and diluted with water (500ml/100 gm wet cake) before chlorine gas was bubbled through at 70°C. The solutions were then progressively passed through a single column containing 10ml IRA402 resin.

A second set of tests were done by fuming with H2SO4, diluting and then adding sodium chloride (about 15g NaC1 per 100g wet cake). Again the solutions were progressively passed through the resin.

The results have been summarised in Table 11. It can be seen that selective eAraction a PGis took place over the base metals. However, Pt and Au absorbed preferentially to Pd once the resin became saturated. The analysis of the resin revealed values of Pt at 5, 83 gel and Pd at 6,68 g/l.

TABLE 11 : Ion Exchange results from Acid Fuming-Leach Filtrates Volume bv Au Pt Pd Cu Ni Co Dissolution mi Cum mg/, method Feed 1 0. 17 12.9 25. 1 8180 16400 457 H2SO4/Cl2 Raffinate 0-80 8 <0. 01 1.83 0. 59 7370 14680 417 Feed 2 < 0. 01 3.79 13. 8 6420 8560 207 H2SO4/Cl2 Raffinate 80-160 16 < 0. 01 1.20 0.20 5910 9480 223 Feed 3 1.10 10.1 22.9 5720 14300 232 H2SO4/C12 Raffinate 160-235 24 < 0. 01 0.91 0.20 5420 13400 223 Feed 4 0.20 0.20 8.89 2545 5795 215 H2S04/Cl2 Raffinate 235-310 31 < 0. 01 0.36 0.15 2670 5850 201 Feed 5 0.23 0.42 13.0 3990 3080 205 H2SO4/Cl2 Volume bv Au Pt Pd Cu Ni Co Dissotution ml Cum mg/l method Raffinate 310-385 39 < 0. 01 0.23 0. 14 3870 3430 205 Feed 6 0. 12 1.01 12.7 3060 2580 205 H2SO4/CI2 Raffinate 385-460 46 < 0. 01 0.21 0.15 3190 2650 206 Feed 7 0.12 0.11 2.77 1210 2380 105 H2SO4/CI2 Raffinate 460-535 54 < 0. 01 0.08 0.07 1520 2440 117 Feed 8 0.12 11.7 64.5 21058 23091 710 H2SO41C12 Raffinate 535-660 66 < 0. 01 0.31 1.32 19931 22021 673 Raffinate 660-765 77 < 0. 01 0. 47 2.56 22545 25112 787 Feed 9 1. 53 39.0 64. 5 12200 9600 3620 H2SO4/Cl2 Raffinate 765-845 85 < 0. 01 0. 28 0.64 13800 13800 410 Feed 10 1.92 84.0 156 25300 38300 760 H2SO4/Cl2 Raffinate 845-985 99 < 0. 01 0.23 1.51 24600 35000 720 Feed 11 0.87 58. 5 104 21900 22200 760 H2SO4/CI2 Raffinate 985-1170 117 < 0. 01 0. 26 2.27 22100 23400 770 Feed 12 1.29 32. 0 46. 0 11275 15274 432 H2SO4/Cl2 Raffinate 1170-1320 132 0. 2 0.37 5.3 11197 15172 442 Raffinate 1320-1540 154 0. 08 0. 31 9. 7 11450 15395 435 Feed 13 0.11 10.4 8.96 6360 7580 308 H2SO4/C12 Raffinate 1540-1740 174 0.05 0.10 13.3 6690 8320 325 Raffinate 1740-1940 194 0. 04 0. 08 16. 1 6380 7630 315 Feed 14 0. 75 30. 5 25. 0 10800 12000 610 H2SO4/NaCI Raffiinate 1940-2145 215 0.07 0. 14 30.7 10800 12100 611 Raffinate 2145-2315 232 0.02 0.36 43.7 11800 12400 638 Feed 15 0. 34 17. 40 29.4 15000 15550 608 H2SO4/NaCl Raffinate 2315-2515 252 < 0. 01 1.22 51.3 14800 15100 633 Feed 16 0. 10 16.9 25. 1 12000 13000 492 H2SOINaCI Raffinate 2515-2745 275 < 0. 01 0.78 31.5 12300 12300 497 Example 8 : Recovery of PGMs from High Pressure/High Temperature Leach in a Two Stage Alkaline/Acid Medium followed by Ion Exchange.

A series of tests were conducted by subjecting the pH2 acidification residue to a two-stage alkaline/acid leach at varying pressures and temperatures from 10-20 bar (1000-2000 kPa) and 150-200°C, respectively.

Acidification residue was prepared as described in Example 4 and slurried with water in an autoclave and then steam heated to the required temperature. A weighed amount of caustic soda (NaOH) was added. The autoclave was pressurised to the set pressure and left to react for 2-4 hrs. A set volume of H2S04 was then added and again left for 1-2 hrs. The conditions for the two-stage leaches are shown in Table 12.

TABLE 12: Alkali and acid pressure leach conditions Test Feed Caustic Leach Conditions Acid Leach Conditions Max pH INaOH Time Pres Temp pH H2SO4 Time Pres Temp (9) (max) (g) (h) kPa (g) °C (min) (g) (h) kPa (g) °C 1 26, 5 12,4 165 2 2000 200 1,3 318, 3 2 2000 200 2 28, 9 12, 2 150 4 1000 155 1, 2 98, 6 1 1000 155 3 424, 3 11, 8 190 4 2000 200 2, 6 192,8 2 2000 200 4 551, 2 9, 84 500 4 1400 150 2,31 493,2 2 1400 150 5 641,1 12,1 200 4 2000 182 1,35 448,4 2 2000 182 6 616 9,87 240 4 1700 170 1,4 493,2 1 1700 170 The results are shown in Table 13. Mass reductions in excess of 95% were achieved with residue PGM grades over 3%.

TABLE 13. Alkali and acid pressure leach results Residue Deportment to Residue Test Ppm (%) % Pt Pd Au Cu Ni fe Pt Pd Au Cu Ni fe 1 (0,54) (1,09) 8,46 (3, 45) (8,96) (39,7) 100 100 4 1 4 119 2 (0,57) (0, 69) 5,15 (6,23) (9,92) (27,5) 100 100 0 3 3 45 3 842 963 3,51 (15, 1) (16,40) (9,52) 87 90 12 97, 5 76,8 100 4 1660 1890 7,57 (17,2) (18,30) (0,67) 100 100 28 90,5 62,1 77 5 (0,39) (2,09) 7,43 (33,3) (9, 37) 1,24) 75 77 7 38,8 8,2 100 6 (1,63) (1, 86) 25,8 (22,9) (8, 03)' (5,00) 100 100 9 8, 6 2,2 45 The solutions generated by the pressure leaches were progressively passed through 10m1 IRA 402 resin. The pH of the feed solutions varied between 0,49 and 5, 83.

The results are summarised in Table 14. It can be seen that selective extraction of PGMs was achieved over the base metals. Pt and Pd extractions were good at all pH values, but Au slippage seemed more pronounced at high pH.

TABLE 14: Ion Exchange results from a two-stage Alkaline/Acid Leach filtrate Volume bv Leach pH Au Pt Pd Cu Ni ml Cum Press Temp °C Feed mg/l (bar) Feed 1 20 200 0.54 0.03 0.66 0.08 641 1065 16 Raffinate 2450-4600 460 < 0. 01 < 0. 01 < 0. 01 634 1090 18 Feed 2 20 200 1. 42 0.09 0.20 0.03 796 1180 42 Raffinate 6110-7290 729 < 0. 01 < 0. 01 0.01 832 1200 43 Feed 3 10 150 0. 56 < 0. 01 < 0. 01 0. 02 747 978 < 1 Raffinate 9460-10860 1086 < 0. 01 < 0. 01 0.02 777 1020 <1 Feed 4 10 155 1.09 0.02 0.25 0.26 498 655 < 1 Raffinate 13020-13570 1357 < 0. 01 < 0. 01 0.05 501 700 < 1 Feed 5 10 155 1. 28 0.31 0.43 0.80 1965 2060 < 1 Raffinate 15420-18030 1803 < 0. 01 < 0. 01 0. 03 1720 1650 17 Feed 6 15 182 1. 62 0.06 0.81 1.41 1700 1600 16 Raffinate 18030-21310 2131 < 0. 01 < 0. 01 0.02 Feed 7 15 182 0. 53 002 0. 15 0.45 732 781 13 Raffinate 21310-24390 2439 < 0. 01 < 0. 01 0.02 850 847 13 Feed 8 20 200 0. 49 0.29 0.06 0.12 2500 1810 80 Volume bv Leach pH Au Pt Pd Cu Ni Co ml Cum Press Temp°C Feed mgn (bar) Raffinate 24390-28150 2815 < 0. 01 < 0. 01 < 0. 01 2460 1800 78 Feed 9 10 155 1. 06 0.32 1.24 2. 19 1700 3890 < 1 Raffinate 28150-30400 3040 0.04 < 0. 01 < 0. 01 1960 4350 3.3 Feed 10 10 155 0.94 0.29 0.26 2.07 1350 2030 2.7 Raffinate 30400-33800 3380 0. 03 < 0. 01 0.01 1790 2870 5.4 Feed 11 10 155 1. 52 0. 25 0.65 1.28 2660 3225 < 1 Raffinate 33800-35350 3535 0.03 < 0. 01 < 0. 01 2990 2690 < 1 Feed 12 20 200 0.66 0. 07 0. 09 0.06 1990 2340 83 Raffinate 35350-36650 3665 0. 03 0. 03 0.04 1990 2340 78 Feed 13 20 200 4. 68 0.07 0. 09 0. 06 151 1800 68 Raffinate 36650-39560 3956 0.07 < 0. 01 < 0. 01 30 2580 80 Feed 14 1. 02 0.11 0.65 1. 58 8190 10500 222 Raffinate 39560-40810 4081 0. 05 0.03 < 0. 01 10200 14100 322 Feed 15 20 150 3. 32 2.21 5. 98 4. 35 482 10100 15 Raffinate 40810-43660 4366 0.88 0.39 0.39 206 10120 <1 Feed 16 0. 72 1.17 1. 53 0. 51 7165 9935 19 Raffinate 43660-45210 4521 0. 28 0.02 0. 15 7180 9760 < 1 Feed 172. 74 2. 34 5. 49 1. 28 3288480 13 Raffinate 45210-48810 4881 3. 26 2. 64 0.79 794 5290 13 Feed 18 5.83 1.88 3.64 1.59 515 18400 34 Raffinate 48810-50660 5066 2.80 2. 32 1.06 335 18310 35 Feed 19'4. 14 1.29 2.18 0.82 6870 20000 432 Raffinate 50660-52310 5231 0.40 0.06 0.04 6570 16800 513 Feed 20 1.92 1.47 1.04 0.23 13700 21000 5 Raffinate 52310-53620 5362 0. 48 0. 10 0.07 11500 15300 430 Feed 213. 68 1. 74 0. 67 0.55 23700 28100 417 Raffinate 53620-56670 5667 0.40 0.08 0.09 18200 26650 463 Feed 22 0. 69 0.82 0.62 0.42 22500 23700 350 Raffinate 56670-58350 5835 3 19400 24900 417 Feed 23 0. 72 4. 02 6. 93 4.42 Raffinate 58350-62100 6210 3.02 1.74 3 Example 9: Recovery of PGM values from pH2 Acidification Residue by Calcining and Leaching In this series of tests, a pH2 acidification residue was calcined at varying temperatures and then leached at high temperature and pressure. The calcine temperature varied from 250°C to 800°C and the calcined material was leached using dilute sulphuric acid at 6 bar (600 kPa) and 140°C for 4 hours.

Figure 12 shows the effect of calcining temperature on the extraction of value metals. It can be seen that good separation is effected at a calcine temperature between 400-600°C where the PGM values remain in the residue and the base metal values essentially report to the solution. The results are summarised in Table 15.

TABLE 15: Extraction into solution during Leaching after Calcining Calcine Leach Extraction into solution % Temp Atm H2SO4 Pres Temp Time Pt Pd Au Cu Ni Fe Co °C g/l kPa (t) °C hr 250 Air 30. 48 600 140 4 62. 6 13.3 98. 2 1 94. 4 2.5 17.4 400 Air 30. 48 600 140 4 7. 1 2.7 8. 4 99. 5 99. 6 19. 8 98. 3 600 Air 30.48 600 14. 0 4 0.1 0.1 0.9 99.2 92.5 51.9 89.9 800 Air 30. 48 600 140 4 1. 1 0. 4 10. 1 97. 0 88. 5 45.1 75. 6 800 CO 30. 48 600 140 4 0. 1 0.3 1. 2 96. 1 99. 4 1. 8 98. 1 Example 10 : Recovery of PGM values from pH2 Acidification Residue by Acid leaching and Fusion In this series of tests, a pH2 acidification residue was leached with sulphuric acid under varying conditions of temperature and pressure. The residue from the leaching step was mixed with concentrated sulphuric acid and heated to 300°C for 60 minutes (fusion step).

The pH2 acidification residue produced from the pilot plant testwork in Example 5 was weighed (800gm wet, 380-41 Ogm dry) into an autoclave and slurried with water.

A known weight of concentrated sulphuric acid was added. The vessel was heated with steam to the target temperature and pressurised with oxygen. The reaction time was 4 hours.

The residue from the pressure acid leach was then subjected to a concentrated sulphuric acid destruct (fusion) by blending with a measured volume of concentrated acid and heated with steam to 300°C for 1 hour. The leach slurry was diluted with water and filtered.

The conditions for the pressure acid leach and fusion are shown in Table 16.

TABLE 16: Leach and Fusion Tests Conditions. Test Leach (3 hours) Fusion (1 hour) Overall Temp Pres H204 Mass Temp Mass Mass ° C kPa (g) Ratio Reduction °C Reduction Reduction 1 180 1500 1, 10 91, 8 300 95,5 99,7 2 180 1500 1,68 88, 3 300 98,5 99,8 3 180 1200 1,50 88, 8 300 95,9 99,6 4 180 1500 1,08 94, 2 300 94, 4 99, 8 The results are summarised in Tables 16 and 17. It can be seen that greater than 99% mass reductions were achieved with final concentrate analyses as follows : Pt 6-25% Pd 8-20% Au 0,01-0, 04% TABLE 17: Analyses of products produced during Leaching and Fusion Test Stream Pt (ppm) Pd (ppm) Au (ppm) 1 Leach feed 1100 1200 9.77 Leach residue/Fusion feed 11500 12400 15.6 Fusion residue 250000 199000 371 2 Leach feed 1100 1200 10. 7 Leach residue/Fusion feed 12300 12900 3.63 Fusion residue Insufficient sample 3 Leach feed 1100 1090 9. 66 Leach residue/Fusion feed 10600 11800 5.27 Fusion residue 61100 85200 109 4 Leach feed 1100 1090 9. 66 Leach residue/Fusion feed 20100 24500 11 Fusion residue 121000 173000 165 Example 11 : Upgrade Process for Acidification Residue This test, run as a pilot campaign, illustrates the upgrade process for the precipitate produced from the pH2 acidification of cyanide solution.

The precipitate as produced in Example 6 was repulped in sulphuric acid, water and a lixiviant sourced from the second stage leach. The slurry was fed to the first stage leach, The flowsheet employed in this pilot campaign is shown in Figure 13.

The first stage leach was an oxidative pressure leach with counter current oxygen flow. The vent gas was scrubbed to recover hydrogen cyanide that was not destroyed in the autoclave. The operating conditions for this stage leach were as follows: Total pressure 1500kPa (g) Oxygen partial pressure 450-500 kPa (g) Temperature 180°C Slurry pH 0.7-0. 8 Retention time 2.5 h Feed rate 0.25 kg/h Number of compartments 1 The mass reduction was only 80% (possibly due to only one compartment). The first stage residue was blended with sulphuric acid, leached at approximately 300°C in mild steel vessels and then quenched in water at 90°C. The conditions for the second stage of leaching were as follows : Pressure Atmospheric Temperature 290 - 310°C Retention time 1 h Water quench temperature 90°C Water quench retention time 1 h Sulphuric acid/residue 3. 6 The residue from the second stage leach was the final concentrate with a typical analysis at steady state as shown in Table 18 TABLE 18 : Final Concentrate Analysis % NI Cu 0.4 Fe 0. 5 Co 0. 5 Pt 36. 7 Pd 43. 8 Au 0. 1 Rh 0. 04 Ru 0. 01 0. 02 S 0. 4 Se 1. 0 Te 0. 02 Si02 1. 8 The PGM split between first stage leachate and the final concentrate is shown in Table 19.

TABLE 19: Recovery of PGMs to outlet streams To Concentrate To Leachate % % Pt 99.64 0. 36 Pd 99. 63 0. 37 Au 59. 1 40. 9 The final concentrate contained well in excess of 99% of the PGMs in the leach feed at a grade of greater than 80% total PGMs. Less than 0. 1% of the base metals in the leach feed reported to the final concentrate.

The first stage leachate was conditioned with sodium chloride at approximately 15g/) and then passed through pilot ion exchange columns filled with IRA-402 in the first, second and fourth columns and activated carbon in the third column. The raffinate values were typically 10 ppb for each element and the load distribution is given in Table 20.

TABLE 20: Ion Exchange columns loading distribution Metal Distribution % Pt Pd Au Column 1 97. 4 98. 5 24. 3 Column 2-0. 3 0. 07-2. 2 Column 3 0. 9-1. 6 Column 4 0.7 - 5.8 Only minor quantities of base metals were loaded on the resin whilst an overwhelming majority was taken to a carbonate precipitation with a composition shown in Table 21.

TABLE 21: Base Metal residue analysis % Cu 12-18 Ni 22-30 Fe 05. - 1. 5 Co 0.15-0. 20 Pt <0. 0010 Pd <0. 0005 Au <0.00001 Rh <0. 0001 Ru<0. 0001 Ir <0. 0005 As <0. 0005 Na ~2. 0