AU5401801A - Processing gold containing copper sulphide feeds - Google Patents

Description

P/00/011 Regulation 3.2

AUSTRALIA

Patents Act 1990 ORI GINAL COMPLETE SPECIFICATI ON STANDARD PATENT Invention title: Processing gold containing copper sulphide feeds The following statement is a full description of this invention, including the best method of performing it known to us:

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S. S e S S 55 agbm 10742791 Processing Gold Containing Copper Sulphide Feeds Field of the invention This invention relates to processing gold containing copper sulphide feeds to recover gold and copper. In a particular aspect, the invention relates to processing of copper sulphide concentrates containing chalcopyrite to recover gold and copper.

Background of the invention Dissolution of the gold content of gold-containing copper sulphide concentrates by direct cyanidation is generally quite unsatisfactory with gold dissolution often being less than 10%, while the actual cyanide consumption is uneconomically high. One reason for the high cyanide consumption is that the copper sulphides show variable but often significant solubility in alkaline cyanide solutions. Moreover, there are problems associated with the recovery of metallic copper from the pregnant copper cyanide-containing leach liquor.

Although many hydrometallurgical processes have been proposed for treating such concentrates, the commercial application of these processes is very limited. Most of the proposed processes involve two separate leaching stages. In the first stage, the copper is leached, usually in an acidic sulphate reaction system. The gold is subsequently leached from the first stage leach residue using the conventional alkaline cyanide leaching system.

Because of the relatively refractory nature of many gold-containing copper sulphide

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20 concentrates, the necessary processing conditions, particularly the initial leaching step, are ::either quite aggressive or involve extended reaction times, in order to dissolve the copper sulphide minerals prior to the separate dissolution of the gold content in the leach residue.

These conditions are required to ensure that the gold, which is often present as submicroscopic inclusions within the sulphide mineral grains, is sufficiently exposed to 'o 25 facilitate the gold-cyanide leach reactions to occur. The aggressiveness of the copper dissolution reaction system is dictated by a number of factors, especially the type of copper sulphide mineralisation. Chalcopyrite, CuFeS 2 is considerably more refractory than chalcocite, Cu 2 S. That is, more aggressive conditions are required to dissolve chalcopyrite Scompared with those for chalcocite. In addition to the copper sulphide minerals, the feed concentrate will typically contain pyrite, FeS 2 and/or arsenopyrite, FeAsS. A variable portion of the total gold content of the feed is quite often associated with these minerals..

portion of the total gold content of the feed is quite often associated with these minerals.

agbm 10742791 Thus the decomposition of these minerals is generally required to achieve an economically sustainable overall gold recovery.

Included in the proposed processes for treating gold-containing copper sulphide concentrates are bacterial leaching using both the direct and indirect bacterial oxidation routes that involve direct bacterial reaction of the sulphide minerals and bacterially assisted regeneration of ferric sulphate respectively, and acidic pressure oxidation. In each of these prior proposed processes, "complete" recovery of the copper and gold values in the feed concentrate can be achieved, especially under laboratory or small-scale pilot plant conditions. However, the estimated capital and operating costs associated with their operation at the commercial scale, combined with the technical difficulties associated with such operations, are such that they are yet to be adopted.

In the bacterial leaching processes, of which there are several versions, quite long reaction times, typically 5-6 days, are required to achieve economically acceptable copper dissolution levels. If the oxidation of the sulphide sulphur component of the feed is not complete, then any elemental sulphur in the leach residue will interfere with the cyanidation of the gold in the carefully washed and neutralised residue once it has been separated from the acidic copper-containing leach solution. During the bacterial oxidation/leaching stage there is significant dissolution of iron and generation of free acid.

These must be precipitated and partly neutralised before the copper can be recovered, 20 typically by a combination of solvent extraction and electrowinning.

o:ooo The capital and operating costs associated with this iron precipitation and partial neutralisation are substantial and tend to make this proposed route uneconomic.

oooo Acidic pressure oxidation can be carried out under a range of conditions, especially temperature. Dissolution of the sulphide minerals occurs via a number of concurrent reactions, including direct oxidation via dissolved oxygen and by reaction with ferric sulphate. The required ferric sulphate concentration is provided in part by several side reactions that occur during the sulphide mineral leach process, as well as that contained in ;.the recycled process water that is returned to the leach circuit from the copper recovery circuit.

30 Because of the refractory nature of chalcopyrite, the most common copper sulphide mineral, reaction conditions generally need to be aggressive to minimise reaction times and maximise overall copper dissolution. One way of making the reaction conditions as mild agbm 10742791 as possible is to fine grind the concentrate ahead of the leach step. This forms the basis of the so-called Activox process. Here the feed concentrate, which will typically have a grind size of minus 45 micron, is reground to a P80 of 6-10 micron. This requires a substantial energy input. The product can then be leached at about 100 0 C in an acidic s sulphate solution under a relatively small oxygen over-pressure. Under these conditions, oxidation of sulphide sulphur to sulphate is limited, most reporting as elemental sulphur.

As noted previously, the presence of the elemental sulphur in the leach residue hinders the subsequent gold cyanidation step.

By carrying the process out at a somewhat higher temperature, say 130-150C, fine grinding is not required, although once again, the leach residue will generally have a significant elemental sulphur content. Various methods for removing this sulphur from the leach residue have been proposed. These include flotation and dissolution in an appropriate organic solvent. While these processes are relatively straightforward at the laboratory scale, operation at the commercial scale is both difficult and costly. As such, they are not widely practiced.

In both of the above acidic sulphate pressure leaching options, and depending upon the relative proportions of chalcopyrite, chalcocite and pyrite in the feed, there is the potential for insufficient ferric sulphate to be regenerated or contained in the recycled process water in order to sustain the leach process on a continuous basis. Under these conditions, it is S 20 necessary to supplement the leach solution with the required amount of imported ferric sulphate. This will substantially increase the operating costs.

Complete oxidation of the sulphide minerals can be achieved by operating the acid .99.

sulphate leach circuit at temperatures above about 220 0 C. Although gold cyanidation of the washed leach residue is relatively straightforward, the process suffers from the high 25 costs associated with the neutralisation of the free sulphuric acid generated during the 9 leach stage. In addition, because of the complete dissolution of the iron-containing sulphide minerals, an excess of ferric sulphate may be generated. Most of this acid and o:oo °°excess ferric sulphate must be neutralised or precipitated before the pregnant leach solution *999 is clarified and forwarded to the solvent extraction-electrowinning circuit for copper metal 30 recovery. Apart from the cost associated with the high limestone consumption, gypsum 9 9(CaSO 4 .2H 2 0) is precipitated. This has a tendency to form as a scale on reactor and pipe walls, agitator blades and shafts, etc,. and even in the solvent extraction mixer-settler agbm 10742791 4 boxes. This can result in equipment failure, blockages, etc. It may be necessary to incorporate substantial downtime in the production schedule to allow for regular gypsum descaling, etc.

Other proposed methods of processing gold-containing copper sulphide concentrates include leaching with ammoniacal ammonium sulphate, acidic ferric/cupric chloride, and nitric acid. As with the previously described processes, these all suffer from a range of technical difficulties and excessive capital and operating costs.

In order to overcome the processing difficulties and high costs associated with the currently proposed treatment options for gold-containing copper sulphide concentrates, a critical analysis of these options was undertaken. From this evaluation, it was concluded that the ideal leaching process would be one that prevented excessive formation of elemental sulphur to simplify gold cyanidation, kept lime and cyanide consumption during gold recovery within acceptable bounds, facilitated integration with a downstream copper recovery circuit, and regenerated acidic ferric sulphate for recycling in the process on a continuous basis.

It is an object of the present invention to provide a process which overcomes or ameliorates one or more of the aforesaid processing difficulties.

Disclosure of the invention S. In one aspect the invention provides a method for recovering gold and copper from a feed 20 which includes these metals comprising the steps of: carrying out a primary leach of the feed in the presence of oxygen, acid and ferric ions at a temperature between 70'C and 110 C to produce a primary leach mix containing a primary leach solution and primary residual solids; ooooo3 (ii) directing the primary leach mix to a separation stage for separating the •or primary leach solution from the primary residual solids; (iii) subjecting a major proportion or all of the primary residual solids to a gold recovery process; S. (iv) subjecting a minor proportion of the primary residual solids to a secondary leach reaction in the presence of oxygen at a temperature in the range 160'C agbm 10742791 to 260 0 C to produce a secondary leach mix containing secondary leach liquid and secondary residual solids; recycling the secondary leach mix to step and (vi) extracting copper from the leach solution separated in step (ii) to produce a S spent solution.

Typically, the feed includes mineral species chosen from the group including chalcopyrite, chalcocite, bornite and covellite and mixtures thereof in association with gold.

The spent solution may be recycled to the primary leach or to another suitable area of the overall flowsheet such as, for example, preacidification of the feed to remove carbonate mineralisation. Suitably the feed is in the form of a concentrate. It may be a floatation concentrate. It may contain minerals such as pyrite, copper sulphides, and/or arsenopyrite.

To improve the rate of leaching and assist with recovery of copper and gold, it is preferred that the feed be ground prior to leaching. Suitably it will be ground as part of the separate flotation process such that the major weight proportion of the feed is composed of particles having a particle size in the range 20 to 200 microns. More preferably the particle size range is 40 to 100 microns.

The primary leach may be carried out under conditions which ensure that a major proportion of copper containing minerals other than chalcopyrite are dissolved in the leach reaction. Suitably, at least 80% and more preferably at least 90% of copper containing 20 minerals other than chalcopyrite may be dissolved.

.o It is to be understood that the operating parameters of the primary leach will be dependent oooo upon the nature of the feed being leached. However, in typical situations, it is anticipated that the acid concentration will be maintained in the pH range 0.5 to 1.5 Similarly, the .:oooi concentration of ferric ions may be maintained in the range 10 to 50 g/L.

a The oxygen taken up in the primary leach reaction may simply be oxygen at atmospheric pressure. However, it is preferred that the oxygen partial pressure be higher than that available through normal atmospheric pressure. An oxygen partial pressure of 4 bar or more or more preferably 6 bar or more and up to 15 bar may be used to speed up the reaction.

o• agbm 10742791 As the primary leach reaction results in dissolution of a variable amount of the non-chalcopyrite containing minerals, the reduction of solids in the feed may be substantial. For example, there may be a weight reduction of 10% to 70% as a result of the dissolution of minerals. About 30-50% would be typical of many feeds.

As a result of the dissolution referred to above, the proportion of chalcopyrite material contained in the solids will substantially increase. Suitably the increase is at least 50% as a proportion of the total copper mineralisation in the remaining solids.

To achieve the desired degree of dissolution for the primary leach, it is anticipated that leach times of the order of 1 hours to 2 hours will be necessary.

Furthermore, the preferred leach temperature range for the majority of copper containing (non-chalcopyrite) minerals is 900C to 110 C.

Should the feed contain significant quantities of pyrite, it is desirable to ensure that the conditions are controlled so that excess ferric ions are not produced as a result of the pyrite being leached excessively.

A proportion of the primary residual solids may be recycled to the primary leach. This may occur by mixing the primary residual solids with the solid feed prior to commencing the main leach. Alternatively the primary residual solids may be simply added to the reactor in which the main leach is carried out.

In another approach the minor proportion of primary residual solids recycled to the main oo: 20 leach may be recycled after all of the primary residual solids have been subjected to a gold recovery process.

Regardless of which method of recycling is used, it is anticipated that 5 to 45 per cent by weight of the primary residual solids will be recycled to the primary leach. More •oo.oi preferably 10 to 20 per cent will be recycled.

The primary residual solids may be subjected to filtration and washing before carrying out a gold recovery process on them. The gold recovery process may involve cyanidation of the primary residual solids followed by recovery of dissolved gold according to the wellknown carbon-in-leach process. Alternatively, gold recovery may be achieved by leaching with alternative lixiviants, for example, under acidic conditions.

agbm 10742791 Copper may be recovered from the primary leach solution using a solvent extraction process. The copper recovered in this fashion will be in solution.

To extract copper in its solid metallic form, an electrowinning process may be employed.

As this will result in the production of excess acid, a proportion of this excess acid may be bled from the electrowinning circuit directed to the primary leach.

The solution containing ferric iron and acid bled from the solvent extraction step as raffinate may be combined with the spent electrolyte prior to recycling to the main leach.

In order to reduce build up of undesirable contaminants, a purifying bleed may be taken from this combined recycled solution. The purifying bleed may itself be subject to further solvent extraction to remove residual amounts of copper.

Where the feed is a flotation concentrate, and where appropriate it is anticipated that the tailings of the flotation process would contain mineralisation capable of neutralising acid.

Accordingly, the tailings may be combined with the purifying bleed to neutralise the bleed before disposing of it. This neutralisation may occur after the further solvent extraction used to remove most of the copper. Alternatively, other suitable neutralizing agents such as limestone, lime, soda ash, or caustic soda, may be used.

As stated above, secondary leach reaction is carried out in the presence of oxygen at a temperature of from 160 0 C to 260 0 C. More typically, the temperature is between 120 0

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and 160 0 C. It is to be understood that the operating parameters of the secondary leach will S 20 be dependent upon the nature of the feed being leached. However, in typical situations, it S is anticipated that the oxygen taken up in the primary leach reaction may simply be oxygen i at atmospheric pressure. However, it is preferred that the oxygen partial pressure be higher than that available through normal atmospheric pressure. An oxygen partial pressure of 4 bar or more or more preferably 6 bar or more and up to 15 bar may be used to speed up .oo..

25 the reaction.

A preferred embodiment of the invention will now be described with reference to the accompanying drawing: Brief Description of the Drawing i °Figure 1 shows a flow chart of a process for carrying out the method of the invention.

Now referring to the drawing, Figure 1 shows a copper and gold recovery circuit generally designated 1.

agbm 10742791 A feed in the form of a flotation concentrate 3 which has been ground to a suitable size for leaching, say 38 microns, is sent to the main leach tank The main leach tank is a pressure vessel which can be operated at temperatures of around to 105 0 C. A source of oxygen 7 at super atmospheric pressure is directed to maintain a partial pressure of oxygen of about 6 bar in the main leach tank. Recycled liquids 8 are also directed to the main leach tank.

The output of the main leach tank is a primary leach mix 9 which is directed to a thickener 11. The thickener separates the primary leach mix into liquid 13 and solid The solids 15 are in turn split into recycled solids 17 and a major solids proportion 27.

The recycled solids typically about 15 per cent of the total solids remaining after the leach reaction are directed to an autoclave 19 maintained at a temperature of about 220'C.

The autoclave has additional inputs in the form of oxygen 7 which maintains the oxygen partial pressure in the autoclave at about 6 bar. Furthermore quench water 23 may be added.

The secondary leach mix produced in the autoclave is then directed to the thickener 11.

The major solids portion is filtered in the filter 29 and washed with washing water 31.

Filtered solids 33 are then treated in a carbon-in-leach facility 35 where gold is extracted 4. using the conventional carbon-in-leach process involving cyanidation and gold recovery from solution using activated carbon.

The liquid filtrate 38 from the filter is then directed to the clarifier 39 where it is combined with the liquid 13 from the thickener to separate the mixture into solids 41 and clarified liquid 43.

Solids 41 are directed to the filter 29 whilst the clarified liquid 43 is directed to a solvent extraction facility 25 At the solvent extraction facility, a suitable extractant in a suitable diluent such as kerosene is used to extract copper in solution from the clarified liquid 43.

A highly acid liquid (49) is in turn used to strip copper from the loaded extraction liquid to provide an acid solution 47 loaded with copper. This is directed to the electrowinning facility 48 for retrieving solid copper by an electro-chemical process. The electrowinning process as a by-product produces acid.

agbm 10742791 The acidic spent electrolyte 49 from the electrowinning is recycled to the solvent extraction process and an acidic spent electrolyte bleed 51 is combined with the spent raffinate 46 to be recycled as recycle liquid 8 to the main leach In order to prevent build up of undesirable contaminants, a bleed 53 is taken from the recycled liquid 8. After removal of a proportion of residual copper using a scavenger solvent extraction process 55, the depleted solution 56 is directed to a bleed neutralisation facility 57. There it may be neutralised with tails from the float process used to produce the feed 3 before being sent to a neutralised solids disposal facility 61.

The flow diagram of Figure 1 also includes a number of alternative approaches falling with the scope of the invention which may be employed, depending on the nature of the feed fed into the process and other factors such as the economics of the process.

Thus for example, a proportion of the feed 3 may be directed as a bleed feed 21 into the autoclave 19 without going through the main leach. In another approach a separate source of chalcopyrite -rich feed (3a) may be directed into the autoclave (19) without going through the main leach.

In another arrangement, a proportion or all of the recycled solids sent to the autoclave may be recycled as recycled solids 37 following gold removal in the carbon-in-leach facility Another variation on the process could involve recycling part or all of the liquid/solids mix from the autoclave 19 directly into the main leach tank S: o 20 In each of the alternatives suggested, the optimum process will depend upon a variety of parameters such as the nature of the feed, the costs of performing each stage including the **cost of the inputs to the various stages and the overall balance of the economics of the various steps of the process being carried out.

Examples The invention will now be further described with reference to the following non-limiting examples. The person skilled in the relevant art will readily appreciate that the examples are indicative only so that actual conditions will depend upon, amongst other things, the mineralogical and chemical composition of the feed and the operating parameters of the S"entire copper-gold processing circuit. The person skilled in the relevant art will also 3o appreciated that the amount of residue from the main, or low temperature, leach step that is subsequently subjected to a high temperature leach step will largely be determined by agbm 10742791 various criteria such as the deportment of the gold, the relative amounts of copper sulphide minerals and pyrite in the feed.

In these examples, reference is also made to the flowsheet depicted in Figure 1 describing an embodiment of the present invention, and in particular to the two separate leach steps and the splitting of part of the residue from the main (low) temperature leach step and directing it as feed to the secondary (high) temperature leach step.

Example 1 In this example, a copper-gold concentrate containing 11.6% chalcopyrite, 20.2% chalcocite and 47.4% pyrite and assaying 20.2% copper and 80 g/t gold was subjected to a high temperature leach test at 210 0 C and an oxygen overpressure of 680 kPa for one hour using a simulated recycle leachant containing 30 g/L acid and 24 g/L iron. Analysis of the resultant leach liquor showed that more than 98% of the copper has been leached as soluble copper sulphate. Following separation of the leach residue by filtration, the carefully washed residue was subjected to a standard gold cyanidation test. More than 94% of the gold was dissolved in this step. The pH of the clarified copper-rich pregnant leach liquor was subsequently increased by addition of lime to a value consistent with forwarding to the copper solvent extraction circuit for recovery of copper. The amount of alkali required for both gold cyanidation and neutralisation ahead of solvent extraction was very substantial. This is a reflection of the amount of free acid generated in the high 20 temperature leach test. The amount of neutralisation required is regarded as being commercially uneconomic -Example 2 In this example, a bulk copper-gold concentrate containing 7.2% chalcopyrite, 3.2% chalcocite and 72.4% pyrite and assaying 6.98% copper and 44.3 g/t gold was subjected to 9 a high temperature leach under the same conditions as outlined in Example 1. In this example, high recoveries of both copper and gold were achieved. However, the amount of acid generated during leach was considerably greater, a reflection of the higher pyrite content of this feed. As a consequence, the amount of alkali required for both gold "i cyanidation and copper solvent extraction was also substantially greater, and once again S. 30 regarded as commercially uneconomic.

agbm 10742791 Example 3 In this example, the bulk copper-concentrate used in Example 1 was leached at 100 0 C and an oxygen overpressure of 1000 kPa. Analysis of the solid and liquid components of the resultant pregnant leach pulp indicated that more than 95% of the chalcocite had been leached, together with about 25% of the chalcopyrite and 7.5% of the pyrite. This indicated that approximately 80% of the total copper content had been leached. The amount of acid generated was very significantly reduced, leading to a much lower alkali requirement for neutralisation purposes.

The residue from above "low temperature" leach test was now subjected to a high temperature leach test under the conditions described in Example 1. It was found that all of the remaining copper in the residue could be leached from this residue. All of the gold in the copper free residue could be recovered by cyanidation although the amount of acid generated during the high temperature leach step was found to be excessive.

Example 4 In this example about 20% of the residue resulting from the low temperature leach outlined in Example 3 was subjected to the second high temperature leach test. The amount of acid and soluble iron generated in the pregnant leach liquor in this high temperature test was found to be appropriate for return to the main, or low temperature, leach step. In addition, it was found that the gold in the final residue was readily recovered by cyanidation with an 20 acceptable lime consumption required for neutralisation.

Example *This test involved integration of the main (low) and high temperature leach steps on a locked-cycle basis using a 20% split factor. The leach residue from the high temperature leach step was combined with that from the main (low) temperature leach step for gold o* recovery by cyanidation. Overall copper and gold extractions were found to be in excess of 85% and 95% respectively with economically acceptable lime and cyanide consumptions.

The word 'comprising' and forms of the word 'comprising' as used in this description does not limit the invention claimed to exclude any variants or additions.

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agbm 10742791 12 Modifications and improvements to the invention will be readily apparent to those skilled in the art. Such modifications and improvements are intended to be within the scope of this invention.

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agbm 10742791 INTEGER LIST 1 3.1 3a 5.

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33.1 35.( 37.1 38.I 39.( 41.

iu/Au gold recovery circuit 'lotation concentrate feed kdtemative feed vlain (primary) leach tank )xygen Zecycle liquids ~rimary leach mix Fhickener .iquid olids ecycle solids kutoclave Meed feed uench Water econdary leach mix \.lternative secondary leach mix vlajor solids proportion ilter Mashing water ~iltered solids 'arbon-in-leach facility ecycled solids iquid filtrate ilarifier olids 9*

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5* 5. 9 5* 9*, agbm 10742791 14 43. Liquid Solvent extraction facility 46. Spent raffinate 47. Copper loaded acid solution 48. Electrowin facility 49. Spent electrolyte 51. Spent electrolyte bleed 53. Bleed solution Scavenger solvent extraction 56. Depleted bleed solution 57. Bleed neutralisation 59. Float tails 61. Neutralised solids disposal facility

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Claims (39)

1. A method for recovering gold and copper from a feed which includes these metals comprising the steps of: carrying out a primary leach of the feed in the presence of oxygen, acid and ferric ions at a temperature between 70'C and 110 C to produce a primary leach mix containing a primary leach solution and primary residual solids; (ii) directing the primary leach mix to a separation stage for separating the primary leach solution from the primary residual solids; (iii) subjecting a major proportion or all of the primary residual solids to a gold recovery process; (iv) subjecting a minor proportion of the primary residual solids to a secondary leach reaction in the presence of oxygen at a temperature in the range 160'C to 260 0 C to produce a secondary leach mix containing secondary leach liquid and secondary residual solids; recycling the secondary leach mix to step and (vi) extracting copper from the leach solution separated in step (ii) to produce a spent solution.

2. A method for recovering gold and copper from a feed including a mixture of chalcopyrite, chalcocite, bomite and covellite associated with the gold, the method comprising the steps of: carrying out a primary leach of the feed in the presence of oxygen, acid and ferric ions at a temperature between 70'C and 110°C to produce a primary "•leach mix containing a primary leach solution and primary residual solids; (ii) directing the primary leach mix to a separation stage for separating the 25 primary leach solution from the primary residual solids; (iii) subjecting a major proportion or all of the primary residual solids to a gold S. recovery process; (iv) subjecting a minor proportion of the primary residual solids to a secondary leach reaction in the presence of oxygen at a temperature in the range 160'C agbm 10742791 to 260 0 C to produce a secondary leach mix containing secondary leach liquid and secondary residual solids; recycling the secondary leach mix to step and (vi) extracting copper from the leach solution separated in step (ii) to produce a spent solution.

3. A method according to either claim 1 or claim 2 wherein the feed is in the form of a concentrate.

4. A method according to any of the preceding claims wherein the feed comprises minerals chosen from the group including pyrite, copper sulphides, arsenopyrite and mixtures thereof.

A method according to any of the preceding claims wherein the feed is ground prior to leaching.

6. A method according to claim 5 wherein the feed is ground as part of a separate flotation process and the major weight proportion of the feed is composed of particles having a particle size in the range 20 to 200 microns.

7. A method according to claim 6 wherein the particle size range is 40 to 100 microns.

8. A method according to any of the preceding claims wherein the spent solution is recycled.

S9. A method according to claim 8 wherein the spent solution is recycled to the primary leach.

10. A method according to any of the preceding claims wherein the primary leach is carried out under conditions which ensure that a major proportion of copper containing •minerals other than chalcopyrite are dissolved in the leach reaction.

11. A method according to claim 10 wherein at least 80% of copper containing 25 minerals other than chalcopyrite are dissolved.

12. A method according to any of the preceding claims wherein the acid concentration of the primary leach is maintained in the pH range 0.5 to

13. A method according to any of the preceding claims wherein the concentration of ferric ions in the primary leach is maintained in the range 10. to 50 g/L. agbm 10742791

14. A method according to any of the preceding claims wherein the oxygen partial pressure is higher than normal atmospheric pressure.

A method according to claim 14 wherein the oxygen partial pressure in the primary leach is between 4 bar and 15 bar.

16. A method according to any of the preceding claims wherein the primary leach time is between 1 hour and 2 hours.

17. A method according to any of the preceding claims wherein the primary leach reaction reduces the weight of solids in the feed by 10% to

18. A method according to claim 17 wherein the reduction in the weight of solids is between 30% and

19. A method according to any one of claims 1 to 16 wherein the feed includes chalcopyrite and the primary leach reaction increases the proportion of chalcopyrite material contained in the solids by at least 50% as a proportion of the total copper mineralisation in the solids.

20. A method according to any of the preceding claims wherein the primary leach temperature is 900C to 1

21. A method according to any of the preceding claims wherein the primary residual solids are recycled to the primary leach.

22. A method according to claim 21 wherein the primary residual solids are mixed with 20 the solid feed prior to commencing the main leach.

23. A method according to claim 21 wherein the primary residual solids are added to the reactor in which the main leach is being carried out. 0

24. A method according to any of the preceding claims wherein the minor proportion of primary residual solids are recycled to the main leach after all of the primary residual 25 solids have been subjected to a gold recovery process.

25. A method according to any of the preceding claims wherein 5 to 45 per cent by weight of the primary residual solids are recycled to the primary leach.

26. A method according to claim 25 wherein 10 to 20 per cent by weight of the primary residual solids are recycled to the primary leach. agbm 10742791

27. A method according to any of the preceding claims wherein the primary residual solids are filtered and washed before being subjected to a gold recovery process.

28. A method according to claim 27 wherein the gold recovery process includes the steps of cyanidation of the primary residual solids followed by (ii) recovery of dissolved gold according to a carbon-in-leach process.

29. A method according to claim 27 wherein the gold recovery process includes the step of leaching under acidic conditions.

A method according to any of the preceding claims wherein copper is recovered from the primary leach solution by a solvent extraction process.

31. A method according to any of claims 1 to 29 wherein copper is recovered from the primary leach solution by an electrowinning process and at least part of any excess acid produced by the electrowinning process is directed to the primary leach.

32. A method according to claim 30 wherein a solution containing ferric iron and acid bled from the solvent extraction step is combined with spent electrolyte prior to recycling to the main leach solution.

33. A method according to claim 32 wherein a purifying bleed is taken from the combination of spent electrolyte and ferric iron and acid solution.

S34. A method according to claim 33 wherein the purifying bleed is subject to further solvent extraction to remove residual amounts of copper. 20

35. A method according to claim 33 wherein tailings are combined with the purifying bleed to neutralise the bleed before its disposal.

36. A method according to claim 33 wherein neutralizing agents are added to the purifying bleed to neuturalise the bleed before its disposal.

37. A method according to any one of claims 1 to 36 wherein the secondary leach 25 reaction is carried out in the presence of oxygen at a temperature of from 220 0 C and 260 0 C.

38. A method according to any one of claims 1 to 37 wherein the overpressure of oxygen in the secondary leach is between 600 and 1200kPa.

39. A method according to claim 38 wherein the overpressure of oxygen in the secondary leach is between 900 and 11 00kPa. agbm 10742791 19 A method according to claim 1 and substantially as herein described with respect to the drawing NEWCREST MINING LIMITED 22 June 2001 S.. S S S S S 5555 S U.S. S S 5.G S S S S. S S S S SS agbm 10742791