CA2648951A1

CA2648951A1 - Process for recovery of antimony and metal values from antimony- and metal value-bearing materials

Info

Publication number: CA2648951A1
Application number: CA 2648951
Authority: CA
Inventors: Dale Harrison
Original assignee: Straits Resources Limited; Dale Harrison
Current assignee: STRAITS RESOURCES Ltd
Priority date: 2006-04-11
Filing date: 2007-04-11
Publication date: 2007-10-18
Also published as: WO2007115377A1; MX2008013141A; ZA200808628B; AU2007236559A1; CN101443466A; EA200870420A1

Abstract

A process for recovery of antimony and metal values from antimony- and metal value-bearing materials is provided. A coarse (typically >250-µm) concentrate of the antimony- and metal value-bearing materials is provided. The coarse concentrate is then subjected to a process for selective recovery of antimony in the form of solid antimony or an antimony containing solution. A residue containing the remaining metal values is then subjected to suitable recovery processes to recover the remaining metal values.

Description

PROCESS FOR RECOVERY OF ANTIMONY AND METAL VALUES FROM
ANTIMONY- AND METAL VALUE-BEARING MATERIALS
Field of the Invention The present invention relates to a process for recovery of antimony and metal values from antimony- and metal value-bearing materials, in particular to a process for recovery of antimony and gold values from materials containing stibnite and gold-bearing materials, such as quartz, pyrite, and arsenopyrite.

Additionally, the present invention relates to a process for recovery of a metal value from a mixture of metal value-bearing material and gangue material, wherein the hardness of the metal value-bearing material is less than the gangue material, the difference in hardness being such that the metal value-bearing material is preferentially ground relative to the gangue.

Background of the Invention In conventional hydrometallurgical metal value recovery, a high grade concentrate is frequently obtained by a process of crushing, grinding and froth flotation before it undergoes further metal recovery through leaching processes, electrowinning, and the like. Typically, in the production of the high grade concentrate there are losses of the metal value, depending on the mineralogy of the concentrate, through grinding, clean-up and upgrading processes. These losses may be greater when one of the metal value bearing materials is softer, for example less than 2 Mohr units, than the other metal value bearing materials or the gangue because the softer material may be ground to a particle size which is too small to be recovered by conventional apparatus or equipment.

Stibnite is an antimony bearing material and it is considered to be a soft mineral. Stibnite can be associated with gold and silver bearing materials and the materials associated therewith, particularly quartz and pyrites which are considered to be hard minerals. Many other metal values are sourced from sulphide ores, such as pyrites and chalcopyrites. Frequently, these sulphide ores are also associated with precious metals, such as gold and silver.
Any process for recovery of mixed metal values from metal sulphide bearing materials, particularly if they also contain precious metals, requires consideration of a first process which will selectively recover the base metal of interest under conditions which do not compromise a second process for selective recovery of the precious metal, and vice versa.

The present invention seeks to overcome at least some of the aforementioned disadvantages.

Summary of the Invention According to one aspect of the invention there is provided a process for recovery of antimony and metal values from antimony- and metal value-bearing materials, the process comprising the steps of:
a) providing a coarse concentrate of the antimony- and metal value-bearing materials;
b) subjecting the coarse concentrate to a process for selective recovery of antimony, thereby producing solid antimony or an antimony-containing solution and a residue containing the remaining metal values; and, c) recovering the remaining metal values from the residue.

In one embodiment of the invention, the antimony bearing material is stibnite, auro-stibnite, tetrahedrite, or a combination thereof. Suitable examples of said antimony-bearing minerals include, but are not limited to, an ore, a concentrate or any other material containing an antimony-bearing mineral.

The metal value-bearing material is an ore, a concentrate, or any other material from which metal values may be recovered. In one embodiment, the metal value-bearing material is a metal sulphide containing one or more metal values. Preferably, the metal value-bearing material is a metal sulphide associated with precious metal value-bearing materials, such as gold and/or silver. Typical examples of such metal sulphides associated with gold and/or silver are pyrites, chalcopyrites, and arsenopyrites.

Generally, the coarse concentrate is a coarse flotation concentrate. The coarse feed stream concentrate may be prepared by crushing and grinding the antimony- and metal value-bearing materials, and subjecting the resulting materials to a flotation process. In one embodiment the flotation process comprises conducting a first flotation process arranged to capture coarse particles of antimony-and metal value-bearing materials, and optionally a second flotation process, arranged to scavenge smaller sized particles and antimony- and metal value-bearing materials.

In one embodiment of the invention, the coarse concentrate has a particle size of >125 m. Preferably, the coarse concentrate has a particle size of >250 pm.

In one embodiment of the invention, the process for selective recovery of antimony comprises leaching the coarse concentrate with an alkaline sulphide solution, thereby producing the antimony-containing solution and the residue containing the remaining metal values, and separating the antimony-containing solution from the residue. The process further comprises the step of electrolyzing the separated antimony-containing solution to recover antimony metal.

Typically, the alkaline sulphide solution comprises 5-50 g/L OH- (g/L) and 50-400 g/L S2-. Typically, the leaching reaction is carried out at a temperature in the range of ambient to 105 C for a period of from 2 to 24 hours.

In one embodiment, prior to undergoing electrolysis, the antimony-containing solution is clarified and eluted through an adsorption medium for recovery of dissolved metal values. Typically, the clarified antimony-containing solution is eluted through the adsorption medium for recovery of dissolved precious metals, in particular gold.
In one embodiment, prior to undergoing electrolysis, the pH
of the antimony-containing solution is adjusted to alkaline. The pH of the solution is raised by dosing it with caustic solution. Typical examples of caustic solutions include sodium hydroxide, potassium hydroxide or calcium hydroxide.

In one embodiment, the electrolysis is performed with a voltage range of 3-4.5 V and a current density of 600-8 0 0 A/m2 at the anode and a current density of 2 5 0- 6 0 0 A/m2 at the cathode. The electrolysis extends for a period of 4-8 hours at 40-60 C.

Advantageously, in one embodiment of the invention, a by-product of the electrolysis reaction is sodium sulphate.

In one embodiment the residue containing the remaining metal values contains gold. Typical examples of the forms in which the gold resides in the residue are gold from the feed stream concentrate, gold liberated from the antimony-bearing material by step b), and gold associated with arsenopyrite. The gold may be recovered from the residue by subjecting the residue to cyanidation and forming a pregnant leach solution containing gold, followed by an extraction process. Typical examples of the extraction process include electrolysis, adsorption onto carbon or resins, or precipitation of the gold by addition of zinc to the pregnant leach solution, such as in the Merrill Crowe process. Prior to cyanidation, the residue may be subjected to an oxidative leach process to liberate gold from the arsenopyrite and other retractory minerals.

The present invention is also based on the surprising realisation that it is possible to improve metal recovery of softer metal value bearing materials from a harder gangue material by providing a coarse, low-grade feed stream concentrate of the mixed metal value bearing material which has a particle size greater than the particle size of conventional feed stream concentrates. This is counterintuitive to accepted practice, as there is typically an economic penalty for treating a coarse grind, low-grade feed stream concentrate in comparison to a finely ground, high-grade feed stream concentrate.

According to another aspect of the invention, there is provided a process for recovery of a metal value from a mixture of a soft metal value-bearing material and a gangue material, wherein the hardness of the soft metal value-bearing material is less than the hardness of the gangue material, the difference in hardness being such that the soft metal value-bearing material is preferentially ground relative to the gangue material, the process comprising the steps of:

a) providing a coarse concentrate comprising the soft metal value-bearing material; and, b) subjecting the coarse concentrate to a process for selective recovery of the metal value, thereby producing a recovered metal or a recovered metal-containing solution and a residue.
In one embodiment of the invention, the difference in hardness between the soft metal value-bearing material and the gangue material is greater than or equal to 2 Mohr units.

In another embodiment, the soft metal value-bearing material has a hardness of 2 Mohrs units or less.

In one embodiment the residue contains one or more further metal values, and the process further comprises the step of recovering the one or more further metal values from the residue.

Brief Description of the Drawings Preferred embodiments, incorporating all aspects of the invention, will now be described by way of example only with reference to the accompanying drawings, in which:
Scheme 1 represents a flow sheet diagram of one embodiment of the process of the invention;
Figure 1 is an X-ray Diffraction Analysis of a coarse concentrate as described in Example 1; and, Figure 2 is an X-ray Diffraction Analysis of a residue remaining after the coarse concentrate has undergone leaching with an alkaline sulphide solution as described in Example 1.

The flow sheet diagram of this embodiment of the invention is based on pilot plant testing of Hillgrove Mine, New South Wales in respect of gold bearing arsenopyrite stibnite ore. It is to be understood that for the purposes of this example stibnite is the antimony-bearing material and gold-bearing arsenopyrite is the metal value-bearing material. Additionally, it is to be understood that for the purposes of providing an example for the second aspect of the invention, the stibnite is the soft metal value-bearing material and the gold-bearing arsenopyrite is a second value-bearing material. The gangue material is the rock mass.

Detailed Description of the Preferred Embodiments of the Invention Before the preferred embodiment of the present process is described, it is understood that this invention is not limited to the particular apparatus described, as these may vary. It is also to be understood that the terminology used herein is for the purpose of describing the particular embodiment only, and is not intended to limit the scope of the present invention in any way. It must be noted that as used herein, the singular forms "a", "an", and "the"
include plural reference unless the context clearly dictates otherwise. Unless defined otherwise, all technical and scientific terms used herein have the same meanings as commonly understood by one of ordinary skill in the art to which this invention belongs.

Referring to Scheme 1, in accordance with various aspects of one embodiment of the present invention, an antimony-and metal value-bearing material is provided for processing. The antimony- and metal value-bearing material may be an ore, a concentrate, or any other material from which antimony and other metal values, in particular precious metals such as gold and silver and tungsten, may be recovered.

The antimony bearing material can be stibnite, auro-stibnite, tetrahedrite, or a combination thereof.
Alternatively, the antimony bearing material can be a metal value-bearing material containing antimony.

Native gold is frequently disseminated in quartz grains, often with metal sulphide ores such as for example pyrite (FeS2), chalcopyrite, galena, stibnite and arsenopyrite.
Accordingly, the preferred embodiment of the present invention is particularly advantageous in connection with the recovery of antimony and precious metals, such as gold, from mixed gold/antimony/arsenic ores or concentrates.

Thus, the antimony- and metal value-bearing material is preferably a mixed antimony/gold/metal sulphide ore.
Typically, the mixed antimony/gold/metal sulphide ore is a stibnite-arsenopyrite-pyrite ore containing gold.
Typically, the relative hardness of each of these minerals is very different. The hardness of stibnite is 2 Mohrs units, arsenopyrite is 3% to 6 Mohrs units, pyrite is 6 Mohrs units and quartz is 7 Mohrs units.

The antimony- and metal value-bearing material undergoes comminution, flotation, blending, and/or slurry formation as well as chemical and/or physical conditioning to afford a coarse concentrate. In contrast to conventional techniques for preparing concentrates, the particle size of the coarse concentrate of the present invention is >125 m.
Preferably, the particle size of the antimony- and metal value-bearing material of the coarse concentrate is >250 m. An advantageous feature of the invention is that conventional sulphide flotation practice may be employed without resorting to further grinding and clean-up processes.

In the preferred embodiment, the mixed antimony/gold/metal sulphide ore is typically treated via a crush/grind/float circuit to yield a concentrate containing an antimony feed grade of >15%, a gold feed grade of >35 g/t and, optionally, an arsenic grade of >4%.

The antimony- and metal value-bearing material undergoes a primary crush before being transferred to a grinding mill.
Any suitable crusher may be used to crush the antimony- and metal value-bearing material including a jaw crusher, high pressure grinding rolls, rolls crusher or a gyratory crusher for a primary crush throughput.

The crushed antimony- and metal value-bearing material is then transferred from the crusher to a grinder. Typical examples of a grinder are a grinding mill, a rod mill, a ball mill, an autogenous or a semi-autogenous (SAG) mill.
The ground antimony- and metal value-bearing material is screened at 3-4 mm with the undersize reporting to a gravity concentrate circuit and the oversize reporting to the grinder. Alternatively, the ground antimony- and metal value-bearing material may report to a cyclone for size separation.

The ground antimony- and metal value-bearing material having a particle size < 3-4 mm is then transferred to a gravity circuit comprising an enhanced gravity concentrator and a primary sizing cyclone. The concentrate formed in the enhanced gravity concentrator circuit reports to a combined feed concentrate stream for further processing in an alkaline sulphide leaching tank. Typically, the particle size of the concentrate is p80 of 150-600 m.

The gravity tail from the enhanced gravity concentrator is directed to the primary sizing cyclone, where the cyclone underflow (coarse particles of >250 m) is directed to a rougher flotation cell, while the cyclone overflow (fine particles of <250 m) reports to a scavenger flotation cell.

Before undergoing flotation, the ground antimony- and metal value-bearing material is optionally conditioned with suitable conditioners known to improve the overall effectiveness and efficiency of processing operations in accordance with the envisaged demands of the antimony- and metal value-bearing material to afford a concentrate feed stream. For example, the mixed gold-stibnite-arsenopyrite-pyrite is conditioned with lead nitrate or copper sulphate at a dose of 250-2,500 g/t, a xanthate collector at a dose of 30-250 g/t and frother. Typically, the ground antimony-and metal value-bearing material is conditioned for 1-5 minutes.

In the preferred embodiment, the mixed gold-stibnite arsenopyrite-pyrite is conditioned with lead nitrate (500-1,000 g/t), PAX collector (80-120 g/t) for 2-4 minutes.

The rougher flotation cell is arranged to capture coarse liberated free minerals (> 250 m), primarily the bulk of the softer antimony- and metal-value bearing materials, such as stibnite, while the scavenger flotation cell is arranged to capture the liberated mineral particles < 250 m. In the preferred embodiment, the scavenger flotation cell captures substantially the remainder of the residual stibnite and gold in the form of an arsenopyrite concentrate.

The rougher flotation cell operates as a coarse flotation cell with a flotation period of about 2-20 minutes, depending on the mineralogy of the concentrate. The rougher flotation cell tails report to the mill for regrinding. A flotation period of 10-15 minutes is typical for the mixed gold-stibnite-arsenopyrite-pyrite conditioned as described above.

The concentrates from each flotation cell and the gravity concentrate are then combined in an antimony leach feed hopper to produce a feed stream concentrate of approximately 5-15% of the feed mass before being directed to a leach reactor where it undergoes an alkaline sulphide leaching process.

The coarse concentrate is then subjected to leaching conditions to selectively recover the antimony in the form of a pregnant leach solution and a solid residue containing the remaining metal values. Antimony metal may be recovered from the pregnant leach solution by electrowinning, and the remaining metal values may be selectively recovered from the solid residue with suitable techniques pertaining to the metal value. For instance, where the solid residue contains gold, the solid residue may be subjected to cyanidation according to well known techniques.

The alkaline sulphide leaching process operates in a leaching vessel at a temperature in the range of about 25 C
to about 105 C at ambient pressure, over a period of 2-24 hours. The leaching solution contains 5-50 g/L OH- and 50-400 g/L S2-. In the preferred embodiment, the leaching solution contains 10-20 g/L OH- and 50-150 g/L Sz-, and the alkaline sulphide leaching process is conducted at 95-105 C
over a period of 6-8 hours.

The alkaline sulphide leaching of antimony sulfides is in accordance with the following reactions:

3Na2S + Sb2S3 (stibnite) -~ 2Na3SbS3 (1) 3Na2S + 4Cu2SSb2S3 (tetrahedrite) --> 4Cu2S + 2Na3SbS3 (2) During the alkaline sulphide leaching process, antimony is substantially fully solubilised to form a pregnant leach solution containing antimony. Typically, greater than 99%
of antimony in the coarse feed stream concentrate is leached into the pregnant leach solution.

Prior to recovery of antimony from the pregnant leach solution, the pregnant leach solution is filtered and clarified. Antimony is typically recovered from the pregnant leach solution by electrolysis in accordance with the following reaction:

4Na3SbS3 + 12NaOH -+ 4 Sb j + 12Na2S + 6H20 + 302 (3) Electrolysis is performed with a voltage range of 3-4.5 V
and a current density of 600-800 A/m2 at the anode and a current density of 250-600 A/m2 at the cathode for a period of 4-8 hours at 40-60 C.

In some instances, depending on the mineral composition of the feed stream concentrate, additional metal values, for example gold and arsenic, may be partially solubilised in the pregnant leach solution. It is envisaged that the other metal values will be recovered from the pregnant leach solution by well understood methods and techniques.
For example, where the solubilised metal value is gold, the solubilised gold is typically recovered from the pregnant leach solution by elution through an adsorption medium.

The metal values, in particular precious metal values such as gold, contained in the feed stream concentrate generally report to the solid residue formed during the alkaline sulphide leaching step. It is envisaged that the metal values will be recovered from the solid residue by well understood methods and techniques. Any gold values, for instance, will be recovered from the solid residue by re-pulping and treating the slurry with a cyanidation process, carbon-in-pulp (CIP), elution, electrowinning and smelting process.

Other metal values, if present in economically viable amounts in the residue, may also be subsequently recovered in accordance with well-understood techniques. For example, in Scheme 1, tungsten is recovered as a tungsten concentrate from the gold leach flotation tailings.

The preferred embodiment of the invention will now be further described with reference to the following Example.
The process conditions and parameters reflected therein are intended to exemplify various aspects of the invention, and are not intended to limit the scope of the claimed invention.

Example The antimony- and metal value-bearing materials of the example are a run-of-mine stibnite-arsenopyrite-pyrite ore containing on average 1% Sb, 0.8o As, and 4 g/t Au The mixed antimony/gold sulphide ore undergoes a primary crush to produce a particle size passing 10 cm at a 24 h throughput of approximately 1000 tonne. The crushed ore is then transferred by feed loader to a grinding mill where it is ground (initial throughput of 30 tph) and directed to a coarse cut cyclone. The cyclone overflow (approximately 250 pm) goes to a rougher flotation cell, while the cyclone underflow goes to a flash flotation cell for very coarse (250 pm - 6-7mm) mineral recovery.

The rougher concentrate is cleaned in a cleaner flotation cell and the cleaned concentrate is then combined with the flash flotation concentrate and fed to a centrifugal or gravity concentrator for free gold recovery.

A feed stream concentrate typically containing 17.25% Sb, 2.75% As and 1.01 o/t Au or higher is produced by the above-described circuit.

The rougher flotation cell is arranged to capture coarse free minerals (> 250 m), primarily the bulk of the stibnite. The rougher flotation cell operates with a conditioning time of about 3 minutes at neutral pH with PbNO3 (500 g/t), PAX collector (50g/t), and frother (IF50) with a flotation time of about 1 minute. The rougher/flash tails report back to the grinding mill for regrind.

The gravity concentrate tails are then fed to an antimony leach feed hopper where they are combined and settled to form the feed stream concentrate before being directed to the antimony leach reactor.

The feed stream concentrate (slurry solids content = 100 g/L) is treated with an alkaline sulfide solution (10 g/L
OH-, 50 g/L S2-) at 100 C for 8 hours resulting in a total mass reduction of 23.7%, and a residue analysis of 0.05%
Sb, 2.63% As, 0.76 o/t Au and 7.93% total S.

X-ray diffraction analysis was performed on the feed stream concentrate and the residue produced by the alkaline sulphide leaching process, and the results are shown in Figures 1 and 2, which confirm that stibnite and silica are the major phases of the feed stream concentrate while the residue contains silica and aluminium oxide as the major phases.

The slurry leaving the antimony leach reactor is separated into solid and liquid components for downstream treatment.
The solid residue (or filter cake) contains the majority of the gold values and is directed to the cyanidation circuit for repulping and gold recovery, whereas the liquid portion (or filtrate stream) of the slurry contains the antimony values and is directed as a pregnant leach solution to a diaphragmless electrolysis cell with mild steel anodes and cathodes. The pregnant leach solution contained Sb 17g/L, OH- 25 g/L, and S9ol,.,hle 55 g/L.

Antimony metal was deposited on the cathode of the diaphragmless electrolysis cell by applying a cathode current density of 600 A/m2, an anode current density of 800 A/m2, and a cell voltage of 3.4V. The electrowinning process was conducted at 50 C over a period of 2 hours with a Faradaic current efficiency of 64%.

The solid residue or filter cake separated after the alkaline sulphide leaching process is repulped and reslurried. The pH of the repulped cyanidation circuit feed stream is adjusted to a pH value of about 10.5 prior to introduction to the leach and adsorption section.
Sodium cyanide solution is dosed to facilitate gold extraction from the solid phase to the solution phase.

The leached gold solution is then electrowon in accordance with known techniques to recover the gold metal values.

In this example over 99.5% of the antimony was selectively leached from the feed concentrate, together with 29o As, 42% Au, and 47% S.

It is to be understood that, although prior art use and publications may be referred to herein, such reference does not constitute an admission that any of these form a part of the common general knowledge in the art, in Australia or any other country.

For the purposes of this specification and claims, it will be clearly understood that the word "comprising" means "including but not limited to", and that the word "comprises" has a corresponding meaning.

Numerous variations and modifications will suggest themselves to persons skilled in the relevant art, in addition to those already described, without departing from the basic inventive concepts. All such variations and modifications are to be considered within the scope of the present invention, the nature of which is to be determined from the foregoing description.

Claims

1. A process for recovery of antimony and metal values from antimony- and metal value-bearing materials, the process comprising the steps of:
(a) providing a coarse concentrate of the antimony-and metal value-bearing materials;
(b) subjecting the coarse concentrate to a process for selective recovery of antimony, thereby producing solid antimony or an antimony-containing solution, and a residue containing the remaining metal values; and, (c) recovering the remaining metal values from the residue.

2. The process according to claim 1, wherein the antimony bearing material is stibnite, auro-stibnite, tetrahedrite, or a combination thereof.

3. The process according to claim 1 or claim 2, wherein the metal value-bearing material is an ore, a concentrate, or any other material from which metal values may be recovered.

4. The process according to any one of the preceding claims wherein the metal value-bearing material is a metal sulphide containing one or more metal values.

5. The process according to claim 4, wherein the metal value-bearing material is a metal sulphide associated with precious metal value-bearing materials.

6. The process according to claim 5, wherein the metal sulphide associated with gold and/or silver are pyrites, chalcopyrites, and arsenopyrites.

7. The process according to any one of the preceding claims, wherein the coarse concentrate is prepared by crushing and grinding the antimony- and metal value-bearing materials, and subjecting the resulting materials to a flotation process.

8. The process according to claim 7, wherein the flotation process comprises a first flotation process arranged to capture coarse liberated particles of antimony-and metal value-bearing materials, and optionally a second flotation process, arranged to scavenge antimony- and metal value-bearing materials.

9. The process according to any one of the preceding claims, wherein the coarse concentrate contains an antimony feed grade at >15%.

10. The process according to any one of the preceding claims, wherein the coarse concentrate has a particle size of >125 µm.

11. The process according to any one of the preceding claims, wherein the process for selective recovery of antimony comprises leaching the coarse concentrate with an alkaline sulphide solution, thereby producing the antimony-containing solution and the residue containing the remaining metal values, and separating the antimony-containing solution from the residue.

12. The process according to claim 11, wherein the alkaline sulphide solution comprises 5-50 g/L OH- (g/L) and 50-400 g/L S2-.

13. The process according to claim 11 or claim 12, wherein the leaching reaction is carried out at a temperature of ambient to 105°C for a period of 2 to 24 hours.

14. The process according to any one of claims 11 to 13 further comprising the step of electrolyzing the separated antimony-containing solution to recover antimony metal.

15. The process according to claim 14, wherein prior to undergoing electrolysis, the antimony-containing solution is clarified and eluted through an adsorption medium for recovery of dissolved metal values.

16. The process according to claim 15, wherein the clarified antimony-containing solution is eluted through the adsorption medium for recovery of dissolved precious metals.

17. The process according to claim 14, wherein prior to undergoing electrolysis, the pH of the antimony-containing solution is adjusted to an alkaline pH.

18. The process according to claim 17, wherein the pH of the antimony-containing solution is raised by dosing it with caustic solution.

19. The process according to any one of claims 14 to 18, whereby the electrolysis is performed with a higher current density at the anode than the cathode.

20. The process according to any one of claims 14 to 19, wherein the electrolysis is performed with a voltage range of 3-4.5 V and a current density of 600-800 A/m2 at the anode and a current density of 250-600 A/m2 at the cathode.

21. The process according to any one of the preceding claims, wherein the residue containing the remaining metal values contains gold.

22. The process according to claim 21, wherein the gold is recovered from the residue by subjecting the residue to cyanidation and forming a pregnant leach solution containing gold, followed by an extraction process.

23. The process according to claim 22, wherein the extraction process is selected from a group comprising electrolysis, adsorption onto carbon or resins, or precipitation of the gold by addition of zinc to the pregnant leach solution, the Merrill Crowe process.

24. The process according to claim 22, wherein prior to cyanidation, the residue may be subjected to an oxidative leach process to liberate gold from arsenopyrite and other retractory minerals.

25. A process for recovery of a metal value from a mixture of a soft metal value-bearing material and a gangue material, wherein the hardness of the soft metal value-bearing material is less than the hardness of the gangue material, the difference in hardness being such that the soft metal value-bearing material is preferentially ground relative to the gangue material, the process comprising the steps of:
(a) providing a coarse concentrate comprising the soft metal value-bearing material; and, (b) subjecting the coarse concentrate to a process for selective recovery of the metal value, thereby producing a recovered metal or a recovered metal-containing solution and a residue.

26. The process according to claim 25, wherein the difference in hardness between the soft metal value-bearing material and the gangue material is greater than or equal to 2 Mohr units.

27. The process according to claim 25, wherein the soft metal value-bearing material has a hardness of 2 Mohr units or less.

28. The process according to any one of claims 25 to 27, wherein the residue contains one or more further metal values, and the process further comprises the step of recovering the one or more further metal values from the residue.