BACKGROUND OF THE INVENTION In the recovery of metal values from metal sulphide containing feed material, the feed material is typically first subjected to a pyrometallurgical process resulting in the formation of a sulphide matte containing the metal values, and a slag phase, and possibly also a fume.

The sulphide matte may then be typically subjected to a further pyrometallurgical process, or in some cases, a separate, hydrometallurgical process in which the matte is leached with a suitable lixiviant, preferably under high temperatures and/or pressures, and often under oxidising conditions.

Normally the pyrometallurgical and hydrometallurgical processes are conducted as separate, distinct operations, often at sites remote to each other.

The leaching process typically generates a pregnant, metal value containing leach solution and a solid phase containing leach residue and/or precipitates. As the solid phase principally contains the waste byproducts of the process, it is usually discarded. However, in order to maximise recovery of metal values into the leach solution, and thereby minimise the metal values in the residue, it is often necessary to extend the leach time or conduct the leach under harsh conditions-such as at very high temperatures and pressures. This requires the utilisation of large, expensive, specialist equipment such as an autoclave, and ancillary equipment including feed heating, feed pumps, oxygen injection, and equipment for temperature control, water balance, flash let down etc. This also results in high energy consumption.

It is an object of the present invention to provide a process for recovery of metal values from a feed material which overcomes, or at least alleviates, one or more of the above disadvantages of the prior art.

SUMMARY OF THE INVENTION According to the present invention, there is provided a process for the recovery of one or more metal values from a metal containing feed material, the process including the steps of: (a) subjecting said feed material to a pyrometallurgical treatment in which the feed material is smelted to form a matte containing said one or more metal values and a slag phase; (b) subjecting said matte to a hydrometallurgical treatment including the leaching of said matte in order to dissolve said one or more metal values into solution, said hydrometallurgical treatment resulting in a metal value containing pregnant leach solution and one or more solid phases; and (c) returning said one or more solid phases to step (a) as a flux for the pyrometallurgical treatment.

The present invention also provides metal values recovered using the process defined above.

As used herein, the term"matte"is defined as a fusion product produced from the pyrometallurgical treatment of the metal containing feed material, and contains metal sulphides, optionally together with metals and/or metal alloys.

The term"concentrate"is defined herein as being a metal value-containing product which is concentrated from ore after removal of gangue.

The process of the invention has the advantage that any metal values which are remaining in the solid phase after the leaching process can be returned to the recovery process for further treatment, instead of being discarded to waste. Further, because any unleached metals are returned to the pyrometallurgical treatment, very high overall recovery of metal values can be achieved by successive leachings of the matte using leaching conditions which are less severe than those utilised in prior art processes, or in a manner to reduce overall processing capital and/or operating costs.

DETAILED DESCRIPTION OF THE INVENTION In a preferred embodiment of the invention, the hydrometallurgical treatment further includes adding a neutralising agent to the pregnant leach solution in order to modify the pH thereof. Typically pH adjustment is effected in order to cause precipitation of one or more metals or impurities from solution, typically a metal other than those of interest. Any unwanted solid material

resulting from the neutralisation step, such as precipitate, residues etc., can be also recycled to smelting step (a) as a flux material for the slag phase. The metal value containing pregnant leach solution can then be treated using conventional chemical or electrolytical extraction means in order to selectively recover the metal values.

By recycling one or more unwanted solid phases from step (b) to the smelting step (a), any metal values contained in those residues or precipitates are returned to the recovery process for further treatment. Accordingly, the overall recovery of those metal values is enhanced, and the amount of waste material produced is minimised, thereby reducing capital and/or operating costs.

In a preferred form of the invention, the pyrometallurgical and hydrometallurgical stages are integrated into a single process, so that residues from the hydrometallurgical stage can be easily recycled into an earlier pyrometallurgical stage.

The invention is applicable for the recovery of base metal values, such as copper, cobalt and nickel possibly together with precious metals, such as silver, gold, platinum, palladium, (rhodium, ruthenium, osmium and iridium) from ores or concentrates, although it is not limited to this application. In one embodiment, the invention also encompasses recovery of lead.

In the case of recovering base metal values, the feed material containing the base metals typically comprises sulphide ores, concentrates and/or mattes.

The feed material may also contain lead values which are typically removed early in the pyrometallurgical treatment as fume. It is noted that any lead which does report to the matte during smelting will usually report to the solid residue remaining after the subsequent leaching step and is therefore recycled to the smelting step.

The smelting step (a) produces a matte containing the metal values. In the case of smelting base metal values, the matte usually comprises a mixture of copper, cobalt and nickel sulphides, possibly together with iron and sulphur.

It may also contain one or more metals and/or alloy phases. The matte may subsequently be separated into one or more segregated mattes and/or alloys by conventional means, with one or more segregated phases removed for use or further processing before the remaining matte is subjected to the hydrometallurgical treatment.

The metal value containing matte produced in step (a) is fed, after grinding if necessary, to a leaching stage in step (b). Although leaching may occur under atmospheric conditions, preferably the leaching is conducted under elevated temperature and/or pressure, for example, in an autoclave. The lixiviant may comprise an acid, preferably a mineral acid, such as sulphuric acid, or a base, such as ammonia. Preferably the lixiviant is an acid, more preferably sulphuric acid. Advantageously, the sulphuric acid is derived from waste gases evolved from furnaces used in the pyrometallurgical stage. The gases are typically cooled and cleaned ahead of a contact sulphuric acid plant.

The temperature of leaching may be from ambient to 270°C and is typically from 90°C to 270°C. Preferably, the temperature of leaching is from 90°C to 220°C.

In the case of processing base metal containing feed material, it is believed that the solid residue remaining after the leaching stage is likely to contain, in addition to any unrecovered metal values, various iron compounds such as iron oxides and/or hydroxides and/or sulphates, such as jarosite.

These solids are separated from the leaching solution and returned to the smelting stage step (a) where they function as a flux for the slag phase. Metal values remaining in the solid residue will generally be separated from the slag and incorporated into the matte.

Because the solid residue from the leaching stage is recycled and not discarded (as in conventional leaching processes), any metal values remaining in the solid residue will be returned to the pyrometallurgical stage, rather than be lost. Accordingly the operating conditions of the leach can beneficially be moderated because it is not critical to maximise the recovery of metal values during the leach. Where leaching is conducted under pressure in an autoclave, the temperature and pressure requirements can be reduced, thereby reducing such requirements as energy requirements, oxygen feed pressure, cooling and cost. Indeed, leaching can be effectively conducted partly or completely under atmospheric conditions, with high recovery, although the length of the process increases.

The metal value containing pregnant leach solution generated in step (b) may be treated with a neutralising agent. Where leaching was conducted using

an acidic lixiviant, the neutralising agent is an alkaline material such as a metal carbonate or hydroxide. The neutralisation may only be partial, in that the pH may only change from very acidic to acidic-for example, from pH = 1 to pH = 3.5. However, in one embodiment, the pH change can be greater, such that the pH of the neutralised solution is from 3.5 to 4.0.

Preferably, the neutralising agent is an alkaline material, more preferably is a metal carbonate such as a magnesium and/or calcium carbonate.

Preferred carbonates are limestone (CaCO3) dolomite (CaMg (C03) 2) or magnesite (MgCO3).

The neutralising agents may advantageously be naturally occurring, such as unpurified dolomite, calcite or magnesite which are convenient and relatively inexpensive to use. Any other metal containing phases present as impurities in the carbonates can be incorporated into the metal recovery process by returning any solid precipitate or residue from the neutralising stage to the smelting stage in step (a). For example, some naturally occurring dolomites contain cobalt and manganese. A solid residue containing cobalt and manganese can be recycled to the smelting stage where cobalt is incorporated into the sulphide matte within the rest of the cobalt value and, under suitable pyrometallurgical conditions, the majority of the (deleterious) manganese constitutes a flux and is incorporated into the slag phase, thereby being separated from the metal values. The solid residue remaining from such a neutralisation step typically includes sulphates, hydroxides and possibly oxides.

The increase in pH of an acidic, pregnant leach solution effected by neutralisation with an alkaline material such as carbonate can cause the precipitation of compounds from solution, in particular iron compounds, principally as iron oxides and/or hydroxides. The neutralisation therefore substantially removes undesirable soluble iron from the pregnant leach solution thereby facilitating subsequent extraction of the metal values of interest with the precipitate returned to smelting step (a) as a flux material. It is believed that a pH change from very acidic (e. g. pH = 1) to acidic (e. g. pH = 3.5) may be sufficient to precipitate most of the soluble iron. In prior art neutralisation processes, a maximum pH of about 3.5 has been found necessary in order to limit coprecipitation of cobalt and nickel compounds with the iron compounds (and any other impurities), which are typically discarded to waste. However,

this limitation is not as crucial to the present invention, because any coprecipitated cobalt and nickel will be returned to the pyrometallurgical stage for further processing, rather than being discarded to waste. Accordingly, the maximum pH of neutralisation can exceed 3.5, such as up to a pH value of 4.0.

This has the advantage of allowing greater removal of iron (and any other impurities) from the pregnant leach solution thereby minimising impurities in the final product.

The process of the invention is particularly beneficial in the recovery of metal values from polymetallic ores and concentrates which are normally difficult to treat using conventional extraction/recovery techniques. Fine grained, polymetallic ores must be crushed to a very fine particle size in order to adequately release and separate the different ore minerals from the gangue and from each other. However, fine particles cannot be successfully subjected to conventional flotation techniques to separate ore particles from gangue particles. Moreover, fine ore particles cannot be successfully subjected to conventional differential flotation techniques to separate different types of ore minerals from each other. The inventive process can avoid these problems by producing a bulk sulphide matte during the pyrometallurgical stage. The different types of sulphides in the matte can then be separated using hydrometallurgical techniques. Furthermore, ore minerals can be separated from gangue minerals, by using fine grained flotation techniques. Of particular interest is oil agglomeration, in which distillate, or other suitable oil, is used as a bridging agent to selectively agglomerate the ore minerals into"clumps"having a size suitable for flotation.

The advantage of the present invention over prior art metal recovery processes is illustrated by a comparison of prior art processes for cobalt recovery with that of the invention.

The invention is particularly advantageous for the direct extraction of cobalt from ore materials containing this element. There is currently very little or no direct processing of cobalt as there are few mines in the world from which cobalt is the principal product. Typically, cobalt is instead produced as a by- product of the processing of nickel, copper-nickel and copper-cobalt ores. In the pyrometallurgical processing of nickel or copper nickel ores, some cobalt is concentrated into a matte rather than being converted directly to metal, with

much of the cobalt being oxidised and reporting to the slag. Moreover, there is a relatively high proportion of slag to cobalt produced. Accordingly, there are significant losses of cobalt to the slag phase. The highest recovery of cobalt to the matte using conventional smelting techniques ranges from about 50% to about 75%.

Cobalt can also be produced hydrometallurgically from lateritic nickel (cobalt) deposits by leaching with ammoniacal ammonium carbonate. This process requires very high tonnages of feed material and produces very high amounts of residue, which are typically discarded. As a result, significant amounts of cobalt are lost in the residue, resulting in low recovery of Co, such as only about 40-50% recovery of cobalt.

Considerably higher recoveries of cobalt from laterites are possible using a high pressure acid leaching process. However this process suffers from the disadvantages of requiring high volume feed and waste materials, specialist equipment and high energy requirements, and is accordingly expensive.

In contrast, using the process of the invention, it is possible to recover in excess of 90% of cobalt values in the base metal ore, as compared to a highest recovery of 50-75% using prior art pyrometallurgical techniques. While better recoveries of cobalt are possible using prior art high pressure acid leaching of lateritic nickel (cobalt) deposits, the feed material has a very low concentration of cobalt. Accordingly, very high volumes of feed material must be processed in order to recover a given quantity of cobalt, with the attendant necessity for high volume equipment and leach solutions and high volumes of waste. In contrast, the pyrometallurgical stage of the present invention can considerably reduce the bulk of ore material (especially due to release of lead fume) resulting in a hydrometallurgical feed material more highly concentrated in cobalt. Moreover, any cobalt and/or nickel which is inadvertently coprecipitated with iron and other impurities and lost to waste in the prior processes, would be recoverable in the process of the invention.

Throughout the inventive process, the amount of feed material required for each subsequent step is consistently being reduced. This means that the metal values are becoming significantly more concentrated in the feed material for each step, leading to greater efficiency and less waste.

DESCRIPTION OF DRAWING The invention will become more readily apparent from the following exemplary description in connection with the accompanying drawing and Example.

FIGURE 1 is a flow chart illustrating the steps and their interrelationship of the process described in the Example.

EXAM PLE The process of the invention is used to extract metal values from a fine grained, polymetallic ore (herein identifie as"Brown's Ore") containing 3.3% lead, 0.6% copper, 0.13% cobalt and 0.11 % nickel. The ore also contains small quantities of zinc and silver. Mineralisation occurs as both oxide and sulphide minerals. The oxide ore is primarily a mixture of malachite and cerussite with minor pyromorphite, along with remnants of pyrite in a matrix of muscovite and quartz with smaller amounts of kaolinite, graphite, goethite, and biotite. Traces of covellite, chalcopyrite, galena, pyrite and pyrrhotite are also present along with tourmaline and zircon. The cobalt and nickel in the oxide ore are included with goethite as a cobalt rich mineral which is optically similar to pure goethite.

Some cobalt and nickel is also associated with the remnant pyrite. The major sulphide minerals present are galena, pyrite, chalcopyrite with some covellite and digenite and a cobalt nickel sulphide, identified as siegenite. The major gangue minerals are feldspar, mica, quartz and graphite.

The oxide component of the Brown's Ore is mined and treated separately to the sulphide component. The oxide ore is leached separately and the leach solutions subsequently added to the pregnant leach solution arising from the treatment of sulphide ores, discussed below.

The sulphide component of the Brown's Ore is finely ground, preferably to a particle size of P8o of-30 microns. Due to the very fine grained nature of the ore, grinding to a small particle size is necessary to ensure that the ore minerals are adequately released. However, the small particle size of the ground ore means that conventional flotation methods cannot be used to separate the ore from the gangue, or indeed to separate the different ore minerals from each other. Instead the ground ore is first subjected to oil agglomeration, in which distillate is used as a bridging agent to selectively agglomerate the"collected"sulphides and native graphite to a size suitable for

flotation. The agglomerated particles are then subjected to flotation in a flotation machine to separate ore from gangue and to produce a concentrate containing mainly lead, copper, cobalt and nickel ore minerals, with small amounts of zinc and silver.

The concentrate is then fed into a first furnace where it undergoes direct smelting. Direct smelting is the combination of a number of different smelting processes in a single furnace. The principle smelting processes occurring in the first furnace includes the initial oxidation of metal sulphides and the subsequent reduction to metal. These processes are effected by utilising "submerged lance technology"and"submerged bath technology"which are licensed by Ausmelt Limited.

During the direct smelting in the first furnace, lead is fumed off under neutral to slightly reducing conditions and copper, cobalt and nickel are formed into a sulphide matte. The lead fume from the first furnace is combined with recycled lead fume from a second furnace, briquette and reduced to lead bullion in the second furnace. Two stages of kettle refining produce a 99.6% Pb bullion.

Most of the zinc present in the concentrate reports to the slag, with the rest either fuming off or a very small amount reporting to the matte. The very low concentration of zinc in the matte simplifies and therefore reduces operation costs in the subsequent hydrometallurgical treatment.

The recovery of copper, cobalt and nickel to the sulphide matte is high- around 95% or higher. The sulphide matte typically contains approximately 37% copper, 12% cobalt, 11% nickel, 13% iron and 27% sulphur. Alternatively, the matte may be subjected to controlled segregation cooling in order to produce a copper matte and a cobalt nickel matte, and/or a metal-cobalt alloy.

The Cu-Co-Ni matte is tapped, granulated and ground to form a feed material for the subsequent hydrometallurgical stage. The ground feed material is subjected to pressure oxidation acid leaching in a small conventional autoclave at a rate of 3.4 tonnes/hour. The autoclave is operated at around 210°C with oxygen injection to provide an operating pressure of around 3,000 kPa. Leaching is effected using sulphuric acid derived from furnace off-gases during smelting. The residence time is around 60 minutes. These operating conditions are significantly more moderate than those required in the

conventional high pressure leaching of nickel laterites, for example at a temperature of around 250°C and a pressure of around 4000 kPa.

The solid residue remaining from the leach process typically contains various iron compounds, such as iron oxides and/or hydroxides and/or sulphates, as well as any unrecovered metal values. The residue is separated from the leach liquor and returned to the first furnace in the pyrometallurgical stage where it functions as a flux for the smelting process. Unrecovered metal values present in the leach residue are also returned to the smelting process and incorporated into the sulphide matte there produced.

The acidic leach liquor is subsequently partially neutralised by addition of a locally occurring cobalt containing dolomite. The neutralisation reaction results in formation of a solid precipitate which largely comprises iron compounds although, some Co and Ni may be coprecipitated. The precipitate, and any solid residue remaining after pH adjustment, are also recycled to the pyrometallurgical stage as flux. Moreover, any metal values derived from the dolomite (e. g. cobalt) and present in the solid residue are incorporated into the sulphide matte produced during the smelting stage, and thereby enhance the overall recovery of those metal values.

The return of leach residue and/or precipitates to the smelting stage facilitates optimum pressure leaching conditions because the recovery of maximum cobalt from the leaching stage is not critical. Similarly, the precipitation and solution purification stages by pH adjustment can be optimised for iron and impurity precipitation as any coprecipitated cobalt and nickel is returned to the pyrometallurgical stage.

The pregnant leach solution is subjected to a series of solvent extraction and/or electrowinning processes, separated by purification steps, in order to recover the metal values of interest. The precipitates from the purification steps were also returned to the pyrometallurgical stage.

Copper is recovered by copper solvent extraction using a chemical consisting of a phenolic oxime derivative having the trade name LIX622N, distributed by Cognis Australia Pty. Ltd. The copper raffinat is then returned to the autoclave to increase the cobalt concentration to about 12.6 g/l and the nickel concentration to about 8.3 g/l. A bleed stream is passed to secondary iron precipitation and then secondary copper solvent extraction.

Zinc, cobalt and nickel are recovered by solvent extraction using a bis (2,4,4-trimethylpentyl) phosphonic acid sold under the trade name Cyanex 272 distributed by Cytec Australia Ltd. Differential extraction is effected by varying the pH of the aqueous leach solution.

The overall percentage recoveries of the metal values from the Brown's Ore are in excess of 90% for copper, cobalt and lead and in excess of 70% for nickel. These recoveries, especially those for cobalt, are a significant improvement over the recoveries using conventional extraction processes.

In summary, the advantages of the present invention include: (i) high recovery of metal values; (ii) a reduction in waste byproducts; (iii) optimised hydrometallurgical processing under moderate conditions; (iv) effective use of solid residues as flux in the pyrometallurgical stage-this enables the use of e. g. dolomite both as a neutralising agent and a flux, thereby maximising the utilisation of resources; and (v) reduced operating costs and capital expenditure.

Finally, it is to be understood that various modifications and/or alterations may be made without departing from the spirit of the present invention as outlined herein.