PROCESS AND SOLUTION FIELD OF THE INVENTION This invention is concerned with a solution and a process for extracting metals from metal containing materials. The invention is particularly concerned with a solution and process for extracting metals by dissolution and/or disproportionation of a metal mixture, metal alloy or metal compound, thereby releasing metal or metals for recovery in such forms as metallic particles, metal ions or precipitation as metal compounds. The process finds particular application in the extraction of metals via an oxidising step.

The process and solution may be used in the extraction of metals from a range of different metal compounds, such as from metal oxides and metal sulphides. However, the invention finds particular application in extracting metals from metal sulphides, such as sulphide ores and concentrates. The process finds particular application in the rapid and economic extraction of valuable metals from difficult to treat refractory type polymetallic ores and concentrates and particularly where dissolution of unwanted compounds contained in these is undesirable.

Such refractory type ores include refractory pyrite (FeS), chalcopyrite (CuFeS2) and arsenopyrite (FeAsS), which act as host minerals for inclusions of valable metals such as silver and gold. For example, arsenopyrite is reacted to form ferric arsenate, FeAsO4, thus allowing easy access to the valable metals for recovery.

BACKGROUND OF THE INVENTION The chemical extraction of metal values from materials via an oxidising step, by processes which do not involve the specific use of an oxidising acid, such as nitric acid, is of great commercial interest. However, the use of oxidising acids or direct oxidation agents, such as hydrogen peroxide, is costly. It has accordingly been proposed to instead use non-oxidising solvents, such as acids (e. g. sulphuric acid) or alkalis (e. g. ammonia solution), incorporating a less expensive oxidising agent, such as air or oxygen, in conjunction with a chemical or biological oxidants. Typically, such processes have involved either reaction of the non-oxidising acid with the material in the presence of oxidising gases or bacteria, or by reaction of the non-oxidising acid with the material in the presence

of the stoichiometric quantity of the chemical oxidant, often combined with a subsequent step for the regeneration of the chemical oxidant.

In the extraction of metals from metal sulphides, particularly from ore minerals, metal ions, such as ferric ions or cupric ions, can advantageously be used as a chemical oxidant. Ferric ions, in the form of ferric sulphate, is cheap, readily available in some form at many mine sites, and can be readily regenerated chemically, or by bacteria and/or aeration. Indeed, iron is frequently generated in solution as a by-product of other leaching processes. Copper ions similarly can be made readily available and also can be generated in solution as a by-product of other leaching processes.

Cupric and ferric ions react with metal sulphides in order to release the metal into solution, in accordance with the following reactions: MeS + 2fie3+ 2+ + 2Fe2+ + S° MeS + 2Cu2+ o Me2+ + 2Cu+ + S° The ferrous and cuprous ions generated by these processes can be re-oxidised to ferric and cupric ions, respectively using oxygen in order to regenerate the chemical oxidants.

Of significant importance in many cases is that in the leaching and oxidant regenerating reactions, a minimum of the sulphur contained in the material is oxidised to sulphate. This is to avoid sulphate build-up in the leach circuit which necessitates expensive sulphate removal steps.

These processes are limited by the speed of the regenerating step, which is often very slow, are stoichiometrically inefficient, and frequently require a different pH range for oxidant regeneration to that required for effective leaching, thus requiring expensive pH adjustment.

The use of a chemical oxidant, particularly in acid mediums, can be disadvantageous in that it can result in the introduction of a non-compatible contaminant into the solution. A recovery or removal step then becomes necessary to separate the contaminant from the desired metal product, which in concentrated solutions entails an expensive neutralising stage. This is particularly the case with contamination by iron which is often used as a chemical oxidant because of its low cost and ready availability and in many cases is dissolved from the feed material along with desired metal or metals to be leached.

There is clearly a need for a more rapid processing method, requiring simple but effective processing equipment, which minimises contamination from chemical oxidants, or from dissolution of undesired compounds which can operate with greater stoichiometric efficiency and with minimal sulphur oxidation.

It is accordingly, an object of the present invention to provide a process and a solution for extracting a metal from a material containing that metal, which overcome, or at least alleviate, one or more disadvantages of the prior art.

SUMMARY OF THE INVENTION According to the present invention, there is provided a process for extracting a first metal from a material containing said first metal, including the step of treating said material with an aqueous alkaline solution containing an oxidising agent and complexes of ions of one or more Transition Elements with ammonia.

The present invention also provides a solution for extracting a first metal from a material containing said first metal, said solution being an aqueous alkaline solution containing complexes of ions of one or more Transition Elements with ammonia.

DETAILED DESCRIPTION OF THE INVENTION As used herein, the term"Transition Elements"is intended to include the main transition elements, or first transition series, from Sc to Cu; the lanthanide elements, or second transition series, from Y to Ag; and the actinide elements, or third transition series, from Hf to Au. Preferably however, the Transition Element is selected from the main transition elements. The Transition Element may be selected from the group comprising copper, nickel, zinc and cobalt. However, typically, the Transition metal is copper and it is present in solution as cupric ions, Cu2+.

It is preferred that the complexes of the Transition Element ions with ammonia are formed through addition of ammonia to the solution, rather than the addition of ammonium salts.

In the preferred form of the invention, the copper ions form complexes with ammonia, in particular: Cu (NH3) 2+ herein referred to as cupric ammine complexes. The copper may be added to the solution, or generated in situ, such as by reaction with a copper containing substance. The cupric ammine

complexes react with the material by oxidation, which releases said first metal into the solution. This oxidation reaction results in reduction of cupric ions (Cu2+) to cuprous ions (Cu+). The oxidising agent then oxidises cuprous ions to cupric ions, thereby replenishing the cupric ammine complexes for further oxidation of the material.

The pH of solution is above 7. Preferably, the pH of solution is (at least initially) above 8. More preferably the pH of the solution is at least initially in the range of 8 to 12, such as from 8.5 to 10.5. However, in some embodiments, the solution pH is allowed to decrease from its initial value over the course of the extraction process. This will be discussed in more detail later.

The oxidising agent is typically an oxidising gas, such as air or oxygen.

The oxidising gas may be introduced by dissolving it in the solution, or by bubbling it through the solution. Preferably, the oxidising gas is introduced into solution as very fine bubbles. The amount of the oxidising gas is preferably at least equal to the stoichiometric amount required to oxidise the Transition Element ions. More preferably, the amount of oxidising gas, such as 02, is between 1 and 5 times that stoichiometric amount.

Typically, the rate of addition of 0 per litre of solution is between 0.1 and 80 I/hour, in the case of 02 (gas), and between 1 and 350 I/hour in the case of air.

Preferably, the rate of addition of the oxidising gas should not be so high as to force NH3 from the solution.

The process preferably is preceded by the step of grinding the material.

Typically, the material is ground to a particle size of less than 40 microns, preferably less than 10 microns. More preferably, the material is ground to such an extent that the particle size of at least 80% of the ground material is 5 microns or less. The step of grinding increases the surface area available for reaction with the solution and enhances the rate and extent of metal extraction. While it is preferred that the material is ground before commencement of the process, the process may be at least partially conducted using unground material. In such an embodiment, the residue remaining from the process is advantageously removed from the leach solution, ground, then returned to the leaching process.

The solution is preferably agitated during the process. The solution of the invention is preferably reacted with the metal containing material in a device which permits agitation of the components. Agitation ensures adequate suspension of the metal containing material in solution and may also ensure the dispersion of an admitted oxidising agent, such as air/oxygen, either into solution, or as fine bubbles. Preferably, a gaseous agent is introduced into the solution by bubbling in a continuous or semi-continuous stream. Such introduction can thereby cause continuous or semi-continuous oxidation of the Transition metal to its higher oxidation state. The solution containing the Transition metal ions, either added in the higher oxidation state, or continuously or semi-continuously generated in situ as described, causes oxidation or partial oxidation of the metal containing material, and dissolution or release of the said metals from the metal containing material, by reactions believed to proceed in the following general manner: MeS + 2Xn+ o Me++ + 2X (n-') 02 and X-). X where in this instance, X is a transition metal and Me is a divalent metal present in the material as a sulphide.

Metals which can be extracted using the process and solution of the invention include copper, lead, zinc, iron, cobalt, silver, gold, antimony, arsenic and barium. However, the process of the invention is particularly useful for extracting copper from a copper sulphide material, such as chalcopyrite, utilising a cupric ammine solution.

It is believed that the copper contained in the solution is present in the forms Cu (NH3) ++ (oxidised state) and Cu (NH3) 4+ (reduced state). In the preferred pH range for this reaction (i. e. 8.5 to 10.5), the Fe released into solution precipitates as a ferric oxide hydrate, leaving a very low Fe concentration, such as less than 10 ppm, preferably less than 5 ppm in the leaching solution. The very low Fe concentration is advantageous as it means that the impurity level of the copper containing solution is low. Moreover, the iron precipitate can be very simply removed by using high strength magnetic separation.

Without wishing to limit the process to a particular reaction mechanism, it is believed that the reactions taking place during the process are: CuFeS2 + 5Cu (NH3) 4 2+ o CU2+ + Fe3+ + 5Cu (NH3) 4 + + 2S° and/or 5Cu (NH3) 4 + + 5/4 °2 + 5/2 Ho o 5Cu (NH3) 4 2+ + 50H- in which the cupric ammine complexes oxidise cuprous and sulphide ions in the chalcopyrite to cupric ions and elemental sulphur respectively, and is itself reduced to cuprous ammine complexes. The cuprous ammine complex is then oxidised back to cupric ammine complex by reaction with oxygen.

The process of the invention can be conducted effectively at atmospheric pressure and at ambient to moderate temperatures. In many embodiments of the invention, the temperature of the process is substantially lower than temperatures typically used in prior art leaching processes. While the process can proceed at any temperature from room temperature to 100°C, an at least initial reaction temperature of 50°C is preferred. Given that most leaching reactions using the process of the invention are shown to be exothermic, once the leaching reaction commences, (as would be experienced in a continuous process) the process will generate its own heat, so that external heating can be discontinued and the reaction will still proceed at an acceptable rate. This is to be compared with many prior art extraction processes which require application of heat. Accordingly, the present process requires less energy.

The initial concentration of the Transition Element in the aqueous alkaline solution is preferably at least 0.5 g/l. More preferably, it is 0.5 to 1.5 g/I. The concentration of the Transition Element may increase over the course of the leaching process, depending on the composition of the metal containing material.

For example, where the Transition Element is copper and the metal containing material is chalcopyrite, during the course of the leaching process, the concentration of copper in solution will increase. However, the initial concentration of copper in the aqueous alkaline solution prior to commencement of the leaching process is preferably 0.5 to 1.5 g/I.

A preferred upper limit on the initial concentration of the Transition Element is around 20 g/I.

The concentration of ammonia is preferably sufficient to form the cupric ammine complexes Cu (NH3) 42+. Typically this requires an amount of NH3 which is from about 1 to about 4 times the stoichiometric quantity of Cu in solution. The concentration of NH3 should preferably also be sufficient to ensure an alkaline solution pH, preferably above a pH of 8.

Accelerants, surfactants, various other catalytic aids such as other cations or anions, defoamers and attriting aids may also be used, depending on the material to be leached. These may function in increasing the rate of oxidation/reduction of the material to be leached or of the Transition metal ions, by decreasing foaming, or by increasing the surface contact between the solution and the material to be leached, by dispersion of reaction products from the reaction sites, or by particle size reduction of the material as it is leached.

The form of metals extracted from solution can be varied by controlling process conditions during the extraction process.

In the case of extraction of Cu from chalcopyrite, the solution pH during the course of the leaching reaction will dictate the form extracted copper will take. If pH of solution is maintained at a high value, such as within the preferred pH range, the extracted copper remains in solution, usually as copper ammine complexes. However, if solution pH is allowed to decrease, during the course of the reaction a copper compound precipitates out of solution. It is thought that the copper compound may be a copper oxide and/or hydroxide and/or a basic copper sulphate compound. In this case, both a copper compound and an iron compound will precipitate out of solution. The iron compound can be separated from the copper compound by magnetic separation.

The decrease in solution pH may occur in a number of ways, but typically by not replenishing NH3 which can be lost from solution by evaporation caused by agitation during the extraction process, or by some sulphation.

The precipitation of both iron and copper compounds from solution is advantageous because it facilitates the separation of other metal species (such as Co, Ni) which remain dissolved in solution.

An advantage of the present invention is the provision of a simple leach process which is able to be performed in a single stage. Another advantage is a rapid leaching rate with high recovery, at atmospheric pressure using inexpensive

reagents and equipment. A further advantage is the ability to provide rapid and economic extraction of metal from a wide range of metal containing materials, including metal containing wastes, from a wide range of mining, metallurgical, chemical or manufacturing processes, including tailings, low grade concentrates, drosses, metal precipitates and metal scrap.

Further advantages of the process and solution of the invention include rapid reaction rates, leading to high recoveries of metal values and minimal levels of sulphur oxidation in the case of metal sulphides. The invention also has the advantage of very low levels of solution contamination from unwanted compounds, which compounds are normally leached from the metal containing materials in acid-based processes. Accordingly, there is provided alternate metal recovery circuits which are not normally available.

Another significant advantage of the invention is that the undesirable co- products of leaching (principally iron) can be separated from solution without the use of chemical reactants. Furthermore, desirable co-products (e. g., cobalt, etc) are present in solution in a form which makes their recovery in a pure state achievable economically. Thus, the resultant solution containing the desired extracted metals may be directly subjected to an electrowinning process, without the need for a costly solvent extraction step.

Also for many cases, a much smaller and compact plant to achieve equivalent metal recovery rates leading to lower capital and operating costs is achievable.

The invention will become more readily apparent from the following exemplary description in connection with the accompanying drawings and Examples.

DESCRIPTION OF DRAWINGS Figure 1 is a plot of the results of Example 1 and depicts rate of recovery of copper from a chalcopyrite copper concentrate under conditions of high pH.

Figure 2 is a plot of the results of Example 2 and depicts rate of recovery of copper from a chalcopyrite copper concentrate under conditions of low pH.

Figure 3 is a plot of the results of Example 3 and depicts rate of recovery of copper from a chalcopyrite copper concentrate and the rise in solution temperature due to the exothermic nature of the reaction.

Figure 4 is a plot of the results of Examples 4 and 5 and compares rate of recovery using air and oxygen, respectively, as the oxidising agent.

Figure 5 is a plot of the results from the first stage of Example 6 and depicts the rate of recovery of lead, zinc and copper from a concentrate treated with a proprietary sulphuric acid based leach solution.

Figure 6 is a plot of the results from the second stage of Example 6 and depicts the recovery of copper from the leach residue of the first stage using a copper ammine solution of the invention.

Figure 7 is a plot of the results from Example 7 and depicts the rate of recovery of lead, zinc and copper from the same concentrate as used in the first stage of Example 6, which is treated with a copper ammine solution of the invention.

Figure 8 is a plot of the results from Example 10 and depicts the rate of recovery of copper from a copper-gold concentrate using a cupric ammine solution of the invention.

Figure 9 is a plot of the results from Comparative Example 3 and depicts the recovery of copper and gold from a concentrate using direct cyanide treatment.

Figure 10 is a plot of the results from Example 11 and depicts the recovery of gold, silver and copper from the leach residue of Example 10 using cyanide treatment.

Figure 11 is a plot of the results of Example 12 and Comparative Example 4 and depicts the benefit of pretreating gold bearing sulphide concentrate with the process of the invention prior to conventional gold extraction.

EXAMPLES Comparative Example 1 An example of extraction of copper from a chalcopyrite concentrate using a conventional acid leaching process will be described. A 100 g sample of chalcopyrite concentrate was leached with a 1600 ml leaching solution containing 50 91-'sulphuric acid and 5 91-: iron. The concentrate was added to the leaching solution at a temperature of 90°C and with an oxygen flow rate of 20 I/hr. The recovery of copper after 5 hours was approximately 60%.

Example 1: Example 1 demonstrates the performance of the process under conditions of high pH, with all the copper going into solution. A leaching solution containing 0.6 gm copper ions and 300 gm of a 25% NH3 solution was diluted to 1.6 litre. 100 gm of finely ground (p80,5 micron) chalcopyrite copper concentrate was added and the solution agitated whilst oxygen gas was introduced into the mixing zone at a rate of 20 L/hr. The reaction was heated to 60°C prior to the commencement of the leach and rose to 76°C during the leach. The recovery rate curve is depicted in Figure 1 with a final recovery (corrected for residue analysis) of 99.90% Cu achieved.

Example 2: Example 2 demonstrates the performance of the process at low pH, where copper leached from the chalcopyrite concentrate is precipitated as a dilute acid soluble salt which is amenable to recovery of copper using conventional Solvent Extraction and Electrowinning processes. A leaching solution containing 0.6 gms copper ions and 120 gm of a 25% NH3 solution was diluted to 1.6 litre. 100 gm of finely ground (p80,5 micron) chalcopyrite copper concentrate was added and the solution agitated whilst oxygen gas was introduced into the mixing zone at a rate of 20 L/hr. The reaction was heated to 60°C prior to the commencement of the leach and maintained at 60°C. No further external heating was applied. The total recovery rate and solution recovery rates are depicted in Figure 2.

Example 3: Example 3 demonstrates the exothermic nature of the process. A leaching solution containing 1.0 gm copper ions and 300 gm of a 25% NH3 solution was diluted to 1.6 litre. 100 gm of finely ground (p80,5 micron) chalcopyrite copper concentrate was added and the solution agitated, whilst oxygen gas was introduced into the mixing zone at a rate of 20 L/hr. The leach was commenced at room temperature and no external heating was applied. The recovery rate is depicted in Figure 3 with a final recovery (corrected for residue analysis) of 99.75% Cu achieved. As most prior art copper extraction processes require the application of heat, the present invention advantageously requires minimal energy input.

Example 4 and Example 5: These Examples demonstrate the effectiveness of air as a source of oxygen when compared with pure oxygen gas as the oxidant. A leaching solution containing 1.2 gm/L copper ions and 300 gm of a 25% NH3 solution was diluted to 1.6 litre. 100 gm of finely ground (p80,5 micron) copper/zinc"bulk"concentrate was added and the solution agitated whilst, in the case of Example 4, air was introduced into the mixing zone at a rate of 90 L/hr, and in the case of Example 5, oxygen gas was introduced into the mixing zone at a rate of 20 L/hr. No external heating was applied. The recovery rate curves are depicted in Figure 4. While the reaction time using air as an oxidant is slightly longer, it is still effective and its lower cost means it is often more economical to use air as an oxidant, rather than oxygen.

Examples 6 and 7: The ability to separate metals in a polymetallic ore is demonstrated: Example 6: In Example 6, a proprietary sulphuric acid-based leach solution is first used to oxidise the Galena and Sphalerite from a mixed zinc/lead/copper/silver/gold concentrate components as a first stage, and subsequently to oxidise the copper minerals using the process of the invention in a second stage.

100g of the concentrate was leached in 1.6 litre of an aqueous solution containing 100g/l H2S04 and 5g/i Fe at 70°C for 2 hours. Oxygen gas was introduced at a rate of 60 I/hour. The rate of copper, zinc and lead oxidation is presented in Table 1.

TABLE 1 Time Elapsed Metal Oxidation Achieved (% of original) Zn Cu Pb 30 min 35.0 2.5 79.0 60 min 55.0 5.0 99.0 90 min 81.0 7.0 100.0 120 min 99.9 8.0 100.0 The results are plotted in Figure 5. The galena was rapidly oxidised to insoluble lead sulphate. The sphalerite was oxidised to soluble zinc sulphate and

8% of the copper content oxidised, probably representing the fine size fraction (<10 micron) present in the concentrate.

A process for selective recovery of Zinc can be developed from this leach method by precipitation of soluble copper with zinc scrap.

The pulp was filtered, and washed, and the leach residue was leached with the aforementioned copper ammine solution using 1.6 litre of a solution containing 1.5 g/I Cu ions at a pH of 10.0 for 90 minutes at 40°C. Oxygen gas was introduced at a rate of 20 I/hr. The rate of copper oxidation is presented in Table 2.

TABLE 2 Time Elapsed Copper Oxidation Achieved (% of original) 0 min 8.0 30 min 45.0 60 min 78.0 90 min 99.9 The results are plotted in Figure 6.

Optimisation of overall leach conditions and sequence then becomes a matter of economics.

Example 7: In Example 7 the same polymetallic concentrate as used in Example 6 was leached in a copper ammine solution as follows. 100 gm of concentrate was leached in a 1.6 litre solution containing 1.5 g/l Cu ions at a pH of 10.0 for 120 minutes at 40°C. Oxygen gas was introduced at a rate of 20 I/hr. The rate of metal oxidation is presented in Table 3: TABLE 3 Time Elapsed Metal Oxidation Achieved (% of original) Zn Cu Pb 30 min 35.0 30.0 79.0 60 min 75.0 60.0 99.0 90 min 99.0 80.0 100.0 120 min 99.9 99.9 100.0 The results are plotted in Figure 7. Lead zinc and copper were all oxidised during the leach. Zinc and copper can be selectively recovered. The lead

remains in the leach residue but can be recovered by a subsequent releach of the residue.

Examples 8 to 11 and Comparative Examples 2 and 3: The recovery of precious metals from refractory-type copper mineral host ores using conventional sodium cyanide leaching is sometimes ineffective or not cost effective due to cyanide consumption by reaction with copper ions. The use of the process of the invention can overcome these problems by rapid and economic leaching of the copper component, enabling subsequent recovery of the precious metals as shown below.

Examples 8 and 9 and Comparative Example 2: In Examples 8 and 9, the leached residue, remaining after the leaching of concentrates in Examples 6 and 7, respectively was treated at 20% pulp density with a solution containing 10 kg/tonne NaCN at pH of 12 for 18 hours.

As a comparison, in Comparative Example 2, the same concentrate as was used in Examples 6 and 7 was directly treated with the NaCN solution under the same conditions as for Examples 8 and 9.

The concentrations of copper, gold and silver over time in the leach solutions of Comparative Example 2 and Examples 8 and 9 are presented in Table 4. The precious metal recoveries were confirmed by fire assay.

TABLE 4 Time Elapsed 2hr 4hr 6hr 18hr Comparative Example 2 Cu ppm 150 160 200 300 Au % Recovered 10.0 20.0 50.0 99.9 Ag % Recovered 4.5 5.0 5.5 8.0 Example 8 Cu ppm 70 75 75 75 Au % Recovered 75.0 80.0 85.0 99.9 Ag % Recovered 85.0 93.0 96.0 99.9 Example 9 Cu ppm 60 65 70 70 Au % Recovered 75.0 81.0 86.0 99.9 Ag % Recovered 85.0 90.0 95.0 99.9 As can be appreciated from Table 4, the treatment of the polymetallic concentrate with the process of the invention prior to the cyanide treatment results in a much higher recovery of precious metals, and a much lower

concentration of copper in the leach solution (see Examples 8 and 9), compared with the direct cyanide treatment of concentrate without a prior treatment according to the invention (without a prior Comparative Example 2).

The lower concentration of Cu in the leach solutions of Examples 8 and 9 is advantageous in that less cyanide is consumed through reaction with soluble copper, thereby minimising the cost and inconvenience of subsequent treatment of the solution to remove copper cyanide.

Examples 10 and 11 and Comparative Example 3: In Example 10,100 gm of a copper-gold concentrate was leached in 1.6 litre of a cupric ammine solution containing 1.5 g/l Cu ions at a pH of 10.0 for 120 minutes at 40°C. Oxygen gas was introduced at a rate of 20 I/hour. The rate of metal oxidation is presented in Table 5.

TABLE 5 Example 10 Time Elapsed Copper Oxidised (%) 30 min 59.0 60 min 96.0 90 min 99.9 120 min 99.9 The results are plotted in Figure 8 and show that very rapid copper oxidation occurred, essentially all copper being solubilised after 90 min. Iron levels in the final solution were <2ppm.

In Example 11, the leach residue remaining from Example 10 was subjected to cyanide treatment at a 20% pulp density with a solution having 10kg/tonne NaCN at a pH of 12. As a comparison, Comparative Example 3 entailed direct cyanide treatment of 100g of the same copper gold concentrate as used in Example 10, under the same conditions of cyanide treatment as in Example 11. The results obtained are presented in Table 6. All precious metal recoveries were confirmed by fire assay.

TABLE 6 Comparative Example 3 Time lapsed 1hr 2hr 4hr 16hr 24hr Cu ppm 700 780 770 720 660 Au % Recovered 0.5 0.8 2.5 25.0 74.0 Ag % Recovered 0.1 0.2 0.5 1.0 2.0 Example 11 30 min 1hr 2hr 4hr 12hr Cu ppm 120 115 110 95 80 Au % Recovered 65.0 70.0 75.0 85.0 99.8 Ag % Recovered 63.0 75.0 83.0 92.0 99.9 The results are plotted in Figures 9 and 10 respectively, and show that the problem of excessive copper pickup into cyanide solution has been minimised, and that the gold and silver were rapidly recovered by the cyanide solution.

Example 12 and Comparative Example 4: This leach demonstrates the effectiveness in the pre-treatment of refractory gold-bearing ores to improve the accessibility of conventional extractants to the contained gold. In Example 12, a leaching solution containing 1.0 gm copper ions, 0.1 gm calcium lignosulphonate, and 120 gm of a 25% NH3 solution was diluted to 1.6 litre. 100 gm of finely ground (p80,5 micron) gold bearing sulphide concentrate was added and the solution agitated whilst oxygen gas was introduced into the mixing zone at a rate of 20 L/hr. The mixture was pre-heated to 60°C, and further heating was discontinued. The reaction was continued for 3.5 hours and the slurry was filtered and washed. The residue was leached with a solution containing 1.0 gm sodium cyanide at pH > 12 to sodium hydroxide.

In Comparative Example 4,50 gm of the same sample of finely ground (p80,5 micron) gold bearing sulphide concentrate was dispersed and leached with a solution containing 1.0 gm sodium cyanide at pH > 12 to sodium hydroxide at room temperature. No pre-treatment was undertaken to the concentrate sample. The slurry was filtered and washed and the residue was subject to a second leach under identical conditions to ensure maximum gold recovery. The recovery of gold and silver from the respective cyanide leaches is shown in Figure 11.

Further, it is to be understood that various alterations, modifications, and/or additions may be introduced into the constructions and arrangements of parts previously described without departing from the spirit or ambit of the invention.