GRAVITY CONCENTRATES Various processes are used in the treatment of gravity concentrates. These processes involve either discrete hydrometallurgical or pyrometallurgical steps or combinations thereof. Selection of a particular process is dictated by the nature of the occurrence of gold in the concentrate, scale of the operation and level of available technical expertise.

The processes that are commonly used are direct smelting, amalgamation, intensive cyanidation, and sulphide roasting to produce a calcine followed by cyanidation. Each of these processes has distinct disadvantages.

Direct smelting is a pyrometallurgical process, which requires a high-grade concentrate containing from 15% to 50% Au. The process produces a low-grade bullion of from 50% to 90% Au and a slag which contains small amounts of Au. This

slag usually requires further treatment to recover the Au.

The main disadvantage of the amalgamation process is that it uses extremely toxic mercury to produce gold mercury alloy or amalgam. Legislation has been passed in many countries which prevents the use of mercury in gold processing.

Intensive cyanidation is a hydrometallurgical process, which by implication requires significantly higher levels of cyanide than that normally used. The process does not yield good recoveries if the gold concentrate is of a refractory nature, eg. a gold sulphide complex.

Sulphide roasting involves a pyrometallurgical step to oxidise the sulphides to gaseous sulphur dioxide. The residue, referred to as calcine, containsthe gold, which is in a form that can be leached in a normal cyanide leach. The roasting process is more suited to higher sulphide content concentrates, while the sulphur dioxide has to be converted into sulphuric acid. Sulphide roasting and associated gas recovery are complex metallurgical operations.

GOLD FEEDS OTHER THAN GRAVITY CONCENTRATES Gold feeds other than gravity concentrates typically include electrolytic sludges, zinc- gold precipitates and general gold scrap.

A pyrometallurgical smelting process is the conventional method for processing feeds of this type ie. excluding gravity concentrates. Depending on the type of the feed material, pre-treatment steps may be required ahead of the main smelting process

step. Smelting is essentially a gold upgrading step, which removes most of the unwanted impurities in a slag to produce a gold bullion containing from 50% to 90 % Au, the remainder being Ag and small levels of base metals. The slag contains finely disseminated gold particles, typically 500-2000 g/t, and therefore requires further treatment to recover the gold.

Feed derived from electrolytic sludges originates from two electrowinning technologies namely, electrowinning onto mild steel wool cathodes or onto stainless steel wool cathodes. Removal of the gold from mild steel wool cathodes requires calcination prior to smelting, or dissolution of the steel wool in an acid, such as sulphuric acid, hydrochloric acid or nitric acid, to yield an insoluble gold sludge, which after filtration and drying is smelted. The leach solution is discarded.

The gold deposited onto stainless steel wool is removed by washing with water, which after filtration and drying is smelted. Unlike the mild steel wool, the stainless steel wool cathode is reused.

Feed derived from older gold operations is typically in the form of a gold-zinc precipitate. The precipitate typically contains a gold content of from 15% to 35%, the balance being predominantly zinc and silica. Such precipitates require two pre- treatment steps ahead of smelting. The first step is optional and involves an acid treatment to dissolve most of the excess zinc. After filtration and drying, the upgraded zinc-gold precipitate is heat treated or calcined to oxidise the remaining zinc and other base metals. This calcine is then fluxed and smelted to produce gold bullion.

Feed derived from scrap gold is smelted directly to produce bullion.

Clearly, current technology for producing primary gold bullion involves numerous steps, with each step generating an effluent or residue. The disadvantages of the present technology are summarised as follows : 'high labour requirements, which translates into a higher security risk for gold theft; a slag residue is generated which requires further treatment to recover entrained gold ; the removal of residues from the smelthouse poses an additional security risk for gold theft; and the smelthouse product is gold bullion of varying gold content, which requires further refining.

Once the gold bullion is produced, it is transported from site to a central refinery where the gold is refined by either a combination of a pyrometallurgical and electrochemical process or a hydrometallurgical process. During the refining process silver is separated and refined, while two grades of gold are typically produced. The one grade has a purity of 99.5% and is used for casting monetary bars, while the other grade has a purity greater than 99.99% and is used for jewellery, coins and other industrial applications. The entire process from primary gold bullion to monetary gold product can take a few days.

The disadvantages of the refining step, after gold bullion production, are: transportation of gold bullion from the primary to the refinery poses a security risk; 'additional refining costs are incurred at the gold bullion supplier's expense; and

the additional time spent in refining the gold bullion translates into loss in revenue for the bullion supplier.

SUMMARY OF INVENTION The invention provides a process for the recovery of gold from gold-bearing concentrate which include the steps of: (a) leaching the concentrate in HCI solution with gaseous chlorine to dissolve at least the gold and base metals in the concentrate, (b) subjecting a slurry produced by step (a) to a solid/liquid separation step to produce a gold bearing solution, and (c) precipitating gold from the solution.

Step (b) may be effected by means of filtration. Solids, separated from the slurry, may be repulped and subjected to a silver recovery process.

Step (c) may be carried out by treating the solution with a suitable reductant to precipitate the gold as a powder.

The reductant may be of any appropriate kind for example gaseous S02, sodium metabisulphite or copper and aluminium powder. Preferably, however, use is made of ferrous sulphate.

If the feed material has a relatively high gold content, say in excess of 10% by weight, and for example is in the form of electrolytic sludge, a zinc-gold precipitate or general gold scrap, then the chlorine used in step (a) may be derived from bottled gas or may

be generated on site.

On the other hand if the feed material is a gold bearing gravity concentrate, typically containing in excess of 1% gold (by weight), then it is preferred for economic reasons to make use of chlorine gas which is generated at site.

BRIEF DESCRIPTION OF THE DRAWINGS The invention is further described by way of examples with reference to the accompanying drawings in which: Figure 1 is a flow chart of a hydrometallurgical gold recovery process applied to a feed material which has a relatively high gold content such as an electrolytic sludge, a zinc- gold precipitate and general gold scrap, Figure 2 is in many respects similar to Figure 1 and illustrates steps in the process of recovering gold from a gold bearing gravity concentrate which has a relatively low gold content, and Figure 3 illustrates a chlorine generation cell which is used in the process of Figure 2.

DESCRIPTION OF PREFERRED EMBODIMENTS Figure 1 illustrates the process of the invention used for the recovery of gold from a feed material 10 which has a relatively high gold content such as an electrolytic sludge, a zinc-gold precipitate or general gold scrap.

LEACHING Referring to Figure 1, the feed material 10 with a relatively high gold content is fed into a leach tank 12. Concentrated hydrochloric acid 14 and water 16 are added to the leach tank to produce a leaching solution with a hydrochloric acid concentration of 5 to 6 M. Mixing of the leach slurry is induced by re-circulating the leach from the upper section of the leach tank through the bottom of the tank. The mixing improves the leaching kinetics by partially fluidising the feed material. Mechanical agitation may also be used to produce the desired mixing.

Chlorine gas 18 from bottles is continuously sparged into the bottom of the tank 12.

The pressure in the leach tank is maintained at as close as possible to atmospheric pressure by a pressure regulating system. Excess chlorine 20 is absorbed in a chemical scrubber 22. As the leach progresses, the gold and base metals are dissolved resulting in a dilute slurry 24 of insoluble silver chloride and small amounts of silica. The silica in the residue originates from feed material consisting of gold-zinc precipitates.

The leach is complete when there is no further demand for chlorine. Air 26 is then sparged into the leach tank to remove entrained chlorine gas from the slurry. The leach slurry 24 is discharged from the leach tank, and filtered (step 28). The residue is washed with water 30 and then discharged from the filter for further processing 32 to produce high purity metallic silver 34. Production of high purity silver is not part of this process. The filtrate 36, containing dissolved gold and base metals, is transferred to a gold precipitation tank 38 for further treatment.

By way of example only, leaching parameters are as follows : Leaching time 2-6 hours Operating temperature 25°C-40°C Au leaching efficiency 99%-100% HCI concentration 3-10 M PRECIPITATION The first step during precipitation is the removal of any residual chlorine from the leach filtrate by sparging with air 40. The resulting air/chlorine mixture 42 is vented to the chemical scrubber 22. A suitable reductant 44, such as gaseous SO2, sodium metabisulphite, or copper and aluminium powder, but preferably ferrous sulphate, is added into the precipitation tank 38 to reduce the aqueous gold to metallic powder.

Very little silver and none of the base metals are precipitated with the gold. The precipitation reaction is enhanced by gentle mechanical agitation. The gold-barren solution and gold precipitate 46 is filtered (step 48) to separate the precipitate, which after washing with water 50 is dried to form a powder 52. The filtrate 54 is transferred into a neutralisation tank 56 for further treatment.

The gold powder 52, which has a purity of approximately 99.9%, can be melted to produce monetary gold bars on site, or may undergo further refining to produce a purity of 99.99%. The melting and additional refining steps do not form part of the process of the invention.

By way of example oniv. precipitation parameters are as follows : Precipitation time 1-4 hours Final Au in solution < 2 mg/L Final redox potential 480-550mV Reductant ferrous sulphate EFFLUENT TREATMENT All gases (20 and 42) extracted from tanks and filters are scrubbed through the chemical scrubber 22 before venting to atmosphere. The scrubber effluent 58 is pumped into the neutralisation tank 56 where it is mixed with the acidicfiltrate 54. The excess acid is neutralised with lime 60 and the pH is raised to pH 10-11 before disposal. The contents of the tank are agitated by sparging with air 62 and water64 is added, as necessary. Failure to neutralise the effluent 54 adequately may result in the generation of hydrogen cyanide when contact is made with a cyanide-bearing stream.

The resulting slurry 66 is fed to an existing gold residue treatment plant.

ADVANTAGES OF PROCESS The replacement of the conventional primary gold smelting and subsequent refining steps by a simple on-site hydrometallurgical process to recover a high purity gold powder, in the manner described, has a number of advantages with include the following: the process is not labour intensive and therefore poses a lower security risk with respect to gold losses due to less handling;

the process operates on a batch basis, which allows for accurate gold accounting; no gold inventory remains in the plant after processing each batch; the gold product is of monetary purity and is recovered in less than 24 hours; the process is cost effective due to the reduced number of processing steps; the nature of the leach and precipitation steps allows for the simultaneous treatment of a larger variety of gold containing feed materials, especially secondary sources of gold, than was previously possible; the process is easily automated; the process produces two discrete products, namely, a high purity gold powder and clean silver chloride solids ; the gold precipitate, with an assay of 99.9%, can be melted directly into monetary bars or may undergo further refining to produce 99.99% gold ; . the clean silver chloride residue, which contains no gold, can be treated conventionally to produce high purity silver metal ; gold losses are insignificant due to efficient precipitation and filtration; 'there is no generation of gold-bearing residues; the only effluent from the process is a gold barren acidic solution, which is treated to allow safe disposal into a cyanide bearing stream with no generation of hydrogen cyanide; and effluents from the process do not pose additional environmental risks.

Figure 2 illustrates steps in a process for the recovery of gold from gold concentrates ie. under conditions in which the gold content is low but in excess of 1 %. In Figure 2 many steps are the same as what has been described in connection with Figure 1

and, for this reason, such steps bear the same reference numerals as the corresponding steps in Figure 1.

An important feature of the process of Figure 2 however is the use of chlorine which is generated on site. Since the chlorine is generated for immediate use at ambient pressure there is no need for expensive chlorine storage or delivery systems. A safeguard is incorporated into the process to ensure that the chloride effluent can be safely discharged into a cyanide circuit.

Figure 3 illustrates a chlorine generation cell used for the generation of on-site chlorine, in the implementation of the method of Figure 2.

LEACHING Referring to Figure 2, a gold gravity concentrate 10 with a grade >1% is fed into a leach tank 12, with or without a draft tube. Concentrated hydrochloric acid 14 and water 16 are supplied to the leach tank to create a leaching medium with a hydrochloric content of about 5 to 6 M. The solids in the tank are partially fluidised by circulating a dilute slurry mixture from the upper section of the leach tank and re- circulating this through the bottom of the tank.

Another stream is removed from the upper level of the leach tank and is circulated through a pipe reactor 70 or a similar item of equipment and returned to the leach tank. The chlorine 18 is generated by a chlorine cell 72 which is started at the beginning of the leach and the generated chlorine gas 18 is continuously induced into the pipe reactor 70. The chlorine gas dissolves in the dilute slurry in the pipe reactor

70 to form aqueous chlorine before entering the leach tank 12. Leaching is preferably carried out at ambient temperature and pressure. Excess chlorine 20, vented from the leach tank, is absorbed in a chemical scrubber 22.

At the end of the leach, the chlorine cell 72 is shut down to stop the generation of chlorine gas 18. Air 26 is then sparged into the leach tank to remove entrained chlorine gas from the slurry. The slurry 24, which contains insoluble silver chloride and residual silica, is discharged from the leach tank and filtered (step 28). The residue is washed with water 30 as required and is then re-pulped in a repulper tank 74 with water 32 before being transferred to a milling circuit 76 in the main plant, where the silver chloride and any residual gold will be recovered in the cyanidation circuit. The filtrate 36, containing dissolved base and precious metals, is transferred to the gold precipitation tank 38 for further treatment.

By way of example only, leaching parameters are as follows : Leaching time 12-18 hours Operating temperature 25°C-50°C Au leaching efficiency > 98% HCI concentration 3-10 M CHLORINE CELL AND CHLORINE GENERATION Figure 3 schematically shows the chlorine generating cell 72. The cell is a simple low- pressure chlorine generator and has a body which is fabricated from plastic and consists of a number of anode and cathode compartments 80 fastened by tie rods 82

which extend between end and pressure plates 84 and 86 to form a sealed unit.

Recessed seals between each compartment ensure that no leakage occurs from the cell structure. A permeable polypropylene diaphragm or ion exchange membrane separates the anolyte and catholyte solutions. Because of the bi-polar design, i. e there are only two electrical contacts 88 and 90. Each electrode, which is fabricated from titanium, acts as an anode and a cathode. The anode side of the electrode is coated with Ru02 to prevent passivation, while the cathode side is uncoated.

At the anode, brine solution is electrolysed to form the chlorine gas 18, while at the cathode, water dissociates to form hydroxyl ions and hydrogen gas 94. The sodium ions in the anode compartment pass through a permeable diaphragm and combine with the hydroxyl ions to form sodium hydroxide-caustic soda 96.

A concentrated brine solution is pumped (91) from a brine storage tank 98 into the anode compartment, via an inlet 100, where it is electrolysed to form chlorine gas. The anolyte, which consists of a mixture of brine and wet chlorine gas, overflows from the top of the anode compartment into a cavity where the gas separates from the brine solution and exits via outlets 102 and 104. The chlorine gas 18 is removed from this cavity by suction. The partially depleted brine solution is pumped (92) back to the brine storage tank 98 where it is re-saturated with salt. Re-saturation is achieved by circulating (106) the depleted brine solution through a saturation tank 108 containing coarse salt 110.

On the cathode side, a caustic solution 112 is pumped from a caustic storage tank 114 into the cathode compartment, via another inlet in the bottom of the cell (opposite the

brine inlet). The catholyte, which contains a higher concentration of caustic soda and hydrogen gas, overflows from the top of the cathode compartment into a cavity where the gas separates from the caustic solution and exits via outlets 116 and 118. The hydrogen gas 94 is vented from this cavity and after dilution with air 120, to produce a gas mixture well below the explosive limit for hydrogen in air, is exhausted to atmosphere. The more concentrated caustic solution 96 is pumped back to the caustic tank 114 where the concentration is controlled by dilution with water 122. Excess caustic solution 124 is bled continuously from the caustic tank to a dechlorination tank 138. Some of this bleed 128 is used to replenish caustic in the gas scrubber system (22).

The anolyte operates at a slightly higher level than the catholyte to ensure a small flow of brine solution through the permeable diaphragm into the catholyte solution.

This flow of brine is necessary to minimise back-diffusion of sodium hydroxide into the brine and consequent formation of undesirable hypochlorite. This flow of brine, designated 130, is shown figuratively in Figure 2. A baffle, at the top of the anode compartment, prevents the mixture of hydrogen and chlorine gases. Furthermore, atmospheric vents in the cell prevent the build-up of pressure in the cell. Since the chlorine 18 is removed from the cell by suction, no gas leaks occur to atmosphere. No brine bleed would occur if an ion exchange membrane were used.

PRECIPITATION A first step in the precipitation process is the removal of any remaining chlorine from the leach filtrate by sparging with air 40. This is followed by the addition of a suitable

reductant 44, such as gaseous S02, sodium metabisulphite, a copper and aluminium powder, but preferably use is made of ferrous sulphate, which reduces the aqueous gold to metallic powder. Very little silver and none of the base metals are precipitated with the gold. The precipitation reaction is enhanced by gentle mechanical agitation.

The gold-barren solution and gold precipitate 46 is filtered (step 48) to separate the precipitate, which after washing (50) is dried to form a powder 52. The filtrate 54 is transferred into the neutralisation tank 56 for further treatment.

The gold powder 52, which has a purity of approximately 99.9%, can be melted to produce monetary gold bars on site, or may undergo further refining to produce a purity of 99.99%. The melting and additional refining steps do not form part of the process of the invention.

By way of example only precipitation parameters are as follows : Precipitation time 1-2 hours Final Au in solution < 2 mg/L Final redox potential 480-550mV Reductant ferrous sulphate EFFLUENT TREATMENT All gases, extracted from tanks and filters, ie. the chlorine gas 20, the gas 136 produced by the repulper tank 74, and the chlorine gas 42 from the tank 38, are scrubbed through the caustic soda scrubber 22 before venting to atmosphere.

The scrubber effluent 136, brine bleed 130, and caustic bleed 124 contain some

sodium hypochlorite, while the leach/precipitation filtrate 54 contains hydrochloric acid and sulphuric acid.

The effluent treatment circuit consists of a dechlorination tank 138 and the neutralisation tank 56 in which dechlorination and neutralisation are carried out respectively. The scrubber effluent 128, bleed 130 and caustic bleed 124 are transferred into the dechlorination tank 138, while the acidic gold-barren leach filtrate 54 is transferred directly into the neutralisation tank 56. A solution of sodium thiosulphate 140 is added to the dechlorination tank to destroy hypochlorite. After dechlorination, the alkaline solution 142 is transferred to the neutralisation tank 56, where it neutralises some of the acid in the leach/precipitation filtrate. The pH in the neutralisation tank is then adjusted with slaked lime 60 to a value of 10 to 11.

Failure to dechlorinate or neutralise the effluent may result in the generation of cyanogen chloride or hydrogen cyanide when contact is made with a cyanide-bearing stream.

As is the case with the Figure 1 process water 64 and air 62 are added to the tank 56 to produce a slurry 66 which is fed to an existing gold residue treatment plant.

ADVANTAGES OF PROCESS The generation of chlorine gas in the chlorine cell 72 using brine as a chlorine source and the use of the gas for the recovery of gold, in the manner described, has a number of advantages with include the following :

the process is not labour intensive and therefore poses a lower security risk with respect to gold losses due to less handling; the process is carried out on a batch basis which lends itself to effective gold accounting; gold recovery takes place in less than 24 hours and no gold inventory remains in the plant ; efficient dissolution of any gold complex and especially refractory gold complexes is achieved; safe and efficient utilisation of low-pressure chlorine gas is effected; the process is easily automated; there is no need for storage of hazardous chlorine; the process results in the recovery of a clean gold precipitate which can be further refined without any pretreatment steps; silver is separated from gold powder as it reports to the final residue as silver chloride ; gold losses are insignificant due to efficient precipitation and filtration ; no bulk storage of hazardous chemicals is required as the chlorine cell produces chlorine gas on demand; effluent is safely disposed of into a cyanide bearing stream with no generation of cyanogen chloride or hydrogen cyanide; and the effluents from the process do not pose additional environmental risks.