US20120067170A1

US20120067170A1 - Extraction of gold from cathode associated gold concentrates

Info

Publication number: US20120067170A1
Application number: US13/266,001
Authority: US
Inventors: Dean R. Butler
Original assignee: Precious Metals Recovery Pty Ltd
Current assignee: Precious Metals Recovery Pty Ltd
Priority date: 2009-04-24
Filing date: 2010-04-23
Publication date: 2012-03-22
Also published as: US20120090433A1; AU2010239156A1; WO2010121317A1; CN102459660A; CA2760354A1; CN102459659A; WO2010121318A1; CA2759804A1; AU2010239083A1; CA2760345A1; US20120067169A1; CN102459661A; WO2010121329A1; ZA201107782B; ZA201107781B; AU2010239157A1; ZA201107780B

Abstract

The invention involves a method for recovering gold from a gold concentrate comprising: dissolving gold from the concentrate in an aqueous liquor to provide a gold liquor; subjecting the gold liquor to electrolysis in an electrowinning cell to provide cathode-associated gold material; leaching the cathode associated gold material in an aqueous liquor under reducing conditions to provide a treated solid residue; and smelting the treated solid residue to recover gold.

Description

FIELD

The invention relates to a method and system for extraction of gold from gold concentrates involving dissolving the gold in an aqueous liquor, electrowinning to provide cathode associated gold concentrates and processing the concentrates.

BACKGROUND

In many gold processing operations a sequence of unit operations is used that lead to the dissolution of gold in a liquor. The components of the liquor are substantially soluble, i.e. the gold values travel in a soluble liquor stream. Furthermore, insoluble moieties associated with the source materials are present in only very low concentration. Thus, for example, gold from an ore may be taken up in a cyanide liquor (the first soluble liquor stream), absorbed onto carbon and then taken up (or stripped) from carbon into a second more concentrated soluble gold liquor stream (the strip liquor). Alternatively gold from a gravity gold concentrate may be taken up directly into a concentrated soluble gold liquor stream.
The concentrated soluble gold liquor stream is typically fed to an electrolysis cell, where the action of electric current leads to the deposition of gold on or under the cathode. This deposited gold is herein referred to as cathode-associated gold concentrate.
Cathode-associated gold in general does not contain valuable amounts of platinum group metals (PGMs)—this may be because little of these materials are present in the source material, or because some of these materials are poorly soluble and do not travel in soluble liquor streams (typically cyanide-based) that are used to solubilise gold values. Cathode-associated gold concentrates are usually further processed by smelting, wherein flux is added to the cathode-associated gold concentrate and the mixture is heated at a temperature sufficient to melt flux and gold. Base metal and base metal associated contaminants present in the cathode-associated gold concentrate pass into the molten flux to form a slag phase, which after cooling can be physically separated from the precious metal (bullion) phase. A characteristic of the smelting operation is that the precious metals are consolidated in a molten button or bar which has a significantly lower surface area than the surface area of the cathode-associated concentrate.
Alternative classes of gold recovery operations are used in the recovery of gold as a by product in the mining of base metals such as copper and nickel. When gold in the source material is found in small quantities in association with valuable quantities of copper or other base metals, impure base metal is cast (by melting) to form an anode for use in an electrolysis cell, and the action of current in this cell leads to the dissolution of copper (with deposition of purified copper on the cathode). Un-dissolved residues from the anode (including any trace amounts of gold) accumulate under the anode, and comprise the anode mud. Components of anode mud are insoluble (in the acidic electrolysis liquor) and do not travel in a soluble liquor stream. These insoluble components may frequently include valuable amounts of platinum group metals (PGMs) as well as gold and silver. The contaminants in an anode-associated gold concentrate are quite different from contaminants in a cathode-associated gold concentrate because (a) the source material is different; (b) many processing steps are different, and in particular contaminants in anode-associated concentrate have not travelled in a soluble liquor stream while contaminants in a cathode-associated concentrate have generally travelled in a soluble liquor stream on two separate occasions; and (c) anode-associated gold concentrates have, in steps prior to electrolysis, been part of a molten-metal environment (for example when the anode is cast) whereas cathode-associated gold concentrates have never, prior to electrolysis, been part of a molten-metal environment. Anode mud concentrates are typically processed using a succession of specialised leaching steps to recover various valuable components in various leach liquors. The smelting of anode muds comprising valuable levels of PGMs as well as gold is counter-productive because smelting leads to a consolidation of gold, silver and PGMs in a molten mass with a significantly reduced surface area (relative to anode mud), and the selective leaching of valuable components from this low surface area molten mass is much more difficult than selective leaching of the original high surface area anode mud.
The current methods of refining cathode associated gold from electrolysis of aqueous gold liquor generally involve smelting of the cathode associated gold without leaching prior to heat treatment.
The discussion of documents, acts, materials, devices, articles and the like is included in this specification solely for the purpose of providing a context for the present invention. It is not suggested or represented that any or all of these matters formed part of the prior art base or were common general knowledge in the field relevant to the present invention as it existed before the priority date of each claim of this application.

SUMMARY

The inventor has found that the recovery of gold from cathode associated gold formed in the electrowinning process can be enhanced by leaching of cathode associated gold with an aqueous reducing liquor.
Accordingly, there is provided a method for recovering gold from a gold concentrate comprising:
dissolving gold from the concentrate in an aqueous liquor;
subjecting the gold liquor to electrolysis in an electrowinning cell to provide cathode-associated gold material;
leaching the cathode associated gold material in an aqueous liquor under reducing conditions to provide a treated solid residue; and
smelting the treated solid residue to recover gold.
In one set of embodiments the gold containing liquor is derived from stripping of gold from carbon on which the gold is sorbed. In another embodiment the liquor is derived from solubilisation of gold from gravity concentrates.
Accordingly in one set of embodiments there is provided a method described above wherein the gold concentrate comprises carbon sorbed with gold and the method further comprises stripping gold from the carbon by contact with an aqueous liquor to provide said gold liquor and in another set of embodiments there is provided a method as described above wherein the gold concentrate comprises a gravity concentrate and the method further comprises dissolving gold from the gravity concentrate to provide said gold liquor.
Throughout the description and the claims of this specification the word “comprise” and variations of the word, such as “comprising” and “comprises” is not intended to exclude other additives, components, integers or steps.

DETAILED DESCRIPTION

In general we have found that leaching of cathode associated gold with an aqueous reducing liquor provides an improved recovery of gold often providing of the order of several percent or more increase in gold yield. An increase of several percent in yield of gold is extremely significant particularly for large gold mining operations.
The gold concentrate may be a carbon sorbent sorbed with gold or may be a gravity gold concentrate. These materials are known in the art and produced in well known mining operations. The concentration of gold in the concentrate is typically at least 0.1% by weight. The concentration of gold in the gold containing liquor that is fed to the electrolysis cell is typically 20 to 2000 ppm or greater.
Aqueous liquor for dissolving gold is known in the industry and preferably is a cyanide liquor such as sodium cyanide or potassium cyanide. The chemical reaction for dissolution of gold by cyanide is called the Elsner Equation and in the case of sodium cyanide is as follows:
4Au+8NaCN+O₂+2H₂O→4Na[Au(CN)₂]+4NaOH
Cathode associated gold includes gold deposited on the cathode or which is formed adjacent the cathode and may for example collect in the electrowinning cell below the cathode.
Gravity gold is gold concentrated by a gravity process. Gravity concentration has been historically the most important way of extracting the native metal using pans or washing tables. However, froth flotation processes may also be used to concentrate the gold. In some cases, particularly when the gold is present in the ore as discrete coarse particles, a gravity concentrate can be in some cases be directly smelted to form gold bars. In other cases, particularly when the gold is present in the ore as fine particles or is not sufficiently liberated from the host rock, the concentrates are treated by cyanidation leaching, followed by recovery from the leach solution. Recovery from solution may involve adsorption on activated carbon and/or electrolysis to form cathode associated deposits.
Throughout the description and the claims of this specification the word “comprise” and variations of the word, such as “comprising” and “comprises” is not intended to exclude other additives, components, integers or steps.
The reducing liquor may be provided by a reducing agent, by contact with a reducing electrode, or a combination thereof.
The reducing agent is preferably compatible with aqueous liquor and may be metal containing or non-metal containing. Examples of suitable metal containing reducing agents include metal containing moieties in a valence state lower then the maximum stable valence state achievable in an aqueous solution. The more preferred metals may be chosen from the group consisting of chromium (Cr II), tin (Sn II), copper (Cu I) and titanium (Ti II, Ti III), most preferably tin (Sn II). In a preferred embodiment, the aqueous reducing liquor comprises stannous ion, for example stannous chloride.
In one preferred set of embodiments the aqueous reducing liquor comprises stannous ion (more preferably in the form of stannous chloride). Without being bound by theory it is believed likely that the use of a reducing leach may facilitate the dissolution of moieties comprising Iron (III), and that these moieties are responsible or partially responsible for immobilizing gold. Evidence for the dissolution of moieties comprising iron III includes decoloration of material after leaching. Leaching may be carried out in liquors comprising 1% HCl and one or more reducing agents such as tin (II) chloride, chromium (II) chloride and oxalic acid. Based on the observed degree of decoloration the effectiveness of reducing agents decreases according to the ranking tin (II) chloride, ≧chromium (II) chloride>oxalic acid
Examples of suitable non-metal containing reducing agents may be selected from the group consisting of sulfites, oxalic acid, formic acid, hydrazine, acetates including acetic acid, citrates including citric acid sulfite and dithionite and preferably sulfites and other organic acids. Organic acids are particularly suitable.
The leaching of the cathode associated gold concentrate under reducing conditions may use an aqueous leach composition selected from the group consisting of aqueous base, aqueous acid, aqueous chelating agent and mixtures of aqueous chelating agent with acid or base. For example, the aqueous reducing liquor may be an acidic aqueous reducing liquor. Alternatively, the aqueous reducing liquor is an alkaline aqueous reducing liquor.
Preferably the reducing liquor in at least one contact between with source material is acidic, preferably the pH is less than about 1.5, more preferably less than about 1.0. Preferably the acid is a non-oxidising acid. Preferably the acid is hydrochloric acid.
In a preferred embodiment the reducing agent is a regenerable reducing agent, for example, a reducing agent which can be regenerated from the oxidised form produced as a result of the process by electrolytic regeneration of the reducing agent.
The method may further comprise leaching the deposited gold concentrate, prior to said reducing leach step, in an aqueous liquor comprising one or more agents selected from the group consisting of hydrochloric acid, nitric acid, alkali, chelating agents, carboxylic acids and their salts, chlorates, perchlorates, chlorides, fluorosilicates, phenol sulfate and peroxydisulfate.
In one set of embodiments the method further comprises subjecting the solid residue from leaching in aqueous reducing liquor to at least one leaching step in an aqueous liquors comprising agents selected from the group consisting of hydrochloric acid, nitric acid, alkali, lead acetate, chelating agents, carboxylic acids and their salts, chlorates, perchlorates, chlorides, fluorosilicates, phenol sulfate and peroxydisulfate.
Examples of carboxylic acid which may be used in leaching prior to, with the reducing leach or after the reducing leach include formic acid, acetic acid, lactic acid, citric acid, isobutyric acid and salts thereof such as the alkali metal and alkaline earth metal salts. Examples of chlorides which may be used in leaching prior to, with the reducing leach or after the reducing leach include ammonium chloride, sodium chloride, potassium chloride, calcium chloride and strontium chloride.
The method may comprise a plurality of steps involving contact of the cathode associated gold material with an aqueous reducing liquor.
Preferably the contact between the source material and the aqueous reducing liquor is carried out at a negative first Eh, and a subsequent contact between the source material and an aqueous reducing liquor is carried out at a more negative second Eh.
Preferably the Eh remains negative throughout the contact period between aqueous reducing liquor and the source material and residue derived therefrom.
In a preferred embodiment of the invention the contact between the source material or residue and aqueous liquor is carried out in conditions that encourage the dislodgement of refractory material from the surface of the solid. Such conditions may include ultrasonic agitation.
In one set of embodiments the leaching is conducted at a temperature of at least 60° C.
Preferably the process of the invention leads to the recovery of a greater quantity of gold from the source material than is apparent in a standard bullion assay test of the source material. Preferably the excess gold recovery over bullion assay grade is at least 1%, preferably at least 2%, preferably at least 5%.
The process preferably comprises removing the liquor from the source material which takes place after contacting the source material with the aqueous reducing liquor. A wide range of methods and apparatus are known in the industry for solid-liquid separation. For example, the liquor may be percolated through the source material in a batch tank and collected as run off, the source material may be filtered from a slurry using suitable filtration equipment known in the minerals processing industry or alternatively the source material solids may submit to gravity separation from liquor, for example, in suitable batch or continuous settling tanks known in the industry. In one embodiment the step of contacting the source material with an aqueous liquor under reducing conditions is carried out by agitating (e.g. stirring, swirling or otherwise agitating) an aqueous slurry of the source material with a reducing agent and the aqueous slurry liquor is removed from the source material by filtration. Other methods such as centrifugal separation may be used if desired but may be less practical on an industrial scale. Such methods may, however, be suitable in use of the process for assay of precious metals.
The process of gold recovery frequently involves a leaching step and adsorption of gold and other precious metals onto an adsorbent such as carbon or a suitable synthetic resin. Improvements in the adsorption process such as the carbon in column (CIC), carbon in leach (CIL) and carbon in pulp (CIP) processes have led to efficient gold recovery which in some cases have even justified reprocessing of mine tailings. Precious metals are stripped from the adsorbent by elution using suitable liquor comprising lixiviant and oxidant.
In one embodiment the gold-rich source material is preferably cathodic material or cathodic sludge from a electro-winning of a strip liquor such as may be used to remove precious metals from sorption onto carbon. Preferably said cathodic material has been treated to remove steel wool.
In another embodiment the gold-rich source material comprises gold sorbed onto carbon.
In a preferred embodiment the reducing liquor in at least one contact with source material comprises at least one chelating agent, preferably selected from the group consisting of beta-diketones, amino polycarboxylic acids, salts of amino polycarboxylic acids, carboxylic acids, salts of carboxylic acids, and polyphosphonates.
In a preferred embodiment the material subject to leaching in an aqueous reducing liquor is finely divided.
For example, the material subject to leaching in aqueous liquor may be finely divided providing on wet sieving at least 80% by weight passing through a 100 micron sieve and more preferably at least 80% by weight passing through a 30 micron sieve.
In one embodiment of the invention the method optionally comprises treatment either prior to leaching in aqueous reducing liquor, after leaching in aqueous reducing liquor or both before and after leaching in aqueous reducing liquor. The optional treatment may include at least one leaching step in an aqueous liquor comprising agents selected from the group consisting of hydrochloric acid, nitric acid, alkali, lead acetate, ammonium chloride, calcium chloride, strontium chloride, acetic acid and citric acid. The optional treatment step preferably comprises leaching in an aqueous alkali metal hydroxide, an aqueous nitric acid, aqueous hydrochloric acid or mixtures of the acids. The optional treatment step may be conducted at elevated temperature such as at from 40° C. to 90° C. and more preferably from 60° C. to 80° C. It may be advantageous in this embodiment to conduct the optional treatment with application of agitation such as ultrasonic agitation. The optional treatment may if desired comprise a plurality of leaches using the same of different aqueous leach liquors before and/or after the reducing leach.
In one embodiment, the solid residue from the aqueous reducing liquor is treated with an aqueous alkaline liquor which has a pH greater than 13, more preferably greater than 14. In one embodiment, the alkaline liquor comprises at least 5% sodium hydroxide.
In a preferred embodiment at least one step selected from contact with the aqueous reducing liquor and leaching prior or after said contact with the aqueous reducing liquor is carried out in conditions that encourage the dislodgement of refractory material from the surface of the source material or solid residue. Examples of such conditions include ultrasonic agitation.
The use of ultrasonic agitation is preferred and in particular a frequency in the range 10-60 kHz, more preferably 20-45 kHz is preferred. In one set of embodiments ultrasonics are applied to a hot leach liquor, for example at a temperature of at least 60° C.
In one preferred set of embodiments the method comprises said reducing leach and a subsequent acid leach wherein the liquor used in the reducing leach comprises hydrochloric acid (preferably 0.5 to 5M hydrochloric acid) and stannous chloride (preferably from 5 to 150 grams per litre of stannous chloride dihydrate, more preferably 10 to 100 and even more preferably 30 to 50 grams per litre of stannous chloride dihydrate) and the liquor used in the subsequent acid leach comprises nitric acid (preferably about 5 to about 70%, more preferably 20% to 60% and even more preferably about 50% nitric acid) in water. The weight ratio of liquid to solid material in the subsequent acid leach step is preferably in the range of 10:1 to 100:1 (preferably 20:1 to 50:1, more preferably about 40:1).
The method may comprise treatment prior to or after leaching with the aqueous reducing liquor. The solid residue which has been treated in accordance with the invention may be further refined to provide precious metal by methods known in the art such as by smelting. In one preferred embodiment of the invention the refining step comprises adding said residue to a crucible, and heating the contents to smelting temperature. The smelting process may include the use of a flux, such as a flux comprising borax, or may be flux-less.
In a preferred embodiment, as an initial step the flux is placed in the crucible and melted prior to addition of the treated solid residue, or the solid residue plus metal. In a further preferred embodiment there is no flux mixed with the treated solid residue or solid residue mixed plus metal.
During the process of smelting gold the present inventor has found that a significant amount of gold, frequently of the order of from 1 to 3% or even more, is lost to slag. Even when the slag is ground and reintroduced into an earlier part of the gold recovery circuit this slag associated gold may be substantially unrecoverable.
In one set of embodiments of the heating step the cathode associated gold concentrate is added to a previously melted pool of metal comprising a material comprising a metal selected from copper, silver, gold and platinum group metals. In a preferred set of embodiments the said material has a concentration of at least 80% by weight (preferably at least 90% and more preferably at least 95% and still more preferably at least 99% by weight) of one of copper, silver, gold and platinum group metals.
In one preference the molten pool has a melting point in excess of 900° C. Preferably the pool metal has one metal selected from the group consisting of gold, silver and copper. In one set of preferred embodiments the metal components are placed in proximity to the treated solid residue and the melting step causes the metal components (preferably selected from gold, silver, copper) to melt.
The smelting may, in one set of embodiments, comprise:
smelting the reduced solid residue by forming a molten pool comprising at least one metal selected from copper, silver, gold and platinum group metals; and
adding the reduced solid residue into the pool of molten metal.
In a particularly preferred set of embodiments the molten pool is formed from a solid particulate mixture comprising particles of treated solid residue and particles of at least one metal selected from copper, silver, gold and platinum group metals. In this set of embodiments the pool metal preferably comprises silver, copper or mixtures thereof. The particulate mixture of the residue and at least one metal selected from copper, silver and gold is preferably gradually added to a heated crucible such that a molten pool is formed during addition and further particulate mixture is added to and becomes part of the molten pool. The particulate mixture of the residue and at least one metal selected from copper, silver, gold and platinum group metals are, in one set of embodiments, gradually added to a preformed molten pool of borax. In another set of embodiments the particulate mixture does not comprise particles of borax or other fluxing agents.
In one set of embodiments the smelting method comprises adding the treated solid residue to a previously melted pool comprising at least one metal selected from copper, silver, gold and platinum group metals.
In a further set of embodiments the smelting method comprises enclosing the treated solid residue is in a metal sheet or foil, preferably selected from at least one of copper, silver, gold and platinum group metals. The process can be conducted to avoid or minimize contact of the treated solid residue with the crucible.
The crucible used in the smelting may comprise a ceramic material (preferably a ceramic material that is relatively inert to corrosion when contacted by molten borax).
The smelting is preferably conducted in a crucible which comprises less than 10% by weight (preferably less than 5%) carbon and less than 10% by weight (preferably less than 5%) of carbides.
In one set of embodiments the reduced solid residue is added to the molten pool through a conduit such as a ceramic pipe that guides said material into the bulk phase of the molten pool. It is preferred that the residue does not encounter the walls of the crucible that contains the molten pool.
At the conclusion of the smelting process the molten pool may be poured into a mold to form an ingot, bullion bar or dore bar.
In a further aspect there is provided a system for recovery of gold from a concentrate that includes gold, the system comprising:

- (i) means for contacting the concentrate with an aqueous liquor to dissolve gold,
- (ii) an electrowinning cell for subjecting the product of step (i) to electrolysis to provide a cathode associated material,
- (iii) a crucible for contacting cathode-associated material from step (ii) with an aqueous leachant under reducing conditions, and
- (iv) means for smelting the product of step (iii) to recover gold.

The system in one set of embodiments further comprises a means for applying thermal or ultrasonic energy during any one of steps (i), (ii) or (iii).
In a preferred embodiment of the system the crucible comprises less than 10% by weight carbon and less than 10% by weight of carbides.
The invention will now be described with reference to the following examples. It is to be understood that the examples are provided by way of illustration of the invention and that they are in no way limiting to the scope of the invention.

EXAMPLES

Example 1

Silver Lake Gold Gravity Concentrate (SLGGC)—Source material
Gold gravity concentrate from the gold processing circuit at Silver Lake's Lakewood Gold production facility (near Kalgoorlie Australia) was stripped in caustic cyanide and the strip liquor processed in an electrowinning cell. Cathode material and cathode sludge from the cell was aggregated and soaked in 25% HCl for 2 hrs, to leach out steel wool from the sample. The residual material was rinsed and dried to provide 12.5 kg of source material. Grab lots of this material were taken and aggregated to procure a 500 g sample of source material.
The source material was homogenised by crushing and chopping, and multiple 10 g sub-samples were riffle split. 6 of these sub-samples were submitted for bullion analysis to the Perth Mint at Hay St East Perth, Western Australia. The bullion assay results were:


	Gold value	Silver value
Weight of sub-sample	(bullion assay)	(bullion assay)

10.17 g	60.43%	7.79%
10.05 g	60.37%	7.65%
10.03 g	60.05%	7.59%
10.01 g	60.37%	7.62%
10.02 g	60.63%	7.68%
10.06 g	60.53%	7.75%
	av = 60.4%	av = 7.68%

Example 2

Reducing Leach Step on SLGGC

One of the10 g sub-samples of SLGGC described above was added to a beaker with liquor comprising 200 ml of 50% HCl and 8 g stannous chloride. The contents of the beaker were heated to 80 deg C. and after 5 minutes the beaker was placed in a Soniclean 160T ultrasonic bath (bath water at 60 deg C., frequency 40 kHz, maximum power 250 W, power setting 60% of 250 W=150 W). After 5 minutes of ultrasonic agitation the beaker was re-heated and the cycle repeated 2 times. The residue was obtained by filtration, rinsed in water and dried. The (un-smelted) residue was found to have a lighter colour than the initial 10 g sub-sample, and was sent to the Perth Mint for bullion analysis, and the gold content (expressed as a percent of 10 g starting material) was found to be 61.77%, an increase from 60.42% in the starting material.

Alkaline Leach Step After Reducing Leach on SLGGC

A 10 g sub-sample of SLGGC source material was provided with a reducing leach as described above. Residue from the reducing leach step was added to 200 ml of a 10% sodium hydroxide liquor, and taken to 80 deg C. for 5 minutes, followed by 3 cycles of ultrasonic agitation as in the above example. The resultant residue was obtained by filtration, rinsed in water and dried. The (unsmelted) resultant residue was sent to the Perth Mint for bullion analysis, and the gold content (expressed as a percent of 10 g starting material) was found to be 61.94%, an increase from 60.4% in the starting material.

Example 3

Silver Lake Gold Carbon-in-Pulp Concentrate (SLGCIP)

Gold-loaded carbon derived from the CIP gold processing circuit at Silver Lake's Lakewood Gold production facility (near Kalgoorlie Australia) was stripped in caustic cyanide and the strip liquor processed in an electrowinning cell. Cathode material and cathode sludge from the cell was aggregated and soaked in 25% HCl for 2 hrs, to leach out steel wool from the sample. The residual material was rinsed and dried to provide 12.5 kg of source material. The source material was homogenised in a kitchen blender, and multiple 10 g sub-samples were riffle split. 6 of these sub-samples were submitted for bullion analysis to the Perth Mint at Hay St East Perth, Western Australia. The bullion assay results were


Weight of sub-	Gold value %	Silver value %
sample	(bullion assay)	(bullion assay

10.05 g	36.46	7.48
10.00 g	36.44	7.47
10.03 g	36.47	7.41
10.09 g	36.33	7.48
10.00 g	36.48	7.48
10.08 g	36.30	7.41
	av = 36.41%	av = 7.46%

Example 4

Reducing Leach Step on SLGCIP Source Material

One of the10 g sub-samples was added to a beaker with liquor comprising 200 ml of 50% HCl and 8 g stannous chloride. The contents of the beaker were heated to 80 deg C and after 5 minutes the beaker was placed in a “Soniclean 160T” ultrasonic bath (bath water at 60 deg C., frequency 40 kHz, maximum power 250 W, power setting 60% of 250 W=150 W). After 5 minutes of ultrasonic agitation the beaker was re-heated and the cycle repeated 2 times. The residue was obtained by filtration, rinsed in water and dried.

Alkaline Leach Step on SLGCIP Source Material

Residue from the reducing leach step (described above)was added to 200 ml of a 10% sodium hydroxide liquor, and taken to 80 deg C. for 5 minutes, followed by 3 cycles of ultrasonic agitation as described above. The resultant residue was obtained by filtration, rinsed in water and dried. The (un-smelted) resultant residue was sent to the Perth Mint for bullion analysis, and the gold content (expressed as a percent of 10 g starting material) was found to be 37.52%, an increase from 36.4% in the starting material. The silver content was found to be 7.58%, an increase from 7.46% in the starting material.

Smelting

A 10 g sub-sample of SLGCIP source material was taken through a reducing leach step and an alkaline leach step according to the above protocols, and the dry residue from the alkaline leach step was added to a 30 g fire assay crucible. The loaded crucible was placed inside an electric furnace pre-heated to 1220 deg C., and kept at this temperature for 1.5 hrs. When the crucible was withdrawn from the furnace, it contained a fluid phase comprising molten gold, and a dark solid phase that adhered to the base of the crucible. The liquid phase was poured into a button mould, and a clean separation achieved from the dark solid phase. After cooling, the button was removed from the mould and sent for bullion assay. The dark solid phase weighed 1.5 g. A portion of the dark solid phase (0.41 g) was added to 250 ml of freshly prepared aqua regia (1 part conc nitric acid and 4 parts conc hydrochloric acid) in a beaker at 80 deg C. After 5 minutes the beaker was placed in a Soniclean 160T ultrasonic bath (bath water at 60 deg C., frequency 20 kHz, bath setting at intensity 250 W). After 5 minutes of ultrasonic agitation the beaker was re-heated and the cycle repeated 2 times. Then 50 ml conc hydrochloric acid was added to the beaker and the beaker was re-heated and given one further 5-minute period of ultrasonic agitation at 60 deg C. Thereafter the liquor in the beaker was immediately filtered and sent for gold assay by flame AAS. The gold content of the button was found to be 2.64 g, and the leach-assay gold content of the dark solid residue was found to be 1.09 g. The total amount of recovered gold from the 10 g sub-sample was thus 3.73 g, an increase from 3.64 g in the starting material

Example 5

Silver Lake Gold Gravity Concentrate (SLGGC)—Source material

Reducing Leach Step on SLGGC

10 g sub-samples of Silver Lake Gold Gravity Concentrate (SLGGC) were prepared as previously described. The gold value (by bullion assay) in each sub-sample was 60.4%.
A 10.06 g sub-sample (particle size sub 250 microns) was added to a 500 ml beaker. Liquor comprising 8 g stannous chloride dihydrate (dissolved), 100 ml concentrated HCl and 100 ml water was added to the beaker, and the beaker was placed in a heated ultrasonic bath (Soniclean, maximum power=250 W) at 60° C. for 8 hours. Ultrasonic agitation (60% max setting) was applied according to the following schedule: 10 minutes initial sonication, 80 minutes pause, 10 minutes sonication, 80 minutes pause and so on to the end of the 8 hour period. No mechanical agitation was used.
After 8 hours, the contents of the beaker were filtered (Whatman 40 ashless filter paper, equivalent in filtration speed to Whatman 2) and the residue on the filter paper washed with water. The residue was then washed from the paper into another 500 ml beaker, and care was taken to use less than 100 ml of water to achieve this transfer. The water level in the beaker was made up to 100 ml, and 100 mls of 8% aqueous sodium hydroxide liquor was added to provide 4% final caustic leach liquor for the second leach. The beaker was placed in a heated ultrasonic bath and treated according to the above protocol. After filtration and water washing, the residue was dried in an oven at 80° C. overnight. The residue cake was readily disrupted to make a fine powder by simple mechanical stimulus with a spatula.
100 g of “fine silver” granules (plus 99.9% silver) was purchased from PW Beck & Co silver merchants of Adelaide, Australia. The granules were approximately 2 mm in diameter. Sheet silver (fine silver grade) of diameter 0.3 mm, with each sheet weighing 10 g was also purchased from PW Beck & Co.
The granules were placed in a 250 ml fire assay crucible purchased from Furnace Industries, of Perth Australia. The loaded crucible was placed inside an electric furnace and brought to 1220° C. Molten silver derived from the granules formed a small pool on the bottom of the crucible.
Dried residue derived from the caustic leach step described above was folded into a 10 g piece of sheet silver. The hot crucible containing the silver pool was withdrawn from the furnace, and the silver sheet envelope was dropped into the crucible directly onto the molten silver pool. The sheet silver melted quickly and the contents of the silver sheet envelope became immersed in the silver pool without making contact with the sides of the crucible. The crucible was immediately returned to the furnace, brought back to 1220° C. and retained at that temperature for 15 minutes. The molten contents of the hot crucible were poured into a hemispherical button mould, and allowed to cool. The button was dislodged from the mould and quenched in water, then allowed to dry. The approximate dimensions of the hemispherical button were: diameter 4 cm, max height 3 cm. The button was drilled out to obtain approx 6 g of shavings and burrs, which were sent for bullion assay Umpire Assay Laboratories, in Perth Australia.
The initial 10.06 g sub-sample comprised gold at 60.4% (multiple bullion assay results on replicate samples). The gold recovered from the button described above was 6.16 g and 0.14 g gold (total) was assayed on the filter papers used in the acid and alkaline leaching steps prior to smelting. This corresponds to a total of 6.3 g gold recovered from the initial sub-sample, compared to 10.06×0.604=6.076 g gold expected from the bullion assay on the initial sub-sample. The 0.368 g gold increment represents the benefit obtained by using the method of the invention.

Example 6

Silver Lake Gold Gravity Concentrate (SLGGC)—Source Material (a)

Gold loaded carbon from the gravity gold circuit at Silver Lake's Lakewood Gold production facility (near Kalgoorlie Australia) was stripped in caustic cyanide and the strip liquor processed in an electrowinning cell. Cathode material and cathode sludge from the cell was aggregated and soaked in 25% HCl for 2 hrs, to leach out steel wool from the sample. The residual material was rinsed and dried to provide 12.5 kg of source material. This source material was homogenised by crushing and chopping, and multiple 10 g sub-samples were riffle split. Apart from the 10 g sub-samples the remainder of the material was smelted using the standard Silver Lake process, and the commercially recoverable gold was found to be 77.06% gold.

Silver Lake Gold Carbon in Pulp (CIP) Concentrate—Source Material (b)

Gold loaded carbon from the C-I-P circuit at Silver Lake's Lakewood Gold production facility (near Kalgoorlie Australia) was stripped in caustic cyanide and the strip liquor processed in an electrowinning cell. Cathode material and cathode sludge from the cell was aggregated and soaked in 25% HCl for 2 hrs, to leach out steel wool from the sample. The residual material was rinsed and dried to provide 12.5 kg of source material. This source material was homogenised by crushing and chopping, and multiple 10 g sub-samples were riffle split. Apart from the 10 g sub-samples the remainder of the material was smelted using the standard Silver Lake process, and the commercially recoverable gold was found to be 35.04% gold.

Reducing Leach Step

Take 10 g sub-sample and add to reducing liquor. The leaching process as described in the first part of Example 4 (“Reducing Leach Step on SLGCIP source material”).
Note: If the reducing leach is not the first leaching step, use leach residue from the previous leaching step. Note that the 10 g sub-samples were used as source material in the initial leaching step.

Alkaline Leach Step

Take 10 g sub-sample and add to alkaline liquor. The alkaline leach is as described in the second part of Example 4 (“Alkaline leach step on SLGCIP source material”).
Note: If the alkaline leach is not the first leaching step, use leach residue from the previous leaching step. Note that the 10 g sub-samples were used as source material in the initial leaching step.

Leaching in 50% Nitric Acid

Take 10 g sub-sample and add to 200 ml of 50% Nitric acid liquor. Perform ultrasonic agitation, filtering, rinsing and drying as described in the first part of Example 4 (“Reducing Leach Step on SLGCIP source material”).
Note: If the acid leach is not the first leaching step, use leach residue from the previous leaching step. Note that the 10 g sub-samples were used as source material in the initial leaching step.

Borax Smelting

Filter and dry leached concentrate; mix 20 g of this material with 20 g borax and add to a crucible and heat the crucible to 1220° C. inside an electric furnace for 1.5 hours. Pour molten material from the crucible into the mold and allow to cool. Remove the contents of the mold and remove slag from the precious metal button. Weigh the button and send a sample of the button to the Perth Mint to establish the concentration of gold in the button.
Calculate the gold content of the source material (i.e. the initial cathode associated gold concentrate used in the leach sequence).

Silver Pool Smelting

Take 100 g of fine silver granules, add to a crucible and heat to 1220° C. in an electric furnace. Take sheet silver of diameter 0.3 mm (fine silver grade, 10 g per sheet) and wrap the sheet around the finely divided material to be smelted, (this material is the residue remaining after previous leaching steps on 10 g of sub-sample) to form a silver envelope. Remove the hot crucible containing molten silver from the furnace and drop the silver envelope into it. Immediately return the crucible to the furnace and reheat to 1220° C. for 15 minutes. Pour the molten contents of the hot crucible into a button mold and allow to cool. Remove slag and drill out the button to obtain shavings and burrs for bullion assay.
Residual gold on filter papers and in leach solutions are added to aqua regia and gold found in this way is added to gold bullion numbers.

TABLE 1

			Gold recovered
	Commercially		in silver pool
Source	recoverable		smelt after
material	gold	Leach sequence	leach sequence

b	35.04%	Reducing leach then	36.37%
		caustic leach
b	35.04%	Nitric leach then reducing	37.74%
		leach then caustic leach
b	35.04%	Reducing leach then caustic	37.22%
		leach then nitric leach
b	35.04%	Reducing leach then caustic	37.87%
		leach then 2 × nitric leach

TABLE 2

			Gold recovered
	Commercially		in silver pool
Source	recoverable		smelt after
material	gold	Leach sequence	leach sequence

a	77.06%	8 × nitric leach, then reducing	77.64%
		leach, then 2 × caustic leach

Example 7

Silver Lakes—500.97 g of wire gold (referred to a CIP-2) was received from Silver City Mining Company this was described as described as being Silver Lake Resources CIP plant material from Lakewood Gold Processing Facility. This sample was taken by representative sampling from a wire gold production run after hydrochloric acid treatment to remove the cathode wire and the gold grade of the sample (calculated by commercial smelting of the production sample with gold determination of the bullion bar by bullion assay from the Perth mint). The gold content was found to be 35.04% by weight.

Pre-Smelt Treatment

10 g of CIP-2 (representative subsample obtained by riffle splitting) was added to 200 ml of 50% by volume nitric acid in water in a 600 ml beaker. The beaker was placed in a heated ultrasonic bath at 60° C. (“Soniclean” brand, maximum power=250 W) and agitated at maximum power for one hour. The liquor was filtered off and the residue washed with water.
The water washed residue was added to a liquor comprising 8 g stannous chloride dehydrate (dissolved) 100 ml conc. hydrochloric acid and 100 ml water in a 600 ml beaker. The beaker was placed in a heated ultrasonic bath at 60° C. (“Soniclean” brand, maximum power=250 W) and agitated at maximum power for one hour. The liquor was filtered off and the residue washed with water.
The water washed residue from the previous step was added to 200 ml of 50% by volume nitric acid in water in a 600 ml beaker. The beaker was placed in a heated ultrasonic bath at 60° C. (“Soniclean” brand, maximum power=250 W) and agitated at maximum power for one hour. The liquor was filtered off and the residue washed with water, and dried.
All filter papers and liquors produced in the above operations were assayed for gold by standard techniques.

Smelt treatments

	Sample
Sample	Number	Particulate mixture	Crucible

CIP2	SLCIP2M19	2 g treated residue plus 2 gms	Clay
		silver powder plus 20 g borax
	SLCIP2M20	2 g treated residue plus 2 gms	Clay
		copper powder plus 20 g borax

Note:
In the 3 above smelts, the particulate mixture was placed in a 250 ml fire assay crucible purchased from Furnace Industries of Perth, Australia. The loaded crucible was placed inside an electric furnace and brought to 1220° C. and retained at that temperature for 15 minutes. The molten contents of the hot crucible were poured into a preheated hemispherical button mold and allowed to cool. The button was dislodged from the mold and quenched in water then allowed to dry. Slag was separated and the button was sent for bullion assay to Umpire Assay Laboratories in Perth, Australia.

Total Gold Recovery (from smelt and solutions/filter paper)

		%
		Upgrade
	Sample	from
Gold Recovered (g)	Gold	com-

	Initial		Solutions/		Head	mercial
Sample	Weight		filter		Grade	smelt
Number	(g)	Smelt	paper	TOTAL	(%)	results

SLCIP2M19	10.04	3.248	0.442	3.69	36.75	4.88
SLCIP2M20	10.18	3.279	0.402	3.68	36.15	3.17

Claims

1. A method for recovering gold from a gold concentrate comprising:

dissolving gold from the concentrate in an aqueous liquor to provide a gold liquor;

subjecting the gold liquor to electrolysis in an electrowinning cell to provide cathode-associated gold material;

leaching the cathode associated gold material in an aqueous liquor under reducing conditions to provide a treated solid residue; and

smelting the treated solid residue to recover gold.

2. The method according to claim 1 wherein the gold concentrate comprises carbon sorbed with gold and the method further comprises stripping gold from the carbon by contact with an aqueous liquor to provide said gold liquor.

3. The method according to claim 1 wherein the gold concentrate comprises a gravity concentrate and the method further comprises dissolving gold from the gravity concentrate to provide said gold liquor.

4. The method according to claim 1 wherein leaching the cathode associated gold material under reducing conditions involves use of an aqueous leach composition selected from the group consisting of aqueous base, aqueous acid, aqueous chelating agent, and mixtures of aqueous chelating agent with acid or base.

5. The method according to claim 1 wherein the aqueous reducing liquor is an acidic aqueous reducing liquor of pH less than about 1.5.

6. The method according to claim 1 wherein the aqueous reducing liquor is an alkaline aqueous reducing liquor.

7. The method according to claim 1 wherein the method further comprises leaching the deposited gold concentrate, prior to said reducing leach step, in an aqueous liquor comprising one or more agents selected from the group consisting of hydrochloric acid, nitric acid, alkali, lead acetate, chelating agents, carboxylic acids and their salts, chlorates, perchlorates, chlorides, fluorosilicate, phenol sulfonate, and peroxydisulfate.

8. The method according to claim 1 wherein the method further comprises subjecting the solid residue from leaching in aqueous reducing liquor to at least one leaching step in an aqueous liquor comprising agents selected from the group consisting of hydrochloric acid, nitric acid, alkali, lead acetate, chelating agents, carboxylic acids and their salts, chlorates, perchlorates, chlorides, fluorosilicate, phenol sulfonate, and peroxydisulfate.

9. (canceled)

10. The method of claim 8 wherein the contact between at least one of the gold concentrate and residue and the aqueous liquor is carried out under conditions of ultrasonic agitation.

11. The method for recovering gold according to claim 10 wherein a mixture of the cathode-associated gold concentrate and reducing liquor is subject to ultrasonic radiation at a frequency in the range 10-60 kHz.

12. The method for recovering gold according to claim 1 wherein the cathode comprises steel wool and the method comprises treating the cathode having deposited gold material to remove the steel wool.

13. The method of claim 1 comprising a plurality of steps involving contact of the cathode associated gold material with an aqueous reducing liquor.

14. The method according to claim 1 wherein the reducing conditions are provided by a reducing agent.

15. The method according to claim 1 wherein the reducing conditions are provided by at least one reducing agent comprising metal species selected from the group consisting of chromium (Cr II), tin (Sn II), copper (Cu I) and titanium (Ti II, Ti III) and non-metal containing reducing agents selected from sulfites, organic acids with sulfites and oxalic acid.

16. The method according to claim 1 wherein the aqueous reducing liquor comprises stannous ion.

17. The method according to claim 1 comprising said reducing leach and a subsequent acid leach wherein the liquor used in the reducing leach comprises hydrochloric acid and stannous chloride and the liquor used in the subsequent acid leach comprises nitric acid in water.

18. The method according to claim 17 wherein the weight ratio of liquid to solid material in the subsequent acid leach step is in the range of 10:1 to 100:1.

19. The method according to claims 14 further comprising subjecting the solid residue from leaching in aqueous reducing liquor to leaching with an alkaline liquor.

20. The method according to claim 19 wherein the aqueous alkaline liquor has a pH greater than 13.

21. The method for recovering gold according to claim 19 wherein the alkaline liquor comprises at least 5% sodium hydroxide.

22. The method according to claim 1 wherein the reducing liquor comprises at least one agent selected from organic acids and their salts.

23. The method according to claim 1 wherein the reducing liquor comprises at least one base metal chelating agent selected from the group consisting of beta-diketones, amino polycarboxylic acids, salts of amino polycarboxylic acids, carboxylic acids, salts of carboxylic acids, and polyphosphonates.

24. The method according to claim 1 wherein the material subject to leaching in an aqueous reducing liquor is finely divided.

25. The method according to claim 1 wherein the material subject to leaching in aqueous liquor is finely divided providing on wet sieving at least 80% by weight passing through a 100 micron sieve.

26. The method according to claim 1 wherein the smelting of said residue in a crucible is conducted without addition of a flux.

27. The method according to claim 1 wherein the smelting of said residue is conducted in the presence of a borax flux.

28. The method according to claim 1 wherein the step of processing the residue to recover gold comprises forming a molten pool comprising at least one metal selected from copper, silver gold and platinum group metals.

29. The method according to claim 28 wherein the molten pool is formed from a solid particulate mixture comprising particles of treated solid residue and particles of at least one metal selected from copper, silver, gold and platinum group metals.

30. The method according to claim 28 wherein the molten pool is poured into a mold to form an ingot, bullion bar or dore bar.

31. The method according to claim 28 wherein the molten pool has a melting point in excess of 900° C.

32. The method according to claim 28 wherein the pool metal comprises silver, copper or mixtures thereof.

33. The method according to claim 28 wherein a particulate mixture of the residue and at least one metal selected from copper, silver gold and platinum group metals is gradually added to a heated crucible such that a molten pool is formed during addition and further particulate mixture is added to and becomes part of the molten pool.

34. The method according to claim 28 wherein a particulate mixture of the residue and at least one metal selected from copper, silver, gold and platinum group metals are gradually added to a preformed molten pool of borax.

35. The method according to claim 34 wherein the particulate mixture does not comprise particles of borax or other fluxing agents.

36. The method according to claim 28 wherein the melting step comprises adding the treated solid residue to a previously melted pool comprising at least one metal selected from copper, silver, gold and platinum group metals.

37. The method according to claim 28 wherein at least part of the treated solid residue is enclosed in a metal sheet or foil, preferably selected from at least one of copper, silver, gold and platinum group metals.

38. The method according to claim 27 wherein smelting is conducted in a crucible comprising a ceramic material, inert to molten borax.

39. The method according to claim 38 wherein smelting is conducted in a crucible which comprises less than 10% by weight carbon and less than 10% by weight of carbides.

40-43. (canceled)