FIELD

This invention relates to a method and system for the recovery of precious metals from source materials containing precious metals and in particular, from ores refined to enrich precious metals.

BACKGROUND

Examples of materials that contain precious metals enriched from ores include:

- (a) precious metals sorbed on activated carbon or absorbent resin;
- (b) concentrates derived by ashing of precious metal containing source material to remove organic material and carbon;
- (c) cathode-associated material formed during electrolysis of a strip liquor. The strip liquor may arise when gold is stripped from an activated carbon;
- (d) anode-associated material formed during the electrolytic refining of copper from a copper anode; and
- (e) gravity gold concentrates.

The process of gold recovery frequently involves a leaching step and adsorption of gold and other precious metals onto and 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 solubilising liquor to form a strip liquor containing the precious metals stripped from the absorbent.

Precious metals including gold and silver may be recovered from the strip liquor in an electrowinning process in which the precious metals are deposited from the strip liquor onto the cathode of an electrowinning cell.

The electrode-associated material includes materials such the direct cathode deposits and electrode-associated sludge which may collect on or below the cathode of a gold electrowinning cell.

The electrode-associated material may also comprise anode mud from a base metal refining process such as a copper refining process.

For cathode associated source materials the cathode used in the process is usually a high surface area cathode, and may comprise steel wool. Both the material deposited on the cathode (often called wire-gold where the cathode is steel wool) and cathode slimes (deposits which collect beneath and in association with the cathode) are rich in precious metals, and the next step in the precious metal recovery process usually involves acid treatment to remove steel wool, followed by smelting and bullion formation.

When copper is refined by electrolysis the anodes are frequently cast from processed blister copper placed into an aqueous solution of 3-4% copper sulfate and 10-16% sulfuric acid. Cathodes are often thin rolled sheets of highly pure copper. At the anode, copper and less noble metals dissolve. More noble metals such as silver and gold as well as selenium and tellurium settle to the bottom of the cell as anode mud, which forms a saleable by-product. The anode mud therefore includes the anode associated gold.

Platinum together with the rest of the platinum group metals is obtained commercially as a by-product from nickel and copper mining and other base metal mining operations. During electrorefining of copper, noble metals such as silver, gold and the platinum group metals as well as selenium and tellurium settle to the bottom of the cell as anode mud, which forms the starting point for the extraction of the platinum group metals.

Silver and gold may be recovered from the anode mud by using concentrated sulphuric acid to dissolve copper and other impurities and casting the remaining noble metals into anodes for electrorefining.

The anode mud may be subject to oxidising fusions to provide a dore bullion of silver, gold and platinum group metals. The nature of the operation and flux may depend on the composition of the anode mud. The dore bar will typically contain the silver, gold and platinum group metals which may be parted to separate and purify the gold and silver and recover platinum group metals. Parting involves the use of selective solvents to recover precious metals. Acid parting may involve boiling the dore material in strong sulphuric acid such as 96%-98% sulphuric acid to dissolve silver and some platinum group metals followed by recovery of the metals from solution. Gold remains undissolved in sulphuric acid. Platinum group metals when present may be separated from the silver for example by casting the silver into anodes and refining electrolytically. The Balbach/Thum and Moebius processes of parting electrolytically may be used and are often better suited to large copper refineries.

Aqua regia may be used as a parting solvent to dissolve gold. The solubilised gold may be recovered by solvent extraction, for example using E444 (butyl diglyme) as the solvent to capture the gold.

A common method used to process gold-rich electrode associated material involves smelting that material. Smelting involves placing the source material in a crucible, adding fluxing agents and heating to about 1250Â° C. Base metal contaminants are collected in the floating slag layer that forms over the molten precious metals. After cooling the slag can be physically separated from the dore metal bar and further processing can take place to obtain more highly purified gold.

Gravity precious metal concentrates comprise precious metal concentrated by a gravity process. Such concentrates are commonly prepared from placer deposits of precious metals such as gold and platinum group metals. Gravity concentration has been historically the most important way of extracting the native metal using pans or washing tables. In some cases, particularly when the precious metal is present in the ore as discrete coarse particles, a gravity concentrate can be in some cases be directly smelted to form, for example, gold bars. In other cases, particularly when the precious metal is present in the ore as fine particles or is not sufficiently liberated from the host rock, the concentrates are treated by leaching, such as cyanide leaching in the case of gold, followed by recovery from the leach solution. Recovery from solution may involve adsorption on activated carbon and/or electrolysis to form cathode associated deposits.

Smelting of wire-gold, gravity concentrates, cathode slimes and/or anode mud followed by first bullion formation is very convenient from a security point of view, and industry standard practice involves use of this method. The selective solubilisation of precious metals from source materials is a convenient method of capturing multiple precious metals in purified form.

There is a need for an efficient recovery of precious metals from concentrates such as precious metals sorbed on activated carbon, concentrates derived by ashing, cathode-associated material, anode mud and gravity gold concentrates.

SUMMARY

There is provided a method for recovering precious metals from source materials containing precious metals wherein the source material is selected from the group consisting of carbon sorbed precious metals, ashed concentrates, electrode-associated material from an electrolytic process, and gravity concentrates, the method comprising leaching in aqueous reducing liquor to provide a reduced solid residue and processing the residue to recover precious metals.

The processing of the treated residue may comprise one or more of smelting the residue, selectively solubilising one or more precious metals and electrorefining.

The source material is preferably, but not restricted to, particulate form such as finely divided particles. For example the method may use a processed source material which on wet sieving provides at least 50% by weight of particles passing through a 100 micron sieve preferably at least 80% passing through a 100 micron sieve. Alternatively, the source material may be coarser, such as the product of gravity concentration.

The source material is preferably an electrode associated material particularly (a) cathode associated precious metals produced in electrowinning of gold from aqueous liquors used to dissolve gold such as strip liquor used to remove gold from sorption on carbon; or (b) anode mud produced in refining of base metals such as copper or nickel (particularly copper).

In one set of embodiments processing the residue to recover precious metals comprises forming a molten pool comprising at least one metal selected from the group consisting of copper, silver, gold and platinum group metals; and adding at least part of the processed residue into the pool of molten metal.

The molten metal may be poured into a mold to form a dore or bullion bar.

In a preferred embodiment the method further comprises isolating the reduced solid residue and treating the residue by leaching in an aqueous liquor comprising at least one agent selected from the group consisting of hydrochloric acid, nitric acid, alkali, lead acetate, ammonium chloride, calcium chloride, strontium chloride, acetic acid, chelating agents or any agent which enhances the solubility of lead, or lead oxide, or other lead moieties in water. In this embodiment the product of the further treatment may be further refined to provide precious metals.

The solid residue or product from treatment of the residue 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 or product of further treatment to a crucible, and heating the contents to smelting temperature. The smelting process may include the use of an included flux or may be flux-less. For example, if copper is used in the process, a pre-formed pool of flux may be preferred in order to inhibit oxidation of the copper. The flux may include any suitable agent known in the art including borax, silica sodium carbonate and the like.

In one embodiment the solid residue is leached with an aqueous alkaline liquor and the solid residue from the alkaline liquor is further refined to provide precious metal.

In one embodiment of the invention the method optionally comprises preparing a material for said leaching in aqueous reducing liquor by a method comprising at least one leaching step in an aqueous liquors comprising agents selected from the group consisting of hydrochloric acid, nitric acid, alkali, lead acetate, ammonium chloride, calcium chloride, strontium chloride, acetic acid, chelating agents or any agent which enhances the solubility of lead, or lead oxide, or other lead moieties in water.

DETAILED DESCRIPTION

Anode Mud refers to a solid substance or mixture that collects at the anode in an electrolytic refining or plating process. It is generally insoluble in the aqueous liquors used as the electrolyte in electrolytic cells. Anode mud is also referred to as anode slime.

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+2H₂Oâ†’4Na[Au(CN)₂]+4NaOH

Cathode associated precious metals include gold deposited on the cathode or which is formed adjacent the cathode and may for example collect in the electrowinning cell below the cathode

Ashing is a pyrolytic process for removing carbon and organic material from a source material such as gold loaded carbon or gold loaded resin.

Precious metals include gold, silver and platinum group metals. The method is particularly suited to recovery of precious metals from precious metal concentrates.

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. 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 â€œcompriseâ€ and variations of the word, such as â€œcomprisingâ€ and â€œcomprisesâ€ 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 combination of two or more 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.

Examples of suitable non-metal containing reducing agents include 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 reducing leach may produce at least partial removal of a base metal from the source material. 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, â‰§chromium (II) chloride>oxalic acid

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 comprises leaching the precious 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 precious metal concentrate 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 dislodgment 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 above ambient, preferably at least 60Â° 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%.

Preferably at least one contact step between source material and reducing aqueous liquor leads to a bleaching of the source material. The bleaching may be measured using quantitative colorimetric methods, such as the LAB method.

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 dislodgment of refractory material from the surface of the solid.

In one set of embodiments the leaching is conducted at a temperature above ambient, preferably at least 60Â° C.

Such conditions may include at least one of those selected from the group consisting of ultrasonic agitation and stimulation by time variant electrical and/or magnetic field.

Preferably at least one contact between source material and reducing aqueous liquor leads to the removal of at least part of at least one base metal from the source material.

Preferably the base metal comprises at least one selected from the group consisting of iron and lead.

In one set of embodiments the leaching is conducted at a temperature of at least 60Â° C.

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 precious metal 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 to form a strip liquor containing the precious metals stripped from the absorbent.

In one embodiment the precious metal-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 precious metal concentrate source material comprises gold sorbed onto carbon.

In yet another embodiment the source material comprises anode mud from a copper refining process.

In a preferred embodiment the reducing liquor in at least one contact with source material comprises at least one base metal 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 of the method the source material is finely divided.

For example, the method may use a source material which on wet sieving at least 50% by weight of particles pass through a 100 micron sieve preferably at least 80% pass through a 100 micron sieve.

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 concentrated nitric acid (preferably diluted to about 5 to about 70%, more preferably 20% to 60% and even more preferably about 50% v/v 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).

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, citric acid or any agent which enhances the solubility of lead, or lead oxide, or other lead moieties in water. 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 40Â° C. to 90Â° C. and more preferably from 60Â° C. to 80Â° 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 or 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 the 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. An example of such conditions is 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 60Â° C.

Without wishing to be bound by theory, it is believed that one benefit of the method of the invention diminishes the impact of interfering substances that impede the recovery of precious metals.

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 method may be refined by one or more of smelting the residue and/or selectively solubilising one or more precious metals.

When the method uses a step of selectively solubilising one or more precious metals the method may comprise at least one of:

- i) removal of silver with an acid particularly sulphuric acid or nitric acid, to form a silver solution;
- ii) aqua regia leach (optionally following removal of silver) to form a gold solution. Gold may be recovered from acid/chloride solution by extraction into an organic solvent particularly E444; and
- iii) electrorefining to form purified precious metal concentrate at the cathode.

The smelting process may include the use of an included flux, such as a flux comprising borax, or may be flux-less.

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 accordance with developments made following this finding the method preferably further comprises:

smelting the reduced solid residue by forming a molten pool comprising a metal selected from gold and metals which form alloys with gold; and

adding the reduced solid residue into the pool of molten metal.

In one set of embodiments, the pool metal is selected from the group consisting of copper, silver, gold, precious metals. In one preference, the pool metal comprises silver or copper. In one set of embodiments of the invention the reducing leaches or other leaches remove sufficient base metals from the source material so that slag formation in fluxless smelt is less than 1% (preferably less than 0.1%) by weight of the molten pool. Slag formation can be determined by observing the presence of a distinct phase other than metal. The slag will typically contain compounds formed between metals and non metals particularly metal oxides.

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.

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.

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 900Â° 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 or other fluxing agents. The flux may include any suitable agent known in the art including borax, silica sodium carbonate and the like. 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 at least partially 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 such as clay).

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.

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.

EXAMPLESExample 1Spent Activated Carbon (SAC)

90 kg of spent activated carbon was taken from a CIP gold processing operation and the carbon riffle-split into 5 kg lots. One of these lots was again riffle split into 100 g sub-samples.

SAC1

100 g spent activated carbon was added to a 5 L beaker. 2 litres of concentrated hydrochloric acid also containing 5 g stannous chloride (dissolved) was then added to the carbon in the beaker. Every hour, the liquor was gently stirred by hand using a glass stirring rod. After 6 hours the contents of the beaker were poured through a 0.35 mm stainless steel screen. The liquor passing through the screen was collected in a bucket and allowed to settle. After settling, clear liquor was decanted, and the residual solids in the bucket were filtered, to provide a first solid fraction. The solids remaining on the screen were well washed with 10% hydrochloric acid to provide a second solid fraction. First and second solid fractions were combined and dried at 40Â° C.

SAC2

100 g spent activated carbon was added to a 5 L beaker. 2 litres of concentrated hydrochloric acid also containing 10 g stannous chloride (dissolved) was then added to the carbon in the beaker, with gentle stirring. The beaker was allowed to stand overnight, and the supernate was decanted. The solids were washed twice with 2 litres of water before screening and drying at 40Â° C.

SAC3

100 g spent activated carbon was added to a 5 L beaker. 2 litres of concentrated hydrochloric acid also containing 40 g stannous chloride (dissolved) was then added to the carbon in the beaker, with gentle stirring. After 65 hours, the supernatant liquor was decanted and 1 litre of 10% HCl solution was added. After 3 hours this was also decanted, and 1 litre of water was added and allowed to stand for 2 hours. This was followed by decanting and the addition of another 1 litre of water with standing for 2 hours. The carbon was screened at 0.5 mm, rinsed with 1 litre of water and dried at 40Â° C.

SAC4

As for SAC3 except in the first leaching step, 2 litres of 50% HCl also containing 40 g stannous chloride (dissolved) were used.

Gold Assay

Samples of un-leached carbon and of carbon leached according to the protocols of SAC1-SAC4 (see above) were sent to Amdel laboratories in Adelaide, South Australia for gold fire assay. The results (duplicates) were:

SampleGold 1(ppm)Gold 2(ppm)Gold average (ppm)unleached carbon385375380SAC1445450448SAC2415425420SAC3435445440SAC4420415418

Example 2Silver Lake Gold Gravity Concentrate (SLGGC)â€”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 hours, 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. Six of these sub-samples were submitted for bullion analysis to the Perth Mint at Hay Street, East Perth, Western Australia. The bullion assay results were:

Weight of sub-Gold value Silver value sample(bullion assay)(bullion assay)10.17 g60.43%7.79%10.05 g60.37%7.65%10.03 g60.05%7.59%10.01 g60.37%7.62%10.02 g60.63%7.68%10.06 g60.53%7.75%av = 60.4%av = 7.68%

Example 3Reducing Leach Step on SLGGC

One of the 10 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 80Â° C. and after 5 minutes the beaker was placed in a Soniclean 160T ultrasonic bath (bath water at 60Â° 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 80Â° 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 4Silver 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 hours, 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. Six of these sub-samples were submitted for bullion analysis to the Perth Mint at Hay Street, East Perth, Western Australia. The bullion assay results were

Weight of sub-Gold value % Silver value % sample(bullion assay)(bullion assay10.05 g36.467.4810.00 g36.447.4710.03 g36.477.4110.09 g36.337.4810.00 g36.487.4810.08 g36.307.41av = 36.41%av = 7.46%

Example 5Reducing Leach Step on SLGCIP Source Material

One of the 10 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 80Â° C. and after 5 minutes the beaker was placed in a â€œSoniclean 160Tâ€ ultrasonic bath (bath water at 60Â° 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 80Â° 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 1220Â° C., and kept at this temperature for 1.5 hours. 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 80Â° C. After 5 minutes the beaker was placed in a Soniclean 160T ultrasonic bath (bath water at 60Â° 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 60Â° 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 6

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 60Â° 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 80Â° C. overnight. The residue cake was readily disrupted to make a fine powder by simple mechanical stimulus with a spatula.

Fine silver granules (plus 99.9% silver) were 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.

100 g of 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 1220Â° 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 1220Â° 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 10.06Ã—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 7Silver Lake Gold Gravity Concentrate (SLGGC)â€”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) Concentrateâ€”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 (â€œReducing Leach Step on SLGCIP source materialâ€ ).

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 (â€œAlkaline leach step on SLGCIP source materialâ€ ).

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 (â€œReducing Leach Step on SLGCIP source materialâ€ ).

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 1220Â° 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 1220Â° 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 1220Â° 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 1Gold recoveredCommerciallyin silver poolSourcerecoverablesmelt after materialgoldLeach sequenceleach sequenceb35.04%Reducing leach then 36.37%caustic leachb35.04%Nitric leach then reducing 37.74%leach then caustic leachb35.04%Reducing leach then caustic37.22%leach then nitric leachb35.04%Reducing leach then caustic37.87%leach then 2 Ã— nitric leach

TABLE 2Gold recoveredCommerciallyin silver poolSourcerecoverablesmelt after materialgoldLeach sequenceleach sequencea77.06%8 Ã— nitric leach, then 77.64%reducing leach, then 2 Ã— caustic leach

Example 8

Silver Lakesâ€”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 60Â° C. (â€œSonicleanâ€ 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 60Â° C. (â€œSonicleanâ€ 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 60Â° C. (â€œSonicleanâ€ 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 treatmentsSampleSampleNumberParticulate mixtureCrucibleCIP2SLCIP2M192 g treated residue plus 2 gmsClaysilver powder plus 20 g boraxSLCIP2M202 g treated residue plus 2 gmsClaycopper powder plus 20 g boraxNote: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 1220Â° 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)%SampleUpgradeGold Recovered (g)GoldfromInitialSolutions/HeadcommercialSampleWeightfilterGradesmeltNumber(g)SmeltpaperTOTAL(%)resultsSLCIP2M1910.043.2480.4423.6936.754.88SLCIP2M2010.183.2790.4023.6836.153.17