A diagram of the laboratory mineral processing equipment used in the precipitation of copper and the subsequent regeneration of cyanide. The general technique employed in these operations was as follows:

1) To pregnant solution in a 1-liter stainless steel precipitator (A) add the sulfide (Na2S; NaHS; CaS) required for precipitation.

2) Bolt the head of the precipitator in place over a Teflon gasket; add the required volume of sulfuric acid to the acid storage container (B).

3) Add acid through valve (C) into the pregnant-sulfide mixture in (A) with valves (H) and (D) closed and with magnetic stirrer (E) operating.

4) Keep system closed and stir for 5 min. to complete precipitation.

5) Open outlet valve (D) to allow the evolved cyanide to pass into traps (F) and (G) containing alkaline absorbing solution.

6) For gas stripping, connect inlet to a source of air or nitrogen, open inlet valve H, and pass gas into and upward through the contents of the precipitator (A) and then out through exit valve (D) into the traps (F) and (G).

7) For steam stripping, a water-cooled condenser and receiver (not shown) are inserted between (A) and (G), and (A) heated and stirred to remove a cyanide-water mixture which is collected (F) and cooled in an ice-water mixture. (F) is vented into trap (G) to prevent loss of cyanide.

8) The cyanide contents of (F) and (G) were determined by direct titration with silver nitrate. The copper bearing precipitate was filtered from the liquors in (A) after the stripping operation, was dried, and analyzed for copper content.

The types of receivers (F) and (G) were varied to facilitate cyanide recovery and manipulation (cooling, volume of distillate handled, quantity of cyanide recovered etc.).


In a fast mineral processing reaction, copper minerals dissolve in cyanide and act as cyanides in the recovery of precious metals by cyanidation. However, the application of cyanide leaching to the recovery of copper from copper ores requires considerably different conditions than used in precious metal cyanidation (higher cyanide concentrations; shorter leaching times; oxidizing conditions unnecessary).

Cyanide Complexes of Copper: The predominant species formed when copper dissolves in cyanide is one containing 3 moles of cyanide to 1 mole of copper (Cu(CN)3). This indicates that at least 2.32 lb NaCN equivalent are required to dissolve one pound of copper, if no side reactions occur.

Cuprous Copper Minerals: The reaction of cyanide with typical cuprous sulfide and oxide copper minimum. Fig. 1:—Simplified precipitation-regeneration system. Not shown in the diagram are a condenser and receiver for steam stripping.

The reaction is generally written as follows:


CU2S + 6 NaCN = 2 Na2Cu(CN)3 + Na2S [1]

CuaS + 3 Ca(CN)a = 2 CaCu(CN)a + CaS [21


Cu20 + 6 NaCN + HaO = 2 Na2Cu(CN)3 + 2 NaOH [SI Cu20 + 3 Ca(CN)a + HaO = 2 CaCu(CN)3 + CalOH). [41

Cupric Copper Minerals: Cupric copper undergoes reduction in cyanide solution. Therefore, when cupric minerals are dissolved in cyanide, the following reactions occur:

2 CuC03 + 8 NaCN = 2 Na2Cu(CN)3 + 2 NaaCOa + (CN).

Cyanogen, (CN)2, in alkaline solution, undergoes a further reaction as follows:

(CN|! + 2 NaOH = NaCNO + NaCN ■ H-.0

The over-all reaction therefore is:
2 CuCOa + 7 NaCN + 2 NaOH = 2 Na2Cu(CN)3 + 2 NaaCOa + NaCNO + HaO
In this reaction, the reduction of the cupric copper and the subsequent oxidation of cyanogen to form cyanate results in the loss of 0.5 mole of NaCN for each mole of copper dissolved (0.39 lb NaCN equiv./lb Cu).

Thiocyanate Formation: A second important source of cyanide loss occurs through the formation of thiocyanates. The exact mechanism of this thiocyanate formation is not known. The reaction of cyanide with sulfides and thiosulfates does not readily produce thiocyanate. However, higher poly-thionates such as the tetrathionates and pentathio-nates react readily with cyanide to give thiocyanate as follows:

NaaSiOo + 3 NaCN + HaO =

NaaSOi + 2 HCN + Na2S2Oj + NaSCN [81 Na_S:,0.; + 4 NaCN + HaO =

NaaSOi + 2 HCN + NaaSaOa + 2 NaSCN [91

Ferrocyanide Formation: In general, ferric iron is not soluble in alkaline cyanide solutions. Ferrous iron, however, does dissolve to form ferrocyanide, Fe(CN),f‘. Since ferrocyanides do not decompose readily in cold, dilute sulfuric acid (used in subsequent copper recovery), such compounds may constitute a source of cyanide loss in the copper cyanidation process.

Leaching Tests on Copper Minerals: Five-gm samples (minus 100 mesh) of several copper min­erals (Ward’s National Science Establishment, Roch­ester, New York) were leached for six hours in 1000 ml. of NaCN solution at various cyanide to copper ratios. In the tests on bornite and chalco- pyrite, 10-gm samples of mineral were leached for four hours using Ca(CN)a solutions. Cyanide con­sumptions were calculated by analysis of the preg­nant solutions for non-regenerative cyanide loss (by distillation) and for ferrocyanide (by iron analysis).

Rate of solubility: The rapid rate of extraction of copper from several copper minerals was demon­strated by leaching 50-gm samples of synthetic ores (2% Cu, —200 +325 mesh) prepared from various copper minerals and quartz. These ores were leached in open beakers at 40% solids and a cyanide ratio of 3.0 gm NaCN equivalent per gm of con­tained copper. Results are shown in Fig. 2. As in­dicated, the rate of extraction of copper from the minerals tested increased in the order of bornite, covellite, chalcocite and malachite. Copper extrac­tions of 81.6% (chalcocite) and 94.2% (malachite) were obtained in 15 min., thus demonstrating that samples of a copper ore (minus 28 mesh and minus 10 mesh) and samples of flotation tailings and cleaner flotation tailings were leached with cyanide. The predominant copper mineral in these samples was chalcocite. Samples of the ore suspen­sion were withdrawn periodically, immediately fil­tered and washed and the filter cakes assayed for copper.

As indicated in Fig. 3, the rate of solubility of copper was rapid in the case of the flotation tailing (1.08% Cu), although the cyanide to copper ratio (2.34:1) was lower than that required to give op­timum copper extraction (i.e., usually 3.0-3.5:1). Leaching of the copper ore at coarser sizes also was rapid, since from the minus 10 mesh sample, 75.1% of the total copper dissolved in the first 30 min and 86.7% in 4 hr. From the minus 28 mesh sample of the same ore, 83.9% of the copper was extracted in 30 min and in 2 hr essentially all the extractable copper was in solution.

Leaching was conducted in a laboratory Fagergren flotation machine and in the first two min. 84% of the copper was dissolved. Thus the rapid solubiliz­ing action of cyanide on copper was demonstrated clearly and was subsequently utilized in flotation operations to eliminate the activation of pyrite by contaminating surface films of copper and to permit the production of final copper concentrates of high grade.

Gold Leaching Plant Flowsheets

The Kalgoorlie ore deposits occur in Pre-Cambrian rocks and essentially consist of calc schist and quartz-dolerite-greenstone. The ores contain free gold and tellurides. Auriferrous pyrite is the most abundant sulphide and is occasionally associated with chalcopyrite, tetrahedrite and arseno-pyrite. The gold associated with the pyrite is in a very fine state and is not liberated even after very fine grinding.

The crushing section reduces run-of-mine ore to minus 1/2 inch which is further reduced by rod mills and ball mills to approximately 75% minus 200 mesh. Rod mill discharge is pumped over strake tables ahead of the ball mills with the concentrate being amalgamated, retorted and finally smelted. Final ground product goes direct to flotation before gold leaching by cyanide.

The pyritic flotation concentrate and tailing are treated in separate thickeners with the overflow from both thickeners going to flotation circuit water storage. Underflows from each thickener are cyanided in separate agitators.

The cyanided concentrate is filtered and the filter cake treated in Edwards roasters with the calcine again being cyanided and filtered. Post-cyanidation is the term used to describe this treatment method. The calcine filter cake is repulped and pumped to the flotation tailings cyanide agitators. Filtrates from both stages of filtering are combined and flow to the high grade precipitation section, clarified, and the gold precipitated with zinc dust. This precipitate is then filtered, roasted, fluxed and smelted.

The tailings cyanide agitators discharge to a washing thickener. The underflow is filtered and the filter cake repulped with barren solution and pumped to the residue dam. Filtrate is circulated back to the washing thickener. Overflow from the washing thickener passes to the low grade precipitation section and is treated similarly to the high grade solution. Barren solution is circulated back to both cyanide sections.

Until 1956, a second 500 ton per day mill operated using a pre-cyanidation flowsheet. In this case, the ground ore was cyanided with the cyanide tailings filter cake) being gassed with sulfur dioxide and then floated for sulfide recovery. Flotation concentrate was then roasted, recyanided, and returned to the grinding circuit. In 1956 this plant was converted to the post cyanide flowsheet.Gold and Silver Leaching

The Emperor flowsheet is so different that it is presented in detail on two pages. It is unique in that tellurium metal is produced along with gold bullion and copper cement both unusual products from a gold ore. The flowsheet is distinguished in that there are three separate but allied circuits; the main cyanidation circuit supported by roaster calcine cyanidation, a tellurium metal recovery circuit, and a copper cement recovery operation.

The mine ore is washed in a blade type mill after which the minus 4 plus 1/2 inch product is hand picked for removal of waste. The slime portions go to flotation for recovery of sulfides with the slime tailings going to waste.

After grinding, the tellurium and copper minerals are recovered by flotation. The concentrate containing some of the gold goes through an oxidation step with NaOH, Na2C03 and Ca (OCl which renders the gold soluble for cyanidation and the tellurium leachable. The concentrate residue containing copper is roasted and leached for production of cement copper.

The flotation tailings are cyanided for gold extraction following usual practice. The cyanide residue is conditioned with SO£ gas from the copper concentrate Edwards roaster and then floated with copper sulphate and xanthate. After filtration, the concentrate is fed to the roasters along with the above main stream copper concentrate. Flotation tailings are used for mine fill.

Roaster calcines, following leaching for copper, are washed and cyanided separately and then the calcine-residue is fed back through the gassing tower circuit.Gold and Silver Plant flowsheet


Alternative reagents for leaching gold

With the growing environmental pressures on the use of cyanide in certain regions and the unsuitability of cyanide for certain complex ores, various reagents have been evaluated. Some of these alternative reagents are summarized in Table 4.

It appears that complexation of copper with ammonia lowers the consumption of CK Similarly, present work on leaching copper-molybdenum sulfides indicates that carbonate also complexes with copper, which depresses the amount of copper leached with CI2. This may have applications to copper-gold ores. Then there are various other processes including the K-process using undisclosed reagents However chlorine and thiourea have received the most attention and have specific applications.

As shown, the advantage of chlorine and thiourea is that they dissolve gold much faster than CN- and sometimes give higher recoveries; but unfortunately greater reagent quantities are often required (Table 5).

The problem with chlorine is that it reacts with sulfide minerals leading to high CI2 consumption and acid production (equation 3).

With thiourea, it is the dimerisation and degradation of thiourea that leads to high consumption (equation 4). With other reagents there is the problem of cost, or stability, or how gold is recovered.


Of all the reagents examined so far, thiourea offers the best prospects for the difficult ores which cannot use cyanide. It is particularly suited to the acidic residues from bacterial leaching or pressure leaching of refractory sulfides because it complexes gold in acid media. A recent examination of a number of Australian oxidised gold ores identified some other promising applications. Furthermore, over the last 3 years there have been developments which could make thiourea even more attractive. Recent work has snown that SO2 inhibits dimerisat’on and degradation of thiourea by controlling the Eh and that a modified thiourea is even more stable. Thus certain types of ore which consume base or CN- could be economically treated with modified thiourea reagents. Furtner research in this area is currently being undertaken.


In the past few years much progress has been made in understanding and developing the C.I.P. process. However, there is still much fundamental worK to be aone ana mprovements to be made with the treatment of refractory ores. No two ores are exactly alike and each needs to be evaluated with a range of options at our disposal.

By understanding the characteristics and properties of carbon, its activity and performance has been improved. Whilst no immediate threat to current C.I.P. practice exists, alternatives to carbon and alternatives to CN- may develop in a few years time.


Solvent extraction and ion exchange

Solvent extraction is established technology for copper and uranium, but the extraction of gold has been hampered by the lack of a selective reagent. In 1984 however, it was reported that mixtures of secondary amnes and TBP were selective for aurocyanide at pH 9. These reagents are widely used in the uranium industry and ODen up the prospect of solvent extracting heap leach eluates or developing solvent-in pulp shows that the solvent mixture is quite selective for gold over copper. Gold is extracted at pH 9 and stripped at pH 11.

Resins, however, have attracted most attent on because they offer high loadings of gold without fouling by organics and are relatively easy to regenerate .Some Russian plants have used resins for years but the main disadvantages are related to their selectivity, stripping and screening. As recently discussed, strong base resins are difficult to strip and require reagents like thiocyanate or thiourea. Weak base resins on the other hand are simply stripped by pH adjustment, but most such resins are not pure and have strong base impurities. Hence they do not strip efficiently when the pH is raised. These resins also load copper zinc and nickel cyanide complexes just as readily as gold and show less discrimination than carbon. Thus the ideal resin has yet to be developed. It is therefore interesting to note that an entirely new approachs being developed which uses ion-exchange fibres woven into cloth. These are polyacrylonitrile fibre with imidazole groups or polybenzimidazole functional groups which represent a new concept in substrate and functional group design.

Cementation of gold onto iron

One of the problems associated with the Anglo elution procedure has been the occasional cementation of gold on the iron pipes and elution vessels. Recently, Kenna [46] has largely elucidated the reasons for this phenomenon, using electrochemical techniques. Initially, it was thought that iron dissolved to form ferrocyanides. In which case, Pourbaix diagrams indicate that geld should cement out over a wide range of pH and potentials. However, Kenna found that cementation occurred only at high temperatures when the pH was allowed to drop below 9 or rise above 14, and was accelerated by the presence of Cl” m the water. The mechanism depends on the corrosion of ircn to Iron oxides. Between pH 9-14 a passive layer of Fe(OHh prevents cementation, but outside this range, soluble iron species exist allowing the iron metal to reduce the gold from solution. Chloride ion interferes with this passive film giving pit corrosion. Thus for most operators, the simple solution is to rubber-line the columns.

Cementation versus eiectrowinning for gold recovery

Because Anglo elution is a batch process a multi-compartment flow-through cell has been designed to treat each batch of eluate. Gold plates onto the steel wool in the normal way but the several banks of steel wool cathodes ensure that the final eluant contains < 1 ppm gold. However although most gold plants in South Africa use these cells, it still takes much time and effort to finally recover the gold from the steel wool. In Australia this problem is being tackled in various ways. Baxter has shown that a pressurised eiectrowinning cell coupled to a pressure Zadra Unit efficiently electrowins gold at 130 degrees leading to high loadings of gold on the steel wool. Currently Costello is examining refining the gold from loaded stainless steel wool by making it the anode of a second cell and plating out gold foil on a steel cathode plate, whilst Biegler is developing a forced circulation cell to produce gold foil directly from dilute eluates.

Meanwhile, in South Africa, zinc cementation is often used to recover gold from batch eluates because it is quick and simple. Recent electrochemical studies have highlighted the important controlling factors and concluded that for this application there was no need for additives of Pb(N03>2 or to de-aerate solutions. Nevertheless a strict range of conditions was still required to avoid passivation or too high consumption of Zn.


Degradation of cyanide

The trend towards using carbon-in-leach, and high temperature elution procedures, has prompted a recent study at Murdoch University on the degradation of cyanide. The extent of the problem is indicated by the 60-70% cyanide loss during the Anglo elution process reported by Davidson. In summary, cyanide can be lost by a variety of mechanisms. For plant operators, cyanide decomposition by oxygen to cyanate and carbonate; and hydrolysis with water to ammonium formate and carbonate, are of most concern. Hydrolysis is important at the high temperatures used for elution.

Initial results from the laboratory show that half the cyanide is decomposed by hydrolysis in 8 hours at 100° whilst at 120 deg the rate is much faster.

Degradation by oxygen is catalysed by carbon and takes place during leaching. As shown the oxidation of cyanide with air and carbon takes place quite rapidly even at 18’C. Half the CN- is lost in 24 hrs; but in the presence of copper the rate is even faster. These results question the desirability of operating a carbon-in-leach circuit and rationalise the significant build-up of CaC03 on carbon as well as evolution of ammonia experienced by plant operators.

Fouling of carbon, regeneration and Elution

In all plant circuits, the carbon activity quickly deteriorates from stage to stage but the reasons why it should do so are complex. It is known that the carbon activity is enhanced by lower pH, higher temperature; the presence of Ca2*, Mg2*, Na*; and by small particle size. On the other hand, its activity is inhibited by excess CN~, build-up of CaCOa and silica, adsorption of xanthates, oils, frothers, huraic acids; and degradatin of active sites [33]. When inorganic foulants are considered, it was found that CaC03 and Mg(OH>2 in particular are adsorbed at high pH, and that copper is adsorbed when the CN~ concentration is low [28]. Silica and iron are invariably present as fine quartz, clay or calcine, which can block the pores of the carbon.

Acid washing with 5-10% HC1 removes many of these foulants but does not restore the carbon activity entirely. As recently shown by La Brooy et al [33] the organic foulants are particularly important, especially the flotation agents, frothers and humic acids. Fortunately most of these get burnt off by heating the carbon, but some require temperatures of 750° to remove them completely.

Carbon regeneration

There is no doubt that the temperature of carbon during reactivation is critical in removing organics and restoring carbon activity and that most plant carbons are not properly regenerated. Reactivation not only removes organics, but also reams out pores, and regenerates surface oxide sites. But as recent studies show, it is important to control temperature, added water/steam, and salts [34]. The key reaction is the water-gas reaction which is rapid above 800° and catalysed by Ca2-, Fe2-. However the type of functional groups produced are also determined by temperature.

Through a proper understanding of key parameters, improved Rotary Kiln designs are evolving, as well as the Electrical Resistive furnace (Rintoul), the Vertical Tubular furnace. Most gold plants currently use rotary kilns with poor control over the carbon temperature. Carbon temperatures vary according to how much water is fed into the furnace with the carbon. It is therefore of interest to compare the various kiln performances.carbon

Elution of gold from carbon

Over the last 10 years there has been much research and development into elution procedures, all trying to improve on the traditional Zadra Process [37] which takes days to complete. Recent studies by Adams and Nicol on the kinetics of elution [38] show that NaCN is more effective than NaOH as an eluant and that high ionic strength solutions retard gold desorption. The optimum NaCN concentration proves to be about 1% w/v

To date, the pressurised Zadra, Anglo, Duval c;nd Micron processes have been commercialised whilst an improved solvent elution system devised at Murdoch University is yet to be developed. All elute the carbon in less than 12 hours using a variety of eluants.

The Anglo process is slowly gaining acceptance over the Zadra process because of its lower overall cost, but the quality of water used is critical to the process [43] which is an important aspect for arid regions. On the other hand the W. Australian Micron process does offer cost advantages in the gold recovery step (see below) and is suited to organically fouled carbons. In South Africa, for example, it is being installed at gold plants which operate a solvent extraction circuit for uranium. The unique advantage of the Micron system (Figure 3) is that less than 1 bed volume of refluxing solvent is used to extract the gold from the carbon charge, and the gold is concentrated in the distillation pot.

The reason why organic solvents are so good at eluting gold is that they increase the reactivity of CN- and stabilise the auro-cyanide in solution.It was the more detailed fundamental knowledge about solvents at Murdoch led to the discovery that acetonitrile elutes gold more efficiently than methanol.

Because the gold is concentrated in the eluant, the Micron process lends itself to elegant gold recovery techniques which give 99.9% pure gold directly. The process selectively discards copper and silver as insoluble cyanides by acidification to pH 3 in the presence of thiourea. The aurocyanide/ thiourea complex is oxidised by chlorine to auric chloride, which is then readily reduced to gold sand by SO2 or sulfites leaving other impurities in solution. Figure 4 shows how a decrease in pH, precipitates out Zn, Ni and Cu cyanides around pH 2-6 leaving gold in solution.

Adsorption of gold onto carbon

Once the gold is leached into solution, there is the problem of selectively concentrating it onto carbon. Again the metallurgists have shown how it works and the scientists are beginning to understand why. Over the last five years the AMIRA Gold Group has been looking at the practical properties of carbon for gold adsorption with postgraduate students looking at some of the fundamental characteristics of carbon.

The essential features of the carbons used today are that they are have and have a graphitic structure with a blend of macro and meso pores. Activated carbon possesses surface oxide sites which ion exchanges with aurocyanide to release OH-, and which also adsorbs cations like Ca2+ and H-.

It is known that more active sites for Au(CN)2- are formed by regenerating carbon at 750“ rather than 550°, and that the carbon potential drops after adsorption of reducible species such as CN- and I-. Carbon also catalyses the air oxidation of CN* into carbonate as discussed below.

Some of the functional groups believed to be formed during activation include carboxylic acid, phenolic, lactone, quinone, hydro-peroxide and chromenol [29] . These groups provide the ion exchange or redox properties of carbon. Aurocyanide is believed to adsorb via ion-exchange with OH” at chromenol sites, Ca2’ adsorb at phenolic sites, whilst CN- is believed to be degraded by peroxide sites.

Significantly the pH of the solution affects both the load ng or au(CN>2′ and associated cations, As shown by Tsuchida less aurocyanide but more cations are loaded at pH < 10. It is believed that adsorption of Ca2* and Mg2* provides excess positive change which allows more aurocyanide to adsorb as an ion-pair.