Gold Leaching Using Thiourea

GOLD LEACHING USING THIOUREA Thiourea, CSN(NH2)2 is an organic compound which dissolves easily I acid solution in a stable molecular form. Gold dissolves in acidic thiourea solution to form a stable complex, 2Au + 4CS(NH2)2 + 2Fe3+ = 2Au(CS(NH2)2)2+ + 2Fe2+ In the thiourea reaction, ferric iron is used as an oxidizing agent, whereas the cyanide process uses oxygen from the air, dissolved in the leach solution. Part of the ferric iron needed is most the time present in the ore. In the case of highly oxidized ores, enough ferric ions will be set free, and then the addition of an oxidant can be reduced. But these ores could content higher amounts of carbonates, which will increase the acid consumption. One could use hydrogen peroxide, sodium peroxide, ozone, potassium permanganate or formamidine disulphide as oxidant; however ferric ion is the most effective. Formamidine is interesting because it can be formed in acid solution by the thiourea oxidation by the presence of an oxidant agent such as ferric iron. Successfully leaching of thiourea depends on an understanding of the role of formamidine which is a compound very active during gold leaching. In the first step thiourea is oxidized to formamidine disulphide, 2CS(NH2)2 + 2Fe3+ = C2S2(NH)2(NH2)2 + 2Fe2+ + 2H+ Oxidation thiourea is reversible. Thus, with a specific reducing agent, formamidine can be converted back into thiourea. In the next step, gold is oxidized by formamidine and forms a cationic gold thiourea complex, 2Au + C2S2(NH)2(NH2)2 + 2CS(NH2)2 + 2H+ = 2(Au(CS(NH2)2)2)+ Formamidine acts as an oxidizer as well as a complexing agent, supplying about 50% of the ligands to the complexation. This explains the higher leaching rates observed with thiourea compared to cyanidation. The general reaction is as follow, 2Au + 4CS(NH2)2 + 2Fe3+ = 2Au(CS(NH2)2)2+ + 2Fe2+ Thiourea must be present in a stoichiometric excess. The ratio of complexing and oxidizing agents must be carefully adjusted. An uncontrolled oxidation of the thiourea solution will lead to unwanted reagent consumption. In a final and irreversible step, formamidine breaks down to cyanamide and elemental sulphur. This forms two effects. First, the elemental sulphur will come in a fine, sticky form, which might passivate the feed ore. Second, loss of silver could occur because of a reaction which leads to precipitation of silver salts. The breakdown can be unpredictable, but it seems that the event is accelerated by high concentrations of formamidine. The corrective measure is to keep the concentration of thiourea itself low and prevent uncontrolled oxidation. High amounts of free thiourea result in rapid leaching, but they are vulnerable for a fast breakdown. The use of acidic solutions at potentials well below those required for the formation of oxide film on the gold surface obviates any possible passivation of the gold by oxide films and extractions in thiourea solutions are high. Of the several oxidants, only ferric ions offer any promise of practical application. The initial rate of dissolution of gold in freshly prepared solutions of thiourea containing ferric ions is high, and is controlled only by the rate of diffusion of the oxidant to the gold surface. However, due to the slow oxidation of thiourea and the partial passivation of the gold surface by the products of oxidation, the rate of dissolution decreases with the age of the lixiviant, and this can lead to excessively high consumption of the reagent. The excess can be more than 5 kg/t. One of the main advantages of thiourea is the high rate of gold dissolution. The leaching rate can be four to five times faster than cyanidation. This could be an important aspect for plants which process ore containing coarse gold particles. The much higher leaching time will reduce the investment for a new processing plant. The thiourea leaching can be adapted to given plant and ore

conditions, whereas the leaching conditions of the cyanide process can only be modified within a narrow range. The economics of gold leaching with thiourea are principally determined by thiourea consumption, which is related to thiourea and oxidant concentrations, pH and the solution potential. Thiourea concentrations between 5 to 50 g/l have been used in lab and pilot test. Sufficient oxidant (i.e. ferric ion) is required to oxidize thiourea to formamidine disulphide for optimal leaching conditions. The presence of excess oxidant increases thiourea consumption significantly. For this reason, close control of solution potential would be required through all stages of leaching in any commercial process. Thiourea consumptions of 1-4 kg/t have been projected for optimized thiourea leaching systems based on currently available technology, although estimates as high as 10 kg/t could be obtained. Such high consumptions, coupled with the requirements of sulphuric acid for pH control, and agents for potential control, trends to be a part very important of the cost, perhaps between 1.5 to 2.1 times the cost of a cyanidation process. The advantage of thiourea is that sulphuric acid has little effect on sulphidic copper minerals which dissolve almost completely in cyanide solutions. Obviously, copper carbonates will be dissolved easily. In general, the observed level of copper in thiourea leach solutions in lower than those in cyanide leaches. Arsenic and antimony present in the form of sulphidic minerals can result in tremendous difficulties for cyanidation. This is true not so much for arsenopyrite, but more for realgar, orpiment and stibnite. These minerals dissolve in alkaline solutions at the high ph level used in cyanidation, resulting in the formation of thioarsenites and thioantimonites. These compounds react with oxygen and then less oxygen is available. Refractory ores of this type are more suitable for thiourea. Arsenic and antimony sulphides will not dissolve at pH range of 1 to 2. This is demonstrated at the New England Antimony Mine. The New England Antimony Mine is located in New South Wales (Australia). The plant leached a refractory Aurostibnite flotation concentrate using thiourea. About 50-60% of the gold content is extracted in 10-15 minutes from the flotation concentrate after as much free gold as possible had recovered by gravity separation. Total recovery is about 80%. Critical parameters for the thiourea leaching were:     

pH, 1.4 adjusted with sulphuric acid. Redox potential, 150 to 250 mV. Thiourea concentration, 1%. Thiourea consumption, 2 kg/t. Leach time, 10-15 minutes.

Over 250 mV, thiourea consumptions increase to excessive levels. Below 150 mV, gold is not leached in thiourea. The potential was initially controlled by adding peroxide, but later was used MnO 2. About 6000 gm/m3 ferric iron is needed for the process to work. Thiourea consumption is reduced by keeping the time of thiourea contact with the sulphide mineral to an absolute minimum. For this reason, leaching time is kept to 10-15 minutes, and the pulp is filtered as rapidly as possible. After leaching, gold is recovered by adding carbon to the pregnant solution, and allowing loaded carbon to settle before decanting the barren solution to recover thiourea values for reuse. Gold thiourea can be recovered from solutions or slurries by adsorption on activated carbon or ion exchange resins, or it can be precipitated on steel wool or lead metal. Approximately, gold thiourea can be loaded in carbon to 10-20 kg/t gold.

GOLD LEACHING WITH ACIDIC THIOCYANATE The oxidation of gold in acidic thiocyanate solutions can be expected to produce either the gold (I) complex Au(SCN)2- or the gold (III) complex Au(SCN)4-. Au(SCN)2- + e = Au + 2SCN-, Au(SCN)4- + 3e = Au + 4SCN-,

ORP = 662 mV ORP = 636 mV

The potentials required for these reactions are below those for the reduction of oxygen and ferric ions, and these reagents can be expected to be effective oxidants for gold in thiocyanate solutions; however, ferric iron is the most suitable oxidant for the reaction since the kinetics of dissolution are slow if oxygen is used and thiocyanate is oxidized rapidly by hydrogen peroxide. The stability of the thiocyanate ion is increased in the presence of ferric ions due to the many complexes it forms with these ions. The standard potential for the reduction of oxygen is 700 mV and this value is higher than the required for the oxidation of gold in thiocyanate solutions, the irreversible nature of the oxygen reduction reaction results in insignificantly slow leaching under normal conditions. The use of elevated temperatures and pressures could be expected to produce acceptable leaching rates, but the decomposition of thiocyanate by oxidation to thiocyanogen is a possible complication. Employing ferric ion as oxidant, thiocyanate forms relatively strong complexes with ferric ions, which reduces both the oxidizing potential of ferric ions and the concentration of free thiocyanate needed for the formation of complexes with gold. Fe(SCN)4 + Au = Fe2+ + Au(SCN)2- + 2SCNThe optimum pH range is 2.0-3.0. The stability of thiocyanate is potential dependent, with stability achieved below approximately 640 mV. Also, higher potentials are required to achieve satisfactory gold leaching rates at practical thiocyanate concentrations. There is a real relationship between these two requirements. The rate of gold dissolution increases with increasing thiocyanate and to a lesser extent ferric ion concentration. Concentrations of 10-15 g/l and 2-10 g/l for the two species have been used. Increasing temperature increases the rate of thiocyanate consumptions and, in view of the high reagent consumption even at ambient temperatures, elevating pulp temperatures is not a good option. A significant drop in gold extraction has been obtained at temperatures above 40oC when using thiocyanate. Silver trends to form an almost insoluble product, silver thiocyanate and is unrecoverable by this process.

LEACHING GOLD WITH BROMINE Bromine has been used in the first years of the last century and can be considered an option when there is an auriferous ore difficult to treat by usual methods. Bromine can be added in several forms. One possibility is to add a bromide salt with chlorine or hypochlorine which converted bromide to bromine. The latter leaches gold as well as platinum at any pH and this is really advantageous. Other possibility is to employ bromocyanide. By using bromine there some advantages such as easy extraction, and adaptability to several pH values. There disadvantages such as high consumption, interference with assays by atomic absorption. Au + 2Br- + Br2 = AuBr4- + e Years ago, bromocyanide was used at the Decloro Mine, Ontario (Canada) on mispickel ore. The process consisted of grinding exceedingly fine, and then agitation with cyanide solution to which bromide of cyanogen was added to intervals. During the treatment protective alkalinity is kept at the lowest possible point due to the instability of the reagent, lime sufficient for settlement being added after the treatment is finished. The quantities of free cyanide and bromocyanide are varied according to the assay value of the ore; any additional bromocyanide needed being added at intervals of several hours. KCN + Br2 = KBr + BrCN BrCN + 2KOH = KBr + KCNO + H2O BrCN + 3KCN + 2Au = 2KAu(CN)2 + KBr The quantity of BrCN may be determined in a cyanide solution by acidifying with hydrochloric acid, adding excess of iodide and titrating with thiosulphate. BrCN is decomposed by alkali, then is very important observe the free alkalinity. Once finalized the process, can be added enough amount of lime for settling. Other option to be considered in this leaching type is to employ chlorine to generate bromine from its salts. The free bromine is converted to bromide ions by reaction with native gold as well as by other reactions with gangue species.

GOLD LEACHING WITH THIOSULPHATE The oxidation of gold in the presence of thiosulphate ions can be expected to produce the gold complex Au(S2O3)23-, but because thiosulphate does not work in acid solutions, alkaline solutions must be used. Few oxidants are suitable for use in alkaline solutions. Although oxygen is an effective reagent for cyanidation, it is not electrochemically active at the required potential for thiosulphate leaching. The reaction is; 4Au + 8S2O32- + O2 + 2H2O = 4Au(S2O3)23- + 4OHThe rate of dissolution is dependent on thiosulphate and dissolved oxygen concentrations and temperature. Leaching rate is improved by copper in ammonium solution. 4Cu(S2O3)23- + 16NH3 + O2 + 2H2O = 4Cu(NH3)42+ + 8S2O32- + 4OH4Au + 4S2O32- + Cu(NH3)42+ = 4Au(S2O3)23- + 4NH3 + Cu(S2O3)23Silver chloride and silver sulphide are easily leached. The thiosulphate compounds are consumed by several oxidations and association reactions. 5S2O32- + S2O62- + 3H2O = 4S4O62- + 6OHThiosulphate can be stabilized by the addition of small amounts of sulphite ions which react with sulphide sulphur and regenerate thiosulphate. This prevents the precipitation of silver as the insoluble sulphide. In the anodic area gold is oxidized and complexed with ammonia. This complex can be replaced by a more stable gold thiosulphate complex. In the cathodic are, the cupric amine complex is reduced and the oxygen present in the ammonia solution oxidizes the cuprous complex to a cupric compound. Ammonia and copper are recycled in the system. It has been reported that the thiosulphate consumption is very high such as 29 kg/t, but it can be reduced by adding reducing agents as chelates which can dive a consumption of 13 kg/t. obviously, there are certain factors that influence thiosulphate consumption. Some of these factors are: sulphide minerals present in the ore react with thiosulphate; oxidant compounds trends to consume more thiosulphate, solid percentage. Thiosulphate does not present a good stability when is exposed to the presence of ultraviolet light, to the presence of heavy metals, or if exist a big amount of cations in the water. Gold recoveries by leaching can be higher than 90%, but as was mentioned with a high consumption of thiosulphate. The pH can be in the range 10.0 to 10.5 and the pulp density can be in the range 40 to 45% solids. Pregnant solutions obtained with thiosulphate don’t give a good recovery with activated carbon or resins. The best results have obtained by cementation with zinc, copper or iron.

LEACH GOLD BY CHLORINATION Chlorination was used two centuries ago before the use of cyanidation. Chlorine was first used to recover gold from residues from amalgamation. Later, was used in big operations in the American and Australian gold fields. At the present time, chlorination for gold recovery is used on extremely small scale where gold is a minor constituent with other precious metals. The most important example is the treatment of matte leach residues to recover the platinum group metals where gold is a by-product. Some companies are Consolidated Murchison, Matthey Rustenburg Refineries, and Brimsdown. Gold dissolves in aqueous chloride solution to form gold+ and gold3+ chloride complexes. Au + 2Cl- = AuCl2 - +e Au + 4Cl- = AuCl4 - + 3e

It’s possible that a solution of auric chloride can be decomposed by many reducing agents such organic compounds, metals, charcoal, and ether. The presence of hydrochloric acid trends to hinder the deposition of metal. Nowadays, the action of carbon can be expressed according to the following reaction: AuCl4 - + 3e = Au + 4Cl- ORP = 1000 mV (1MCl-) From pure solutions of AuCl4- in very dilute acid, it is possible to obtain gold loadings on activated carbon of about 60% by mass, but the gold could be lightly attached and fragments are easily dislodged. Silver and lead form insoluble chlorides in chloride media. This is important because insoluble products can reduce the solubility of gold due to the formation of an insoluble layer. Nevertheless, passivation can occur to any important degree when the silver and lead content of the gold alloy is more than 12%. Copper and zinc form relatively unstable chloro-complexes. All these compounds are less stable than auric chloride. It is important the large relatively differences in the stability constants of the copper cyanide chloride and gold complexes; it means that less copper can be dissolved in chloride media than in cyanide. Gold telluride minerals are soluble in acid chloride media when there is an oxidant such as iron ferric or chlorine, and dissolve to form some complexed compounds of gold and tellurium. The behaviour of carbonates is interesting because the decomposition of these minerals help to recover gold because there is a better exposition of locked gold. The process has been applied at Emperor (Fiji) that treats a gold-tellurium ore does not easily to cyanidation. Other plant that employed this process was Antimony Products, Pty. Ltd. (South Africa) for the recovery of gold from antimony rich slag that was stockpiled for a long time. The process consisted in getting a gold concentrate by flotation which was leached in chloride media. Antimony was separated by hydrolysis and the liquor was treated with activated carbon for recovering the precious metals. The gold content in the stockpile was 85 g/t Au with 9.9% Sb, an the flotation concentrate reported 935 g/t Au and 31.6% Sb. Gold recoveries were more than 80%. The treatment has a bottleneck in the filtration step because the rate was low and special care was considered during this stage.

Gold Leaching With IODINE Used in the past for gold dissolution, iodine trends to forms the most stable gold complexes of all halogens. As a result, the redox potential at which gold dissolves in an iodine solution is about half that at which it dissolves in acidified chlorine. The redox potential at which gold dissolves in iodine is, however, considerably higher than in cyanide. Iodine leaches gold over a wide pH range. See the next table. Leaching Agent

Redox Potential, Volts

Cl-

+1.15

Br-

+0.96

I-

+0.56

Thiourea

+0.36

CN-

-0.61

Iodine is an element very expensive, it cost about $7 to10/lb, but it can be regenerated from the electrolyte. The high cost demands the efficiency recovery and recycling of iodine. Less pure grades of iodine are said to be available at around $1 to 3/lb. Iodine does not absorbs to any great extent on gangue particles, so good recoveries of reagent is possible.

Iodine at very low concentrations leaches gold. If the ore or pulp has to be washed with a lot of water to recover the lixiviant, iodine can be recovered from the resulting dilute solution using a simple ion exchange technique. When recovered in that manner, iodine is readily eluted by hot water, with leaves it in a recyclable form. When iodine attacks a gold ore where there is pyrite or other reducing agents, gold is solubilized as iodize complex AuI2- and AuI4-, and pyrite and other minerals are oxidized by iodine forming iodide. For an economical process iodine must be recycled, thus iodine must return to its oxidizing state and the gold have to be removed. These requirements are difficult and expensive. The difficulty is on the fact that iodine and gold-complex behave very similar in the presence of reducing materials such as resins or activated carbon which are employed to recover gold from liquors. Ergo, gold cementation with iron or zinc dust results in solubilization of big amounts of these elements because iodine and gold will be oxidized. Recovery by resins or activated carbon has a problem, gold and iodine competes for the same places at the same time. In this process gold is leaches by iodine which is complexed by iodide. The oxidation of sulphide minerals produces iodide as a reaction product. The latter is combined with iodine favoring its dissolution because iodine has a low solubility (0.3 g/l). Oxidants are no necessary to be added. Iodine concentration must be in the range 2 to 10 gr/l. The reactions are: FeS2 +I2 = Fe2+ + 2S + 2IAu + I- + I2 = AuI2Iodine penetrates gold particles particularly well and, above pH 8, it does no complex iron and attacks sulphides weakly. Below pH 7, iodine does attack sulphides, at which time is converted in iodine ions rather than combining with sulphur. Iodine can be regenerated to I 2 electrolytically in a diaphragm cell. During this electrolytical process iodide is reduce to iodine in the anode and gold is deposited in the cathode. Thus, iodine can be recycled and gold can be recovered. Leaching Agent

Redox Potential, Volts

Cl-

+1.15

Br-

+0.96

I-

+0.56

Thiourea

+0.36

CN-

-0,61