The leaching of chalcopyrite with cupric chloride

A comparative study of electrochemical leaching and chemical leaching of chalcopyrite was performed mainly at 343 K to elucidate the leaching mechanism of chalcopyrite with CuCl₂. Also, the morphology of the leached chalcopyrite surface was studied by using a single chalcopyrite crystal. The leaching with CuCl₂ produced a porous elemental sulfur layer on the chalcopyrite surface, showing a similar morphology to that produced during leaching with FeCl₃. The leaching kinetics were found to be linear over an extended period, followed by an acceleration stage, as a result of an increase in the reaction surface area. The leaching rate of chalcopyrite was proportional to C(CuCl₂)^0.5, whereas it was inversely proportional to C(CuCl)^0.5. The mixed potential of chalcopyrite exhibited a 66 mV decade⁻¹ dependency upon C(CuCl₂), and—69 mV decade⁻¹ upon C(CuCl). Based on these observations together with other findings, an electrochemical mechanism involving the oxidation of chalcopyrite and CuCl ₋ ² and the reduction of CuCl⁺ was proposed. The Tafel plot between the mixed potential and the current density obtained by converting the rate of chemical leaching gave a straight line whose slope was in good agreement with that of the electrochemical leaching. These findings strongly support the electrochemical mechanism of chalcopyrite leaching with cupric chloride.     

The leaching of sintered CuS disks with ferric chlorides

A comparative study of chemical and electrochemical leaching of a sintered disk of CuS was carried out at 343 to elucidate the leaching mechanism of CuS with FeCl₃. The leaching,rate of CuS exhibited a half order dependence on the FeCl₃ concentration, but the addition of FeCl₂ slightly reduced the leaching rate. The mixed potential of CuS exhibited a 53 mV decade^-1 dependence upon the FeCl₃ concentration; no significant effect on the mixed potential was observed by the addition of FeCl₂. Based on these observations, an electrochemical mechanism is proposed which involves the oxidation of CuS and the reduction of FeCl₂ ⁺ . The activation energy in chemical leaching of CuS was 46.7 kJ mol^-1, while in electrochemical leaching, it was 50.4 kJ mol^-1. By converting the leaching rate to electric current density,i, a 110 mV decade^-1 dependence of mixed potential, E, against log i is obtained. This dependence of the mixed potential during chemical leaching of CuS on FeCl₃ concentration agrees with the value theoretically expected from the electrochemical model of the leaching process. These findings strongly support an electrochemical mechanism for the FeCl₃ leaching of CuS. HIROSHI HIAI, formerly Graduate Student with Kyoto University     

The leaching of chalcopyrite with ferric sulfate

The leaching kinetics of natural chalcopyrite crystals with ferric sulfate was studied. The morphology of the leached chalcopyrite and the electrochemical properties of chalcopyrite electrodes also were investigated. The leaching of chalcopyrite showed parabolic-like kinetics initially and then showed linear kinetics. In the initial stage, a dense sulfur layer formed on the chalcopyrite surface. The growth of the layer caused it to peel from the surface, leaving a rough surface. In the linear stage, no thick sulfur layer was observed. In this investigation, chalcopyrite leaching in the linear stage was principally studied. The apparent activation energy for chalcopyrite leaching was found to range from 76.8 to 87.7 kJ mol⁻¹, and this suggests that the leaching of chalcopyrite is chemically controlled. The leaching rate of chalcopyrite increases with an increase in Fe(SO₄)_1.5 concentration up to 0.1 mol dm⁻³, but a further increase of the Fe(SO₄)_1.5 concentration has little effect on the leaching rate. The dependency of the mixed potential upon Fe(SO₄)_1.5 concentration was found to be 79 mV decade⁻¹ from 0.01 mol dm⁻³ to 1 mol dm⁻³ Fe(SO₄)_1.5. Both the leaching rate and the mixed potential decreased with an increased FeSO₄ concentration. The anodic current of Fe(II) oxidation on the chalcopyrite surface in a sulfate medium was larger than that in a chloride medium.     

Oxidative leaching of an offgrade/complex copper concentrate in chloride lixiviants

Laboratory studies have been conducted on chloride leaching as a possible route for the simultaneous recovery of copper, zinc, and lead from an off grade and complex chalcopyrite concentrate (from Sikkim, India) associated with appreciable amounts of sphalerite, galena, and pyrite. The effects of temperature, concentration, and quantity of ferric chloride, stirring speed, and leaching time on metal dissolution have been investigated. Leaching tests have also been conducted with in-dividual (HC1, NaCl, CuCl₂, FeCl₃) and mixed chlorides (two-, three-, and four-component mix-tures). Results show the possibility of recovering not only 99 pct Cu and 89 pct Zn but also 82 pct Pb and 58 pct elemental S by treatment of the concentrate with 4 M FeCl₃ at 383 K (110 °C) for 7.2 ks (2 hours) employing 25 pct excess FeCl₃ and a stirring speed of 700 rev min⁻¹. Though 64 pct iron of the concentrate is found to dissolve, the pyrite seems to remain unattacked. Kinetic studies indicate that the chalcopyrite, sphalerite, and galena of the concen-trate dissolve simultaneously in the FeCl₃ lixiviant as if each mineral is separately leached, and the Cu and Zn dissolution reactions are under chemical control (linear kinetics). The addition of NaCl to the chloride lixiviants is found to be beneficial only up to a common salt concen-tration of 100 g/l. Leaching of the copper concentrate with CuCl₂ or mixed FeCl₃-CuCl₂-NaCl has not been as effective as its direct leaching with 4 M FeCl₃. N.V. NGOC, formerly Visiting Scientist with the Department of Metallurgical Engineering, Institute of Technology, Banaras Hindu University.     

Leaching of sulfidized chalcopyrite with H<Subscript>2</Subscript>SO<Subscript>4</Subscript>-NaCl-O<Subscript>2</Subscript>

The recovery of copper from chalcopyrite by leaching is complex not only due to the slow dissolution kinetics of this mineral in most aqueous media but also due to the production of solutions that are heavily contaminated with iron. On the contrary, the leaching of sulfidized chalcopyrite is very attractive because of a faster and more selective dissolution of copper compared to the leaching of the untreated chalcopyrite. In this work, the results of leaching in H₂SO₄-NaCl-O₂ solutions of sulfidized chalcopyrite concentrate are discussed. Experiments were carried out with chalcopyrite concentrates previously reacted with elemental sulfur at 375 °C for 60 minutes. The results showed that the concentration of chloride ions below 0.5 M, temperature, and leaching time are important variables for the extraction of Cu. On the other hand, Fe extraction was little affected by the same variables, remaining below 6 pct for all the experimental conditions tested. Microscopic observations of the leached particles showed that the elemental sulfur produced by the reaction does not form a coherent layer surrounding the particle, but rather concentrates in certain locations as large clusters. The leaching kinetics can be accurately described by a nonreactive core-shrinking rim topochemical expression for spherical particles 1 − (1 − 0.45X)^1/3=kt. The activation energy found was 76 kJ/mol for the range 85 °C to 100 °C.     

The leaching kinetics of chalcopyrite (CuFeS2) in ammonium lodide solutions with iodine

The leaching kinetics of chalcopyrite (CuFeS₂) in ammonium iodide solutions with iodine has been studied using the rotating disc method. The variables studied include the concentrations of lixiviants, rotation speed, pH of the solution, reaction temperature, and reaction product layer. The leaching rate was found to be independent of the disc rotating speed. The apparent activation energy was measured to be about 50 kJ/mole from 16 °C to 35 °C, and 30.3 kJ/mole from 35 °C to 60 °C. The experimental findings were described by an electrochemical reaction-controlled kinetic model: rate =k [NH₃]^0.69[OH⁻]^0.42[I ₃ ⁻ ]^0.5.     

A mathematical model for calculation of equilibrium solution speciations for the FeCl3-FeCl2-CuCl2-CuCl-HCl-NaCl-H2O system at 25 ‡C

Equilibrium solution speciation computations were performed for the FeCl_-FeCl3-CuCl₂-CuCl-HCl-NaCl-H₂O system at 25 ‡C. In dilute solutions, complexation of Fe(III), Fe(II), and Cu(II) is insignificant but the major Cu(I) species is CuCl₂ ^-. In concentrated solutions, FeCl ₃ ⁰ , FeCl ₂ ⁰ , and CuCl ₂ ⁰ are the major Fe(III), Fe(II), and Cu(II) species, and CuCl ₃ ^2- is the most important cuprous complex. High Cu(I)/Cu(II) ratios are apparently more readily attainable in CuCl₂ than in FeCl₃ media. The Cu(I)/Cu(II) ratio is increased by making the solution more concentrated in any component except FeCl₃ or CuCl₂. Neither the ionic strength nor the total chloride concentration is a good predictor of the Cu(I)/Cu(II) ratio.     

Calculation of phase diagrams for the FeCl2, PbCl2, and ZnCl2 binary systems by using molecular dynamics simulation

Recently, molecular dynamics (MD) simulation has been widely employed as a very useful method for the calculation of various physicochemical properties in the molten slags and fluxes. In this study, MD simulation has been applied to calculate the structural, transport, and thermodynamic properties for the FeCl₂, PbCl₂, and ZnCl₂ systems using the Born—Mayer—Huggins type pairwise potential with partial ionic charges. The interatomic potential parameters were determined by fitting the physicochemical properties of iron chloride, lead chloride, and zinc chloride systems with experimentally measured results. The calculated structural, transport, and thermodynamic properties of pure FeCl₂, PbCl₂, and ZnCl₂ showed the same tendency with observed results. Especially, the calculated structural properties of molten ZnCl₂ and FeCl₂ show the possibility of formation of polymeric network structures based on the ionic complexes of ZnCl ₄ ²⁻ , ZnCl ₃ ⁻ , FeCl ₄ ²⁻ , and FeCl ₃ ⁻ , and these calculations have successfully reproduced the measured results. The enthalpy, entropy, and Gibbs energy of mixing for the PbCl₂-ZnCl₂, FeCl₂-PbCl₂, and FeCl₂-ZnCl₂ systems were calculated based on the thermodynamic and structural parameters of each binary system obtained from MD simulation. The phase diagrams of the PbCl₂-ZnCl₂, FeCl₂-PbCl₂, and FeCl₂-ZnCl₂ systems estimated by using the calculated Gibbs energy of mixing reproduced the experimentally measured ones reasonably well.     

Chalcopyrite leaching by CuCl2 in strong NaCI solutions

Chalcopyrite oxidation by CuCl₂ in NaCl solutions has been studied as a function of leaching time, temperature, grain size, Cu(II)/Cu(I) ratio, and Cl⁻ activity. The experiments have been performed under conditions where the solution potential remains almost constant in each test. Leaching rates are high above 85 °C, and high copper recovery can be obtained provided that high Cu(II)/Cu(I) ratios are maintained. Elemental sulfur is obtained with no sulfate formation. It was observed that the combined influence of Cu(II)/Cu(I) ratio and Cl⁻ activity on the leaching rate is related to the solution potential, explaining the action of the concentration of chloride ion in chalcopyrite leaching. M.BONAN, formerly with Ecole des Mines, is now with Israel Mining Industries.     

Kinetics of the sulfidation of chalcopyrite with gaseous sulfur

The sulfidation of chalcopyrite with gaseous sulfur in the temperature range of 325 °C to 400 °C occurs with the formation of covellite and pyrite as the final products. The rate of sulfidation depends strongly on the temperature, with nearly complete conversion in less than 30 minutes at 400 °C. Microscopic analysis of partially and completely reacted particles showed that the sulfidation proceeded topochemically, with a shrinking core of unreacted chalcopyrite surrounded by successive layers of FeS₂ and CuS. The experimental data exhibited an induction period at the beginning of the reaction. An electrochemical mechanism is proposed for the sulfidation reaction, which involves simultaneous diffusion of Cu⁻ and electrons through the product layers. The rate data showed that the fraction reacted is well represented by a shrinking-core model controlled by the reaction occurring at the chalcopyrite-pyrite interface, resulting in the conversion-vs-time relationship 1−(1−X)^1/3=k(t−t _ind). An activation energy of 98.4 kJ/mol was determined for the temperature range of 325 °C to 400 °C.     

Reaction mechanism of gold dissolution with bromine

The dissolution of gold with elemental bromine was studied by using a rotating disc technique. The main parameters studied were bromine and bromide concentrations, stirring speed, pH, and temperature. The effect of various salts, manganese, and hydrogen peroxide was also examined. The dissolution kinetics of gold with Br₂ and NaBr mixture is complex. The reaction mechanism is a function of solution composition, which determines the kind of adsorbing species. For an excess concentration of bromide ions, the rate expression is Rate = (2k _cl7 k _al6)^1/2 K ₁₅ [Br ₃ ⁻ ] and for an excess concentration of bromine, the rate expression is Rate = (2k _c27 k _a29)^1/2 [Br⁻]^1/2 {K₂₅ [Br₂]³/(1 +K ₂₅ [Br₂]³)}^1/2 Gold in bromine solutions dissolves according to electrochemical/chemical (EC) mechanisms. The electrochemical component of the mechanism is responsible for the formation of AuBr₂. In the chemical component of the mechanism, this monovalent gold bromide disproportionates into gold and stable AuBr ₄ ⁻ , which reports into solution. With respect to pH, there are two characteristic dissolution regions. In the pH range of 1 to 7, gold dissolution rates were insensitive to pH. Above pH 7, gold dissolution rates decreased with increase of pH.     

The development of a flow injection analysis method for the quantification of free cyanide and copper cyanide complexes in gold leaching solutions

The quantification of free cyanide, Cu(CN)₂⁻ and Cu(CN)₃²⁻, is essential in investigating the leaching of copper–gold ores in cyanide solutions. A flow injection analysis (FIA) method has been developed for this purpose which utilizes a flow-through electrochemical cell containing silver and platinum electrodes. The basis of this method is that the measured charge due to either silver oxidation or copper reduction at an applied potential is related to the species concentration. A potential of −150 mV was chosen for free cyanide measurement, at which the measured charge due to silver oxidation is related to the free cyanide concentration. While silver oxidation at 100 mV and copper reduction at −650 mV were used for Cu(CN)₃²⁻ and Cu(CN)₂⁻ analysis, respectively. For the solution containing both free cyanide and Cu(CN)₃²⁻, a two-step analysis technique was developed for the quantification of each of these species. A pretreatment method was also developed to remove the interference of sulfide ions. The versatility of the FIA method developed in this study is demonstrated by measuring the cyanide species formed during the leaching of both Cu₂O and Cu₂S in cyanide solutions.     

Leaching kinetics of digenite concentrate in oxygenated chloride media at ambient pressure

The leaching of digenite concentrate in CuCl₂-HCl-NaCl oxygenated solutions is very rapid. From the effect of variables on the leaching rate and measurements of the concentrations of cuprous and cupric species in the solution as a function of time, it was concluded that the leaching in O₂ atmosphere proceeds by the attack of cupric ions on the copper sulfides to produce cuprous ions which are subsequently oxidized to cupric by the O₂ present in the system. The kinetic study showed that the leaching proceeds in two sequential stages. In the first stage, the digenite is transformed to covellite, and in the second stage, the covellite is dissolved to copper and elemental sulfur. In the first stage, the fraction of copper extracted varied linearly with time according to α=k _l t, whereas in the second stage, the dissolution of covellite was well represented by a shrinking core model controlled by diffusion through a porous product layer kinetic equation: 1−2/3α _cv−(1−α _cv)^2/3=k _cv t. The calculated activation energies were 15.8 and 80.0 kJ/mol for the first and second stages, respectively. These results were explained by an electrochemical mechanism of digenite dissolution.     

Leaching of CuFeS2 by aqueous FeCl3, HCl,and NaCl: Effects of solution composition and limited oxidant

Batch leaching experiments were performed in which the initial amounts of chalcopyrite and ferric chloride were selected to ensure that the oxidant was significantly depleted over the course of an experiment. Solution samples were analyzed for Cu(II) and Fe(III) by visible spectrophotometry and for total copper and total iron by atomic absorption, making it possible to measure changes in the solution component concentrations as leaching progressed. For selected samples, the solution potential was also measured. In all experiments, the Cu(II) concentration passed through a maximum and, simultaneously, the Cu(I) concentration increased very sharply. An acceleration in the total rate of leaching was normally observed at the same time. Early in a leach, the solution potential was too high for the reduction of Cu(II) to Cu(I) to take place at the time of the increase in the overall leaching rate, however, the solution potential dropped sharply during a span of a few hours, reaching a value low enough that reduction of cupric ion became possible. The amount of Cu(I) present at the completion of a leach was dependent on the total chloride concentration of the system. The highest Cu(I)/Cu ratios were observed in systems with the highest chloride concentrations. The ultimate extent of CuFeS₂ leaching was dependent on the initial FeCl₃ and total chloride concentrations; the FeCl₃ was virtually completely consumed and the total chloride concentration controlled the extent to which Cu(II) was reduced by reaction with chalcopyrite.     

Effect of solution composition on the optimum redox potential for chalcopyrite leaching in sulfuric acid solutions

The leaching rate of chalcopyrite (CuFeS₂) by Fe³⁺ in H₂SO₄ solutions depends on the redox potential determined by the Fe³⁺/Fe²⁺ concentration ratio, and there is a maximum leaching rate at an optimum redox potential. The present study investigated the effects of solution composition on the optimum redox potential by electrochemical measurements using a CuFeS₂ electrode and electrolyte solutions containing 0.01–1 kmol m^− 3 of H₂SO₄, Fe²⁺, and Cu²⁺ at 298 K in nitrogen.Anodic-polarization curves of the CuFeS₂ electrode showed that there was a current peak on the curves in the presence of Cu²⁺ and Fe²⁺, corresponding to the maximum leaching rate. The redox potential of the peak increased markedly with increasing Cu²⁺ concentration, while it was little affected by the H₂SO₄ and Fe²⁺ concentrations. These results agree with the results of leaching experiments reported previously, and indicate that the optimum redox potential for chalcopyrite leaching is a function of the Cu²⁺ concentration. An empirical equation for the optimum redox potential for CuFeS₂ leaching is proposed.     

Leaching of chalcopyrite with ferric ion. Part III: Effect of redox potential on the silver-catalyzed process

This study examines the effect of redox potential on silver-catalyzed chalcopyrite leaching. Leaching tests were carried out in stirred Erlenmeyer flasks with 0.5 g chalcopyrite mineral, 1 g Ag/kg Cu and 100 mL of a sulphate solution of Fe³⁺/Fe²⁺ (with redox potential ranging between 300 and 600 mV Ag/AgCl) at pH 1.8, 180 rpm and 35°C or 68 °C. Unlike uncatalyzed leaching, an increase of the redox potential increased copper dissolution in the presence of silver ions, as the regeneration of Ag⁺ requires a high concentration of oxidizing agent, Fe³⁺. Additionally, the high reactivity of the mineral surface when silver was present could have been responsible for inhibiting the nucleation of hydrolysis products of Fe³⁺ on it. Excessive addition of silver transformed the chalcopyrite surface into copper-rich sulphides such as covellite, CuS, and geerite, Cu₈S₅, preventing the formation of CuFeS₂/Ag₂S galvanic couple and the recycling of silver ions.     

Kinetic and Mechanism Studies Using Shrinking Core Model for Copper Leaching from Chalcopyrite in Methanesulfonic Acid with Hydrogen Peroxide

ABSTRACT

An investigation on copper leaching from a chalcopyrite concentrate in methanesulfonic acid (MSA) and hydrogen peroxide at 75°C was carried out. Periodic additions of H₂O₂ were applied to enhance chalcopyrite dissolution and the reaction mechanism was analyzed using a shrinking core model. The results revealed that compared with the addition of H₂O₂ at the very beginning, periodic additions of H₂O₂ enhanced copper extraction and leaching kinetics, and the rate-determining step shifted from the diffusion of oxidant to the surface chemical reaction. The reaction orders with respect to MSA and H₂O₂ were found to be 0.19 and 1.26, respectively, suggesting the leaching process was highly dependent on H₂O₂ concentration. Calculated activation energy between the temperature range of 25–75°C was 79.8 kJ/mol. Detailed study also indicated the reaction mechanism is diffusion-controlled through a protective sulfur layer at lower temperature and surface chemical reaction-controlled at temperatures higher than 55°C. Overall, the MSA–H₂O₂ leaching system is a green method for chalcopyrite leaching and a possible flowsheet in the industrial application with the periodic additions of H₂O₂.     

The dissolution of chalcopyrite in ferric sulfate and ferric chloride media

The literature on the ferric ion leaching of chalcopyrite has been surveyed to identify those leaching parameters which are well established and to outline areas requiring additional study. New experimental work was undertaken to resolve points still in dispute. It seems well established that chalcopyrite dissolution in either ferric chloride or ferric sulfate media is independent of stirring speeds above those necessary to suspend the particles and of acid concentrations above those required to keep iron in solution. The rates are faster in the chloride system and the activation energy in that medium is about 42 kJ/mol; the activation energy is about 75 kJ/mol in ferric sulfate solutions. It has been confirmed that the rate is directly proportional to the surface area of the chalcopyrite in both chloride and sulfate media. Sulfate concentrations, especially FeSO₄ concentrations, decrease the leaching rate substantially; furthermore, CuSO₄ does not promote leaching in the sulfate system. Chloride additions to sulfate solutions accelerate slightly the dissolution rates at elevated temperatures. It has been confirmed that leaching in the ferric sulfate system is nearly independent of the concentration of Fe³⁺, ka[Fe³⁺]^0.12. In ferric chloride solutions, the ferric concentration dependence is greater and appears to be independent of temperature over the interval 45 to 100 °C.     

Henrian activity coefficient of As in Cu-Fe mattes and white metal

A transpiration method was used to evaluate the Henrian activity coefficient of As (γ _As ^o ) in Cu-Fe mattes and white metal. Values for the activity coefficient of As (γ _As) have been evaluated as a function of the Cu/Fe molar ratio from 1 to ∞, as a function of the sulfur deficiency (defined as SD=X _s−1/2X _Cu−X _Fe, where X _i is the mole fraction of the ith species) from −0.02 to +0.02 and at temperatures between 1493 and 1573 K. The activity coefficient for arsenic in the matte was found to have a weak dependence on both temperature and the Cu/Fe molar ratio, but a strong dependence on SD. Analysis of γ _As as a function of the trace-element concentration reveals that the activity coefficient is highly dependent on the As content, even at trace concentrations where Henrian behavior is expected. That dependency is attributed to uncertainty in the reported value of the saturation pressure of monatomic arsenic (P _As ^o ) and highlights problems in comparing results and specifying Henrian values for the activity coefficient. A method is presented whereby the impact of P _As ^o on computed values of γ _As is significantly reduced to obtain an approximate Henrian value of the activity coefficient.     

Direct versus indirect bioleaching

The dissolution of metal sulfides is controlled by their solubility product and thus, the [H⁺] concentration of the solution, and further enhanced by several chemical mechanisms which lead to a disruption of sulfide chemical bonds. They include extraction of electrons and bond breaking by [Fe³⁺], extraction of sulfur by polysulfide and iron complexes forming reactants [Y⁺] and electrochemical dissolution by polarization of the sulfide [high Fe³⁺ concentration]. All these mechanisms have been exploited by sulfide and iron-oxidizing bacteria. Basically, the bacterial action is a catalytic one during which [H⁺], [Fe³⁺] and [Y⁺] are breaking chemical bonds and are recycled by the bacterial metabolism. While the cyclic bacterial oxidative action via [H⁺] and [Fe³⁺] can be called indirect, bacteria had difficulties harvesting chemical energy from an abundant sulfide such as FeS₂, the electron exchange properties of which are governed by coordination chemical mechanisms (extraction of electrons does not lead to a disruption of chemical bonds but to an increase of the oxidation state of interfacial iron). Here, bacteria have evolved alternative strategies which require an extracellular polymeric layer for appropriately conditioned contact with the sulfide. Thiobacillus ferrooxidans cycles [Y⁺] across such a layer to disrupt FeS₂ and Leptospirillum ferrooxidans accumulates [Fe³⁺] in it to depolarize FeS₂ to a potential where electrochemical oxidation to sulfate occurs. Corrosion pits and high resolution electron microscopy leave no doubt that these mechanisms are strictly localized and depend on specific conditions which bacteria create. Nevertheless, they cannot be called ‘direct’ because the definition would require an enzymatic interaction between the bacterial membrane and the cell. Therefore, the term ‘contact’ leaching is proposed for this situation. In practice, multiple patterns of bacterial leaching coexist, including indirect leaching, contact leaching and a recently discovered cooperative (symbiotic) leaching where ‘contact’ leaching bacteria are feeding so wastefully that soluble and particulate sulfide species are supplied to bacteria in the surrounding electrolyte.