A partial equilibrium model to characterize the precipitation of ferric ion during the leaching of chalcopyrite with ferric sulfate

A partial equilibrium model has been developed and used to characterize the conditions under which precipitation of ferric ion occurs during the dump leaching of chalcopyrite ores. The precipitates which have been considered include amorphous Fe(OH)₃, α-FeOOH (goethite), and Na⁺, K⁺, Ag⁺, Pb²⁺, and H₃O⁺ jarosites. Solution of the model equations makes possible the determination of the concentrations of the solution species during leaching of the mineral. The concentration product for Fe(OH)₃ (am) and α-FeOOH was calculated for changing solution concentrations and compared with the solubility product constants to determine when precipitation would be expected thermodynamically. The K⁺, Na⁺, Ag⁺, and Pb²⁺ concentrations that would be necessary to satisfy the solubility product constants for the corresponding jarosites were calculated for various initial concentrations and varying amounts of O₂ consumption. Formerly Graduate Assistant, Ames Laboratory USDOE and Department of Chemical Engineering, Iowa State University, Ames, IA 50011     

Chloride leaching of chalcopyrite

Two new process flowsheets have been developed which combine chloride leaching of copper from chalcopyrite with solvent extraction, to selectively transfer copper to a conventional sulfate electrowinning circuit. Chloride leaching with copper(II) as oxidant offers significant advantages for copper including increased solubility and increased rates of leaching. Both process flowsheets were similarly designed with a two stage counter-current leach but differ with respect to iron deportment. The goethite model flowsheet includes sparging of air or oxygen to the second leach stage to aid precipitation of iron as goethite (FeOOH). The hematite model flowsheet precipitates iron as hematite (Fe₂O₃) downstream from the leach in a dedicated autoclave. A mass balance has been completed for both process flowsheets and this determined the concentrations of copper and iron species in feed liquor returning to the leach following copper solvent extraction.The optimum leach extraction conditions were determined by varying grind size, temperature and residence time for both leach model scenarios. Leach tests were conducted using a chalcopyrite concentrate from Antamina in northern Peru, which contains a low to moderate amount of gangue material. The hematite model was also examined using a Rosario concentrate from Chile which contained chalcocite in addition to chalcopyrite and significant pyrite. Leach tests based on the hematite model were successful in achieving copper extractions > 95% in 4–6 h at 95 °C after fine grinding the concentrate (P₉₀ = 41 μm). However, copper extraction exceeded 99% from the finely ground Rosario concentrate (P₉₀ = 37 μm). In the goethite model leach tests, 89% copper extraction was achieved under optimum conditions in the atmospheric conditions tested.     

A general model for the reaction of several minerals and several reagents in heap and dump leaching

The rate of recovery of metal values from heap and dump leach operations has been stimulated by mathematically modelling reactions among several minerals and reagents. The principal steps are: (i) representation of all possible reactions in a matrix, (ii) division of the system into coherent groups of reactions that proceed at the same rate, (iii) calculation of the rate of reaction of the coherent groups.The procedure models sequential and complementary reactions, including those which involve substances generated within the heap. (It is assumed that the rates of all reactions are controlled by molecular diffusion within the particles.) A general equation for multi-mineral, multi-reagent systems is derived and used in the algorithm.The procedure has been used with a number of systems of different complexity. It has shown that not all of the possible reactions occur under a particular set of conditions, times and location within the heap. The algorithm has been incorporated into a generalised, comprehensive simulation of heap and dump leaching processes, which is independent of any process testwork data. Comparison of predicted results and test results reported by others is included.     

Activated sulfating roasting of the chalcopyrite concentrate for sulfuric-acid leaching

The sulfating roasting of the chalcopyrite concentrate with various pyrite contents is investigated in activated conditions in order to obtain the preferentially sulfate cinder most suitable for leaching and for the effective recovery of copper and accompanying iron. The activating effect that the trivalent iron sulfate which forms as a result of the low-temperature oxidation of pyrite, as well as the steam-air medium, has on the kinetics and completeness of the sulfatization of the concentrate are shown.     

Ferric chloride leaching of sulphidized chalcopyrite

Reactions between chalcopyrite and elemental sulphur at between 320°C and 400°C produce various combinations of covellite, nukundamite (Cu_5,5 FeS_6,5, and pyrite depending on the sulphur content of the products. If all the contained copper is transferred to covellite, and iron to pyrite, at least 95% of the copper can be leached by ferric chloride with negligible iron dissolution at up to 70°C. At above 70°C, iron is extracted from pyrite. When nukundamite is present, the iron and copper contents of this compound are leached simultaneously, and the consequent loss of selectivity is further enhanced at above 70°C by attack of pyrite. The increased rates of copper extraction obtained following sulphidation result from both the breakdown of chalcopyrite and a greater porosity.     

A kinetic study of the leaching of chalcopyrite at elevated temperatures

A study of the rate of dissolution of chalcopyrite (CuFeS₂) in acidic solutions under oxygen overpressures was carried out by measuring the rate of formation of cupric ions in solution. Effects of temperature, oxygen partial pressure, surface area, and concentration of sulfuric acid were evaluated. A sized batch of chalcopyrite was leached in the temperature range of 125 to 175° and in the pressure range of 75 to 400 psi of oxygen. In 0.5N H₂SO₄ all products of reaction went into solution except for trace amounts of elemental sulfur. The dissolution of chalcopyrite followed linear kinetics and was essentially independent of hydrogen ion concentration for H₂SO₄ concentrations between 0.2 and 0.5 JV. The oxygen dependence indicated adsorption approaching limiting values with increasing oxygen pressure. The linear mechanism was explained in terms of steady-state adsorption of oxygen at the chalcopyrite surface followed by a surface reaction. The enthalpy of activation for adsorption of oxygen was found to be approximately 33 kcal per mole. An activation enthalpy of approximately 9 kcal per mole was observed for the surface reaction. Charge transfer reaction are not rate controlling in the process.     

Reaction mechanism for the acid ferric sulfate leaching of chalcopyrite

The acid ferric sulfate leaching of chalcopyrite, CuFeS₂ + 4Fe⁺³ = Cu⁺² + 5Fe⁺² + 2S⁰ was studied using monosize particles in a well stirred reactor at ambient pressure and dilute solid phase concentration in order to obtain fundamental details of the reaction kinetics. The principal rate limiting step for this electrochemical reaction appears to be a transport process through the elemental sulfur reaction product. This conclusion has been reached in other investigations and is supported by data from this investigation in which the reaction rate was found to have an inverse second order dependence on the initial particle diameter. Furthermore, the reaction kinetics were found to be independent of Fe⁺³, Fe⁺², Cu⁺² and H₂SO₄ in the range of additions studied. The unique aspect of this particular research effort is that data analysis, using the Wagner theory of oxidation, suggests that the rate limiting process may be the transport of electrons through the elemental sulfur layer. Predicted reaction rates calculated from first principles using the physicochemical properties of the system (conductivity of elemental sulfur and the free energy change for the reaction) agree satisfactorily with experimentally determined rates. Further evidence which supports this analysis includes an experimental activation energy of 20 kcal/mol (83.7 kJ/mol) which is approximately the same as the apparent activation energy for the transfer of electrons through elemental sulfur, 23 kcal/ mol (96.3 kJ/mol) calculated from both conductivity and electron mobility measurements reported in the literature. formerly Metallurgy Graduate Student, University of Utah.     

Direct selective leaching of chalcopyrite concentrate

The ammonium persulphate (APS) leaching of chalcopyrite concentrate in the presence of ammonium carbonate was studied. The effects of ammonium carbonate concentration, APS concentration, leaching time, leaching temperature, solid/liquid ratio and stirring speed were investigated. Optimum leaching conditions were found as follows: APS concentration is 200 g L^?1; ammonium carbonate concentration is 200 g L^?1; leaching time is 180 min; leaching temperature is 60°C; solid/liquid ratio is 0·04 g mL^?1; and stirring speed is 400 rev min^?1. Under these conditions, copper extraction yield was obtained at about 72%. Furthermore, iron extraction yield decreased with increasing ammonium carbonate concentration and iron did not pass into solution under this condition. X-ray and SEM analysis also supported these results. It was determined that the copper extraction results were satisfactory by way of all experiments were performed under atmospheric conditions (i.e. low temperature and atmospheric pressure) and achieved selective copper leaching from chalcopyrite concentrate.

On a étudié la lixiviation au persulfate d’ammonium (APS) du concentré de chalcopyrite en présence de carbonate d’ammonium. On a examiné l’effet de la concentration du carbonate d’ammonium, de la concentration d’APS, de la durée et de la température de lixiviation, du rapport solide-liquide et de la vitesse d’agitation. Les conditions optimales de lixiviation suivent: concentration d’APS?=?200 g L^?1; concentration de carbonate d’ammonium?=?200 g L^?1; durée de lixiviation?=?180 min; température de lixiviation?=?60°C; rapport solide-liquide?=?0·04 g mL^?1; et vitesse d’agitation?=?400 rpm. Avec ces conditions, le rendement d’extraction du cuivre était d’environ 72%. De plus, le rendement d’extraction du fer diminuait avec l’augmentation de la concentration de carbonate d’ammonium, qui empêchait également le fer de passer en solution. L’analyse aux rayons x et au SEM supportait également ces résultats. Grâce à toutes les expériences effectuées en conditions atmosphériques (c’est-à-dire à basse température et à pression atmosphérique), on a déterminé que les résultats d’extraction du cuivre étaient satisfaisants et qu’on obtenait une lixiviation sélective du cuivre à partir du concentré de chalcopyrite.     

A chemical equilibrium model of the imperial smelting blast furnace

Abstract

A mathematical model of the Imperial Smelting lead-zinc blast furnace was developed to determine the critical reactions and to evaluate certain parameters of its operation. The model was based on the assumption that equilibrium existed throughout, except that carbon was assumed not to react until it reached the tuyeres. The two major reactions occurring in the lower parts of the furnace are found to be:

Zn(g) +CO₂ = ZnO +CO (1)

ZnS + Pb = Zn(g) + PbS(g) (2)

The second one has heretofore been neglected, but it is critical in predicting the furnace behaviour. Large quantities of lead, zinc, and sulphur circulate inside the furnace in the tuyere-equilibrium zone. Results calculated by the model show that sinter sulphur content and tap temperature have a major effect on this circulation and on zinc recovery and slag composition. The predicted results and published furnace data are in good agreement, indicating that a calculation that assumes equilibrium is a useful approximation of a real furnace.

Résumé

Un modèle mathématique du haut--fourneau «Imperial Smelting» a zinc et plomb est proposé pour etablir les réactions critiques et pour évaluer certains paramètres de fonctionnement. Le modéle est basé sur la supposition que le système est entièrement en équilibre, sauf le carbone qui est supposé ne pas réagir avant d'atteindre la région des tuyères. On trouve que les deux réactions principales qui s'effectuent au fond du four sont:

Zn(g) +CO₂ = ZnO +CO (1)

ZnS + Pb = Zn(g) + PbS(g) (2)

Bien que la deuxième réaction ait été ignorée jusqu'à présent, la prise en considération de cette réaction est essentielle si on veut prévoir Ie comportement du four. De grandes quantités de plomb, zinc et soufre circulent dans le four dans la région d'équilibre des tuyères. Des calculs a partir du modèle indiquent que la teneur en soufre des boulettes ainsi que la température de la coulée ont une influence importante sur cette circulation ainsi que sur la récupération du zinc et sur la composition des scories. Le fait que les résultats prévus et les données publiees pour ce haut-fourneau soient en bon accord indique qu'un modèle base sur la supposition de l'équilibre est une approximation valable pour un vrai four.     

Copper leaching from nanoparticles of chalcopyrite concentrate

In this research, the effect of mechanical activation, via ball milling, of copper sulfide concentrate on the efficiency and mechanism of copper leaching in an acidic ferric chloride solution was investigated. Copper concentrate, containing chalcopyrite as the main constituent, was supplied from Sarcheshme mine, located in the central part of Iran. By 24 hours of ball milling, the size of the concentrate particles was decreased to about 50 nm. The leaching of ball-milled samples was completed in a relatively more dilute solution and in shorter leaching times than that of unmilled samples. Ball milling of the concentrate from 0 to 180 minutes led to an increase in the efficiency of copper leaching in a acidic ferric chloride solution from 43 to 86%. Copper recovery was completed after 360, 540, 960, and 1440 minutes of leaching for the samples subjected to ball milling for 300, 240, 120, and 60 minutes, respectively. The activation energy of leaching process in the unmilled sample was 60.23 kJ/mol, while the activation energy of leaching for 24 hours of the ball-milled concentrate was 5.56 kJ/mol. The intensive decrease of activation energy after 24 hours ball milling was thought to be the result of the acceleration of the chemical reaction in very fine particles of the nanometrical scale. The low amount of activation energy suggested that the leaching rate controlling step was mass transfer in the concentrate milled for 24 hours. This article was submitted by the authors in English.     

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.     

The leaching of chalcopyrite with ferric chloride

A comparative study of electrochemical leaching and chemical leaching of chalcopyrite was done to elucidate the leaching mechanism of chalcopyrite with FeCl₃. The leaching rate of chalcopyrite exhibits a half order dependency on the FeCl₃ concentration, whereas it is independent of the FeCl₂ concentration. The mixed potential of chalcopyrite exhibits a 72 mV · decade⁻¹ dependency upon FeCl₃ concentration; no influence on the mixed potential was observed by the addition of FeCl₂. In FeCl₃ solutions acidified with HC1, the predominant chemical species of Fe(III) was found to be FeCl ₂ ^u+ from equilibrium calculations. The concentration of this species is approximately proportional to the amount of FeCl₃ added to the solutions. Based on these observations, an electrochemical mechanism is proposed which involves the oxidation of chalcopyrite and the reduction of FeCl ₂ ⁺ , the predominant species of Fe(III). By converting the leaching rate to electric current density,i, 140 mV · decade⁻¹ dependency of mixed potential,E, against logi is obtained. This dependency of the chemical leaching of chalcopyrite with FeCl₃ as well as its activation energy agree with those for electrochemical leaching. These findings strongly support the electrochemical mechanism of FeCl₃ leaching of chalcopyrite. Formerly Graduate Student, Kyoto University     

氯化物浸出黄铜矿的试验研究

研究了硝酸催化氧化、氯化物浸出黄铜矿的机理,考察了氧化和浸出反应的影响因素。研究结果表明:硝酸催化氧化的优化条件是酸度0.50mol/L、温度95℃、硝酸量0.5%～1.0%。采用浸出、氧化分离式新工艺,4.5h铜的浸出率达97%,氧化除铁率97%,可实现黄铜矿浸出的综合优化。     

氯化物浸出黄铜矿的试验研究

研究了硝酸催化氧化、氯化物浸出黄铜矿的机理，考察了氧化和浸出反应的影响因素。研究结果表明：硝酸催化氧化的优化条件是酸度。0．50mol／L、温度95℃、硝酸量0．5％～1．0％。采用浸出、氧化分离式新工艺，4．5h铜的浸出率达97％，氧化除铁率97％，可实现黄铜矿浸出的综合优化。     

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.     

A model of the dump leaching process that incorporates oxygen balance,heat balance,and air convection

A one dimensional, nonsteady-state model of the copper waste dump leaching process has been developed which incorporates both chemistry and physics. The model is based upon three equations relating oxygen balance, heat balance, and air convection. It assumes that the dump is composed of an aggregate of rock particles containing nonsulfide copper minerals and the sulfides, chalcopyrite and pyrite. Leaching occurs through chemical and diffusion controlled processes in which pyrite and chalcopyrite are oxidized by ferric ions in the lixiviant. Oxygen, the primary oxidant, is transported into the dump by means of air convection and oxidizes ferrous ion through bacterial catalysis. The heat generated by the oxidation of the sulfides promotes air convection. The model was used to simulate the leaching of copper from a small test dump, and excellent agreement with field measurements was obtained. The model predicts that the most important variables affecting copper recovery from the test dump are dump height, pyrite concentration, copper grade, and lixiviant application rate.     

矿堆非饱和渗流中的界面作用

为探明矿堆非饱和浸出渗流规律,以界面作用为切入点,分析了浸出液的运动状态.矿堆中吸力由界面作用产生的基质吸力和吸收扩散产生的渗透吸力组成.孔隙中介质分布的不均匀性和矿石形状的随机性是导致界面作用多样性以及浸出液运动状态复杂的原因.采用毛细上升实验很好地证明了矿堆中吸力的存在,在驱动力的作用下,浸润初期的液面上升速度较快,浸润后期液面上升相对平缓.通过拟合得知液面毛细上升高度与浸润时间符合幂函数关系.理论研究表明可以通过改变固相、液相和气相的物理性质来实现对矿堆非饱和渗流中界面作用的调控.

     

Equilibria associated with cupric chloride leaching of chalcopyrite concentrate

The leaching of chalcopyrite in aqueous cupric chloride solutions is treated with a thermodynamic model which incorporates the equilibria associated with the formation of various cupric, cuprous and ferrous chloride complexes. The precipitate from a refluxing slurry of elemental sulfur and aqueous cuprous and ferrous chlorides contained predominantly covellite (CuS) and a small amount of pyrite (FeS₂). Mössbauer spectra indicate that there was no chalcopyrite formation. Cupric chloride leaching experiments on both chalcopyrite concentrate and cupric sulfide indicated that the extent of cuprous ion formation in these aqueous cupric chloride solutions is controlled by the thermodynamics of the equilibrium:

$\text{CuS+Cu}{}^{\mathrm{2+}}\text{? 2 Cu}{}^{\mathrm{+}}\text{+S}{}^0$

Conditions which favor the desired complete reduction of cupric ion are discussed.     

Microbiological leaching of a chalcopyrite concentrate by Thiobacillus ferrooxidans

The microbiological leaching of a chalcopyrite concentrate has been investigated using a pure strain of Thiobacillus ferrooxidans. The optimum leaching conditions regarding pH, temperature, and pulp density were found to be 2.3, 35 degrees C, and 22% respectively. The energy of activation was calculated to be 16.7 kcal/mol. During these experiments the maximum rate of copper dissolution was about 215 mg/liters/hr and the final copper concentration was as high as 55 g/liter. This latter value is in the range of copper concentrations which may be used for direct electrorecovery of copper. Jarosite formation was observed during the leaching of the chalcopyrite concentrate. When the leach residue was reground to expose new substrate surface, subsequent leaching resulted in copper extractions up to about 80%. On the basis of this experimental work, a flow sheet has been proposed for commercial scale biohydrometallurgical treatment of high-grade chalcopyrite materials.