Kinetics of sulphuric acid leaching of a zinc silicate calcine

Recent developments of acid leaching and solvent extraction of zinc silicate ores have produced renewed commercial interest. However, the leaching kinetics of these concentrates has received little attention. This work, therefore, addresses the leaching of a zinc silicate concentrate in sulphuric acid. The effects of particle size (0.038–0.075mm), temperature (30–50°C) and initial acid concentration (0.2–1.0mol/L) were studied. The results show that decreasing the particle size while increasing the temperature and acid concentration increase the leaching rate. As leaching occurs, there is a progressive dissolution of willemite while the quartz and iron-containing phases remain inert. Among the kinetic models of the porous solids tested, the grain model with porous diffusion control successfully described the zinc leaching kinetics. The model enabled the determination of an activation energy of 51.9 ± 2.8kJ/mol and a reaction order of 0.64 ± 0.12 with respect to sulphuric acid, which are likely to be a consequence of the parallel nature of diffusion and chemical reaction in porous solids.     

Dissolution of MoS2 concentrate using NaClO from 283 to 373 K

This study analyses the leaching process of molybdenite (MoS₂) concentrate using sodium hypochlorite (NaClO) at temperatures ranging from 283 to 373?K. The following variables were studied: leaching time, stirring velocity, temperature, NaClO concentration, NaOH concentration, particle size and liquid:solid ratio. The optimum parameters for molybdenite dissolution were: Time?=?0.81?h, stirring speed?=?800?rev?min^??1, temperature?=?303?K, NaClO concentration?=?1.49M, NaOH concentration?=?8.01M, particle size?=?45?μm and liquid:solid ratio?=?300?:?1.?Under these conditions, molybdenite dissolution reached around 99%, while Cu and Fe recovery were insignificant. Analysis using XRD, QUESCAM and SEM showed release indices and subsequent reaction of the MoS₂ and the constituent parts of Cu and Fe.     

微波活化铜精矿加压浸出动力学

铜精矿经微波活化后,在初始硫酸浓度1.23 mol/L,液固比（mL/g）30/1,氧分压0.6 MPa,搅拌转速度500 r/min条件下,在408～453 K范围内研究了加压浸出动力学,并与微波活化前浸出动力学进行比较.结果表明,微波活化前后,铜精矿铜、锌浸出行为规律基本一致.在408～438 K范围内,铜精矿微波活化前后,铜浸出速率未见明显变化,而当温度升高至453 K后,铜浸出速率较微波活化前略有增大.当温度低于423 K时,锌浸出速率较微波活化前略有增大;当温度高于438 K时,锌浸出过程反而略有放缓.微波活化铜精矿铜、锌浸出反应的表观活化能分别为56.33、49.77 kJ/mol,铜、锌浸出过程均遵循界面化学反应控制的收缩核模型.与活化前相比,铜精矿经微波活化后铜、锌浸出过程得以促进.     

Reaction kinetics of the ferric chloride leaching of sphalerite—an experimental study

Chloride leaching processes have significant potential for treating complex sulfides. One advantage of chloride leaching is fast dissolution rates for most sulfide minerals. This experimental study is concerned with ferric chloride leaching of sphalerite, a common component of many complex concentrates. The effects of stirring, temperature, ferric ion concentration, and particle size have been examined. In addition, reaction residues at various levels of zinc extraction were examined by SEM, and the products of reaction were identified by energy dispersive X-ray analysis and X-ray diffraction. These observations indicated that the dissolution reaction is topochemical. Moreover, the leaching results fit a surface reaction control model. The activation energy was calculated to be 58.4 kJ/mole which is reasonable for a rate limiting surface reaction. The order of the reaction was 0.5 with respect to Fe³⁺ at low concentrations and zero at high concentrations. The change in reaction order occurred at similar Fe³⁺ concentrations for various particle sizes. This is believed to be indicative of an electrochemical reaction mechanism at low Fe³⁺ and an adsorption mechanism at higher Fe³⁺. A kinetic model for the ferric chloride leaching of sphalerite was also obtained for the lower Fe³⁺ concentrations and is given by: (ie5-01) This model is in excellent agreement with the experimental results for fractions of zinc extracted up to 0.95.     

The bioleaching of different sulfide concentrates using thermophilic bacteria

The bioleaching of different mineral sulfide concentrates with thermophilic bacteria (genusSulfolobus @#@) was studied. Since the use of this type of bacteria in leaching systems involves stirring and the control of temperature, the influence of the type of stirring and the pulp density on dissolution rates was studied in order to ascertain the optimum conditions for metal recovery. At low pulp densities, the dissolution kinetic was favored by pneumatic stirring, but for higher pulp densities, orbital stirring produced the best results. A comparative study of three differential concentrates, one mixed concentrate, and one global concentrate was made. Copper and iron extraction is directly influenced by bacterial activity, while zinc dissolution is basically due to an indirect mechanism that is activated in the presence of copper ions. Galvanic interactions between the different sulfides favors the selective bioleaching of some phases (sphalerite and chalcopyrite) and leads to high metal recovery rates. However, the formation of galvanic couples depends on the type of concentrate.     

Selective and Efficient Extraction of Zinc from Mixed Sulfide–oxide Zinc and Lead Ore

Mixed sulfide–oxide lead and zinc ores are generally composed of both sulfides and oxides. The dissolution of sulfides is more difficult than oxides thus the addition of oxidant is necessary. In this paper, oxidative leaching of mixed ore in NH₃-(NH₄)₂SO₄ solution using ammonium persulfate as oxidant under atmospheric pressure and relatively low temperature was investigated for the first time. The effects of factors on the leaching of pure ZnS were studied and the optimal conditions with zinc 98.7% were determined. Selective and efficient extractions of 93.9% and 94.9% zinc from zinc sulfide ore and mixed ore were also achieved, respectively.     

The ferric fluosilicate leaching of lead concentrates: Part I. Kinetic Studies

A new hydrometallurgical leaching process, which dissolves lead concentrates with acidified ferric fluosilicate solution, has been investigated for the selective extraction of lead and zinc from lead concentrates containing galena. The leaching of the Pine Point lead concentrate by ferric fluosilicate solutions was studied under various experimental conditions in the temperature range 20 °C to 95 °C. Temperature had a pronounced effect on the dissolution of the concentrates. The rates of lead leaching are very rapid over the temperature range 38 °C to 95 °C. The kinetics of zinc extraction are much lower than those of lead extraction. The reaction rates for the dissolution of galena were found to be controlled by surface chemical reaction. The apparent activation energy of the leaching reaction was calculated to be 62.1 kJ/mol. The initial concentrations of Pb²⁺, H⁺, and Fe³⁺ in the lixiviant do not have a significant effect on the rate or extent of lead extraction under the experimental conditions in this study.     

Pressure Leaching Kinetics of Multimetallic Sulphide Concentrates in H2SO4 + O2 Solution

Abstract

A multimetallic sulphide concentrate containing sphalerite, galena, chalcopyrite and silver in the matrix of pyrite was decomposed at elevated temperature and oxygen pressures in dilute sulphuric acid solutions for sufficient residence time to yield 95% of the zinc in the pregnant solution while most of the lead and silver remained in the residue together with most of the pyrite. The selective leaching process appeared to follow the diffusion controlled mechanism. The effects of concentration of the leachant, temperature and time of leaching, particle size, oxygen pressure and agitation on the leaching process were investigated. Results indicate the prospect of extracting not only all the metals but also appreciable amounts of elemental sulphur under optimized experimental conditions.     

Kinetics of Ammonia Leaching of Multimetal Sulphides

This paper briefly describes the studies carried out on oxidative ammonia leaching of Cu-Zn-Pb multimetal sulphides. Kinetics of zinc and copper dissolution were studied with ? 200 + 300 mesh BSS fraction and 1% solids in the slurry. It is observed that the dissolution of sphalerite proceeds by a phase boundary reaction model and that of copper via diffusion through product layer in the temperature range of 70-100°C. The rate equations for zinc and copper dissolution are given by:

1 ? (1 ? α)^1/3 = k _Zn[NH₃][pO₂]^1/2

1 ? 2/3α ? (1 2/3α )^2/3 = k^Cu[NH₃]²[pO₂]^1/2

where the symbols have the usual meanings.

Activation energies for zinc and copper dissolution reactions are estimated to be 66.5 and 55.4 kJ/mole, respectively. Activation energy values thus obtained are also comparable to those obtained using a differential approach.

The leaching results obtained with 10% solids using a wide range of particle size (? 140 + 500 mesh) indicate that copper dissolution is chemically controlled in ammonia as well as ammonia-ammonium sulphate medium in the temperature range of 115-135°C. However, at lower temperature (?55°C). the leaching reaction follows a diffusion model. Zinc dissolution data show deviations from the shrinking core model due to high extractions in the initial stages.     

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.     

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.     

Mechanochemical leaching of refractory zinc silicate (hemimorphite) in alkaline solution

A process was developed for extracting zinc from refractory zinc silicate (hemimorphite) by mechano-chemical leaching in sodium hydroxide solution with calcium oxide added to inhibit the leaching of silica. The effects of mechanical activation, calcium oxide dosage, temperature, leaching time, liquid–solid ratio and sodium hydroxide concentration were investigated. The dissolution of silica was considerably inhibited by adding CaO, but zinc extraction was almost unchanged. Under optimum conditions in 4.5 mol/L NaOH and twice the stoichiometric CaO at 383 K, the dissolution of Zn and SiO₂ were 93.6% and 7.3%, respectively, after 1 h using a liquid–solid ratio (L/S) of 10:1.     

Oxidative Pressure Leaching of Silver from Flotation Concentrates with Ammonium Thiocyanate Solution

The thermodynamics and technologies of the selective pressure leaching of silver from flotation concentrates were investigated in an ammonium thiocyanate medium. Thermodynamic analyses, which include silver solubility in NH₄SCN solution and Eh-pH diagrams of the Me-MeS-NH₄SCN-H₂O system at 25 °C, were discussed. The effects of several factors, such as temperature, leaching time, oxidant, pH value, flotation concentrates concentration, surfactant concentration, and so on, on the extraction percentages of silver and zinc were investigated. The following optimal leaching conditions were obtained: NH₄SCN concentration 1.5 M, lignin concentration 0.5 g/L, Fe³⁺ concentration 2 g/L, flotation concentrates addition 200 g/L, and oxygen pressure 1.2 MPa at 130 °C for 3 hours. Under these optimum conditions, the average extraction percentage of silver exceeded 94 pct, whereas the average extraction percentage of zinc was less than 3 pct. Only 7 pct of ammonium thiocyanate was consumed after 4 cycles, which indicated that ammonium thiocyanate hardly was oxidized under these oxidative pressure leaching conditions.     

Oxidative Leaching of Chalcopyrite by Cupric Ion in Chloride Media

In this study, chalcopyrite concentrate produced in Sarcheshmeh copper plant was subjected to oxidative leaching investigation by cupric ion to determine the effect of several parameters on the copper and iron dissolution, including temperature of leaching, time of holding and cupric concentration as oxidant agent in a range of 38–97 °C, 1.5–8.5 h and 0.2–0.8 M, respectively. The leaching media was chloride providing with 3 M HCl 37% and CuCl₂. The experiments were designed by central composite design method. The dissolution of copper and iron was examined. The maximum dissolution of copper 62.64%, was obtained at 85 °C, 7 h and oxidant concentration of 0.7 M. The kinetics model of chalcopyrite leaching and an optimized condition with maximum dissolution of copper and minimum dissolution of iron was obtained by Minitab^®18 software. Additionally, the thermodynamics of leaching was investigated by Pourbaix diagrams of copper and iron composition, using HSC Chemistry6 software. It was found that the oxidative leaching process is controlled by diffusion through passivation layer with an activation energy of 19.57 kJ/mol.

     

Determination of the Optimum Conditions for Leaching of Zinc Cathode Melting Furnace Slag in Ammonium Chloride Media

This research is part of a continuing effort to leach zinc from zinc cathode melting furnace slags (ZCMFSs) to produce zinc oxide. The slag with an assay of 68.05 pct Zn was used in ammonium chloride leaching for zinc extraction. In this paper, the effects of influential factors on extraction efficiency of Zn from a ZCMFS were investigated. The Taguchi’s method based on orthogonal array (OA) design has been used to arrange the experimental runs in order to maximize zinc extraction from a slag. The softwares named Excel and Design-Expert 7 have been used to design experiments and subsequent analysis. OA L ₂₅ (5⁵) consisting of five parameters, each with five levels, was employed to evaluate the effects of reaction time (t = 10, 30, 50, 70, 90 minutes), reaction temperature [T = 313, 323, 333, 343, 353 (40, 50, 60, 70, 80) K (°C)], pulp density (S/L = 20, 40, 60, 80, 100 g/L), stirring speed (R = 300, 400, 500, 600, 700 rpm), and ammonium chloride concentration (C = 10, 15, 20, 25, 30 pctwt), on zinc extraction percent. Statistical analysis, ANOVA, was also employed to determine the relationship between experimental conditions and yield levels. The results showed that the significant parameters affecting leaching of slag were ammonium chloride concentration and pulp density, and increasing pulp density reduced leaching efficiency of zinc. However, increasing ammonium chloride concentration promoted the extraction of zinc. The optimum conditions for this study were found to be t ₄: 70 minutes, T ₅: 353 K (80 °C), (S/L)₂: 40 g/L, R ₃: 500 rpm, and C ₄: 25 pctwt. Under these conditions, the dissolution percentage of Zn in ammonium chloride media was 94.61 pct.     

Decomposition of mixed rare earth concentrate by NaOH roasting and kinetics of hydrochloric acid leaching process

A new clean extraction technology for the decomposition of Bayan Obo mixed rare earth concentrate by NaOH roasting is proposed.The process mainly includes NaOH roasting to decompose rare earth concentrate and HCl leaching roasted ore.The effects of roasting temperature,roasting time,NaOH addition amount on the extraction of rare earth and factors such as HCl concentration,liquid-solid ratio,leaching temperature and leaching time on the dissolution kinetics of roasted ore were studied.The experimental results show that when the roasting temperature is 550℃ and the roasting time is 60 min,the mass ratio of NaOH:rare earth concentrate is 0.60:1,the concentration of HCl is 6.0 mol/L,the ratio of liquid to solid(L/S) 6.0:1.0,and the leaching temperature 90℃,leaching time 45 min,stirring speed 200 r/min,and the extraction of rare earth can reach 92.5%.The relevant experimental data show that the process of HCl leaching roasted ore conforms to the shrinking core model,but the control mechanism of the che mical reaction process is different when the leaching temperature is different.When the leaching temperature is between 40 and 70℃,the chemical reaction process is controlled by the diffusion of the product through the residual layer of the inert material.The average surface activation energy of the rare earth element is E_a=9.96 kJ/mol.When the leaching temperature is 75-90℃,the chemical reaction process is controlled by the interface transfer across the product layer(product layer interface mass transfer) and diffusion.The average surface activation energy of rare earth elements is E_a=41.65 kJ/mol.The results of this study have certain significance for the green extraction of mixed rare earth ore.     

Dissolution behaviour of a beryl ore for optimal industrial beryllium compound production

This study involves the leaching of the beryl ore with sulphuric acid (H₂SO₄) solution for predicting optimal beryllium extraction conditions with the aim of assessing the importance of leachant concentration, reaction temperature and particle size on the extent of dissolution. A kinetic model to represent the effects of these variables on the leaching rate was developed. It was observed that the dissolution of beryl ore increases with increasing H₂SO₄ concentration, temperature, decreasing particle size and solid to liquid ratio. At optimal leaching conditions, 89.3% of the ore was reacted by 1.25?mol/L at 75°C temperature and 120 minutes with moderate stirring, where 1612.0?mg/L Be²⁺, 786.7?mg/L Al³⁺, 98.1?mg/L Fe³⁺ and 63.4?mg/L Ag⁺ were found as major species in the leach liquor. The unleached products constituting about 10.7% were examined by X-ray diffraction (XRD) and found to contain primarily, siliceous compounds such as Xonotlite, Antigorite, Chrysolite and Kaolinite.     

Kinetics of copper dissolution during pressure oxidative leaching of lead-containing copper matte

The kinetics of pressure oxidative leaching of lead-containing copper matte with sulfuric acid was investigated. The effects of particle size, leaching temperature, oxygen partial pressure and sulfuric acid concentration on the kinetics and mechanism of copper extraction were studied. It was found that the reaction kinetic model follows the shrinking core model of chemical reaction control and the apparent activation energy was determined as 39.1 kJ/mol. The order of the reaction with respect to total pressure was found to be 0.64. The kinetic equations for the effect of particle size, leaching temperature, total pressure and sulfuric acid concentration were obtained and a mathematical model of copper extraction from lead-containing copper matte was developed as:

This equation estimates the extraction of copper with very good agreement (r = 0.99) between the experimental and calculated values.     

硫氧混合铅锌矿中锌的氧化氨浸动力学

常压低温条件下在NH₃-（NH₄）₂SO₄体系中使用过硫酸铵作为氧化剂对硫氧混合铅锌矿中的锌进行浸出实验,系统研究了搅拌速度、浸出剂浓度、氧化剂浓度与温度对于锌浸出率的影响.结果表明,在最优条件下锌的浸出率可达93.2%,且浸出过程中几乎没有其他离子进入溶液,实现了锌的选择性高效浸出,从而简化了后续的浸出液净化与材料制备过程.动力学研究表明,硫氧混合铅锌矿中锌的氧化氨浸过程遵循固体产物层扩散控制的未反应核收缩模型,浸出反应的表观活化能为17.89 kJ·mol-1.      

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.