Model for ferric sulfate leaching of copper ores containing a variety of sulfide minerals: Part II. Process modeling of in situ operations

The Lin-Sohn Leach simulation (DLS) computer model was used to simulate copper extraction from run-of-mine fragment size distributions expected for underground modifiedin situ solution mining. Ventilation was found to be a key variable in obtaining good copper recoveries, and the rate of ventilation was found to vary with ore type and through time. Improving recoveries and recovery rates with improved fragmentation was found to be more difficult than expected. Limitations on surface kinetics and bacterial ferric ion generation often erode benefits expected from reduction in diffusion limitation. Control of fines formation is difficult with the current state of the art in blasting. Fines may also improve the initial leaching rate at the expense of increased ventilation cost and more limited ultimate recovery. Ore grade was found to be less important to solution mining than to conventional copper recovery. Mineralogy was found to be crucial to solution mining economics. Mineralogy affects the grade of the ore and the importance of that grade. Pyrite was found to be helpful in increasing the ultimate copper recovery from ore but, at the same time, was found to reduce the rate of that recovery. Recovery rates were found to be very sensitive to mineral grain size, and calibration of DLS to a given ore may be especially important in this variable. Tortuosity and porosity of ore fragments appear relatively unimportant over the ranges typical of sulfide copper porphyry ore. Modifiedin situ solution mining appears to be a competitive option when a favorable ore type is available. The DLS computer code may be a valuable aid in screening options and selecting the most favorable orebodies.     

Mixed-control kinetics of oxygen leaching of chalcopyrite and pyrite from porous primary ore fragments

The leaching of individual copper ore fragments was investigated under oxygen pressure of 690 to 3100 kPa. After 500 hours of leaching at 90°C under 3100 kPa oxygen pressure, 50 pct of copper was extracted from an ore particle of 0.57 cm size, while 1300 hours were needed to extract 90 pct of copper under the same experimental conditions. A mathematical model incorporating the intrinsic leaching kinetics of chalcopyrite and pyrite grains and the pore diffusion of dissolved oxygen within the ore fragment was formulated to interpret the experimental results. Both chemical reaction and diffusion were found to be important to the overall leaching rate in the particle size range tested. The model can also be used to predict the leaching behavior of particles of different sizes at different temperatures and oxygen partial pressures. H. K. LIN, formerly Graduate Student in the Department of Metallurgy and Metallurgical Engineering, University of Utah     

In-place leaching of primary sulfide ores: Laboratory leaching data and kinetics model

Experimental results obtained in laboratory leaching studies of primary copper sulfide ore in sulfuric acid systems pressurized with oxygen are interpreted by a computerized geometric model involving the movement of a reaction zone through the ore fragments. Physical properties of the ore, including size, shape, and mineral content, are considered. The leaching mechanism involves mixed kinetics and includes a surface reaction within a moving reaction zone plus pore diffusion of dissolved oxygen through the reacted portion of the ore fragment to the reaction zone. The results are applicable to conditions that would exist innuclear solution mining or similar processes in which the ore is converted into rubble and then inundated by a leach solution adequately supplied with oxidants. Experimental results at 90°C are correlated with the model, which includes temperature-dependent parameters.     

Evaluation of velocity-dependent in situ leaching processes: Single-porosity model

A methodology to characterize processes of in situ leaching is developed to study various parameters that may affect the recovery of a valuable mineral and to assure a successful application of the in situ leaching technology. A leaching kinetics model is first derived based on the concept of representative elementary volume (REV) in porous ore deposits. Every parameter in the model is clearly defined and may be easy to obtain in practice. Then the governing equations are obtained for transport of both the reagent and the dissolved mineral in porous ore deposits. These equations are solved by a fully three-dimensional computer simulator. Effects on the mineral recovery of parameters such as lumped rate constant, ore porosity, injection flux, and macrodispersivities are investigated through the three-dimensional case with one injection well. It is found that all of these four parameters, interconnected through the ore porosity, significantly affect the effectiveness of the in situ leaching. Furthermore, the flow velocity of leach solution within an ore deposit may be the key parameter to the design of a real leaching mine because all of the other three parameters are site specific.     

Leaching kinetics of copper from natural chalcocite in alkaline Na4EDTA solutions

The leaching behavior of copper from natural chalcocite (Cu₂S) particles in alkaline Na₄EDTA solutions containing oxygen was examined at atmospheric pressure. The EDTA leaching process took place with consecutive reactions, where the solid product of the first reaction, covellite (CuS), became the reactant for the second. The copper leached into the alkaline solutions was immediately consumed by the chelation of copper (II) with EDTA, and the mineral sulfur was completely oxidized to sulfate ion. The experimental data for the leaching rate of copper were analyzed with a familiar shrinking-particle model for reaction control. The conversion rate of chalcocite to covellite was found to be about 10 times as high as the dissolution rate of covellite. The time required for complete dissolution of covellite was directly proportional to the initial particle size and was inversely proportional to the square root of the product of the hydroxide ion concentration and the oxygen partial pressure, but it was independent of the Na4EDTA concentration in the presence of excess Na₄EDTA. The observed effects of the relevant operating variables on the dissolution rate were consistent with a kinetic model for electrochemical reaction control. The kinetic model was developed by applying the Butler-Volmer equation to the electrochemical process, in which the anodic reaction involves the oxidation of covellite to copper (II) ion and sulfate ion and the cathodic reaction involves the reduction of oxygen in alkaline solution. The rate equation allowed us to predict the time required for the complete leaching of copper from chalcocite in the alkaline Na4EDTA solutions.     

LEACHING OF MASSIVE RICH COPPER ORE WITH ACIDIFIED FERRIC CHLORIDE

Massive rich copper ore, taken from the Küre region of Turkey and containing 7.95% Cu, 2.13% Pb, 0.341% Zn, 0.087% Co, 0.039% Ni, 0.9 (g/t) Au, and 17 (g/t) Ag, was subjected to the experimental study. Effects of leaching time, ferric ion concentration, solid/liquid ratio, acid concentration and temperature on the metal dissolution efficiencies were investigated in order to determine optimum leaching conditions and related leaching kinetics. As a result of these experiments carried out under the optimum leaching conditions, 76% of Cu, 55% of Co, 96% of Ni, 100% of Pb, and 91% of Zn were extracted. The value of apparent activation energy for copper dissolution was found as 43.8?kJ/mole, which is in agreement with the published data. The rate-controlling step was found as the diffusion of ferric ions into the sulfur layer formed on the surface of partially leached ore particles.     

Catalytic decomposition of hydrogen peroxide by ferric ion in dilute sulfuric acid solutions

Hydrogen peroxide decomposition in acidic solutions is catalyzed by the free ferric ion, Fe³⁺. The following rate law for this reaction is determined by the initial rate method in solutions similar to those used for acidicin situ uranium leaching: wherek = 4.3 × 10⁻³ s^°1 at 25 °C. From 25° to 50 °C, the activation energy is 85.6 kJ/mol. The decomposition of hydrogen peroxide proceeds by a particular redox reaction sequence that depends on the ratio of the concentrations of hydrogen peroxide to free ferric ion. The rate law determined here is consistent with the form derived from the redox sequence for the case where the ratio of hydrogen peroxide to free ferric ion concentration is greater than 1.0. The magnitude of the rate constant indicates that the decomposition of hydrogen peroxide may cause rapid loss of this oxidant in leaching solutions containing ferric ion. Formerly a Graduate Student with the Department of Geochemistry and Mineralogy, Pennsylvania State University,     

含碳酸盐脉石氧化铜矿的酸浸动力学

针对碳酸盐脉石对氧化铜矿酸浸动力学的影响进行探讨,研究了温度、酸度、矿石粒径、液固质量比、振荡速度等因素对含碳酸盐脉石氧化铜矿浸出的影响.结果表明,高温、高酸度、高液固质量比、小粒径和高振荡速度利于矿石的浸出,但碳酸盐脉石使得酸耗增加.考虑浸出成本确定合理的浸出条件为温度303 K、酸度35 g·L-1、矿石粒径0.074~0.125 mm、液固质量比3∶1以及振荡速度180 r·min-1,浸出180 min后铜浸出率达53.6%.对浸出前后矿石表面形貌进行分析.结果显示碳酸盐脉石与酸反应后在矿石表面形成CaSO₄·2H₂O沉淀,覆盖在颗粒表面,限制了矿石颗粒孔裂隙的发育.基于收缩未反应核模型对浸出动力学进行分析,发现碳酸盐脉石反应生成的沉淀阻碍了浸出反应,固体产物层扩散为浸出反应的控制步骤,反应的表观活化能为8.65 kJ·mol-1.      

Acidic rate- and flow-controlled dissolution of uraninite ores

A mathematical model is developed to calculate rates of uranium leaching as a function of flow rate and oxidant concentration for acidic solutions. The model is based on a simple plug-flow formulation for the conservation of mass and on experimental rate expressions for the aqueous oxidation of uraninite and pyrite by either ferric ion or hydrogen peroxide. Model calculations of oxidant and dissolved uranium concentrations are generally in good agreement with those measured in column experiments in which synthetic ores consisting of mixtures of uraninite, pyrite, and quartz sand were leached by dilute sulfuric acid solutions (pH = 1.7) that contained ferric ion or hydrogen peroxide as the oxidant. The experimental and modeling results show that uraninite is preferentially leached from pyritic ores by ferric ion solutions, but is less selectively leached by hydrogen peroxide. The calculated and measured rates of uranium leaching for pyritic ores indicate that it is most efficient to use high oxidant concentrations and high flow rates to take advantage of the more rapid reaction rates of uraninite with oxidant compared to reaction rates of pyrite with oxidant. At high flow rates, uranium extraction is maximized throughout a proportionately greater amount of ore before other competing reactions deplete the oxidant concentrations in the leaching solution. Formerly a Graduate Student with the Department of Geochemistry and Mineralogy, The Pennsylvania State University     

缅甸某砷质金矿中金的氰化浸出的研究

缅甸某砷铜炭质金矿中的金系难浸矿物之一。直接用碳浆工艺氰化浸出金的浸出率为60%。矿物中的铜在氰化过程中易形成Cu(CN)2 沉淀覆盖在矿物表面上,阻碍金的进一步氰化反应。研究中在金矿石中添加助浸剂后,矿石中的一些矿物与之反应,使金的氰化速度增加,浸出率达到了90%-92%。     

Investigation of Leaching Kinetics of Copper from Malachite Ore in Ammonium Nitrate Solutions

The production of metallic copper from low-grade copper ores is generally carried out by hydrometallurgical methods. Leaching is the first prerequisite of any hydrometallurgical process. Solutions containing ammonia may allow for selective leaching of the copper from the ore. In this study, the leaching kinetics of malachite, which is an oxidized copper ore, in ammonium nitrate solutions was examined. The effects of some experimental parameters on the leaching process were investigated, and a kinetic model to represent the effects of these parameters on the leaching rate was developed. It was determined that the leaching rate increased with increasing solution concentration, temperature, and agitation speed, as well as decreasing particle size. It was found that the leaching reaction followed the mixed kinetic controlled model, which includes two different leaching processes including the surface chemical reaction (303?K to 323?K [30?°C to 50?°C]) and diffusion through a porous product layer (323?K to 343?K [50?°C to 70?°C]). The activation energies of these sequential steps were determined to be 95.10 and 29.50?kJ/mol, respectively.     

A REVIEW OF GOLD LEACHING IN ACID THIOUREA SOLUTIONS

Since the initial report of thiourea as a complexing reagent for gold leaching, considerable research has been directed toward the use of thiourea as an alternative to cyanide for gold extraction from different auriferous mineral resources. At the same time, some fundamental investigations of the system have been reported. In this article, a review of both applications and fundamental research is made, including a review of the recent results from laboratory studies at the University of Utah. Recent research results demonstrate that thiourea decomposition is quite slow in the presence of ferric sulfate for simple solutions. Ferric sulfate and formamidine disulfide (FDS) are effective oxidants with fast kinetics. No passivation of the gold surface is observed in simple acidic solutions. However, in actual leaching systems, some sulfide minerals significantly catalyze the redox reaction between thiourea and ferric ion, causing high thiourea consumption if ferric ion is present in excess. The presence of copper has a deleterious influence on this leaching system.     

Pressure leaching of a sulphide copper concentrate with simultaneous regeneration of the leaching agent

An experimental study is presented of leaching of a sulphide copper concentrate with aqueous ferric sulphate under oxygen pressure. The effects of oxygen pressure, sulphuric acid concentration and ferrous iron additions on the copper leaching rate have been determined. Partial pressure of oxygen has been found to govern the rate of copper leaching while the concentration of sulphuric acid only slightly influences this rate. Oxygen enhances the leaching rate mainly by oxidizing ferrous iron to ferric iron — the major leaching agent — and not by direct action on the minerals.Leaching of a sulphide copper concentrate under oxygen pressure with aqueous ferric sulphate leads to a three-fold reduction of both the necessary leaching time and ferric sulphate concentration in the leaching solution over those for conventional leaching with concentrated aqueous ferric sulphate.     

A mathematical model for the solution mining of primary copper ore: Part II. leaching by solution containing oxygen bubbles

The one-dimensional model developed previously to simulate thein situ leaching of copper from deeply-buried low-grade copper ore deposits is used to simulate thein situ operation in which the oxygen-saturated solution containing oxygen bubbles is introduced at the bottom of the chimney. The physical and chemical processes incorporated in the present model include the axial convective transport of mass and heat, axial dispersion of mass, mass transfer between the liquid and gas phases, fluid-solid mass transfer, diffusion of oxygen in the pores of ore fragments, and the dissolution of sulfide minerals. The coupled model equations are solved numerically by an implicit finite-difference method. Calculations have been made for various values of the volume fraction of oxygen bubbles (up to 0.1) in the fluid just downstream of the oxygen sparging nozzle. Calculated results indicate that, for a specific chimney considered, the total amount of copper extracted increases with increasing volume fraction of undissolved oxygen bubbles in the inlet fluid and increasing superficial velocity of the solution (up to 20 m per day). However, a further increase in the superficial velocity of liquid or undissolved oxygen bubbles does not enhance copper extraction. Calculated results also indicate that the total fractional recovery of copper increases with decreasing pyrite to chalcopyrite molar ratio, ore grade, particle size, and shape factor.     

Mineralogical changes occurring during the ferric lon leaching of bornite

The reaction products formed during the leaching of bornite in either ferric chloride or ferric sulfate media depend on the leaching conditions as well as the particle size of the bornite. The extent of dissolution is always more vigorous in the ferric chloride system and increases with increasing temperature in either system. The reaction initially involves the rapid outward diffusion of copper to form slightly nonstoichiometric bornite (Cu_5-xFeS₄), chalcopyrite, and covellite. The non-stoichiometric bornite is progressively converted to a Cu₃FeS₄ phase, which varies considerably in its composition, and to covellite. Although the reaction at low temperature terminates at the Cu₃FeS₄ phase, leaching at higher temperatures results in further dissolution to elemental sulfur and soluble Cu²⁺ and Fe²⁺. The leaching ofmassive bornite illustrates the complexities of the leaching reaction more clearly than is observed for the finelypaniculate bornite. In leached massive bornite, a distinct covellite zone appears in the Cu₃FeS₄ phase; as well, chalcopyrite exsolution lamellae rimmed by a copper sulfide (possibly digenite) appear in the covellite zone, in the Cu₃FeS₄ phase, and in the nonstoichiometric bornite. The experimental leaching results, especially those involving massive bornite, are generally consistent with the mineralogical trends produced by supergene alteration of bornite ores, but a significant difference is that the Cu₃FeS₄ phase does not correspond closely to the mineral idaite.     

Leaching of high-solids,attritor-ground chalcopyrite concentrate byin situ generated ferric sulfate solution

Leaching of attritor-ground chalcopyrite concentrate by acidified ferric sulfate, generatedin situ, has been studied. Air and oxygen alone and mixed with sulfur dioxide were tested as oxidizing agents. A specially designed reactor, which allowedn situ oxidation of ferrous sulfate, permitted the leaching of a 20 pct solids slurry with high copper recovery. The leachate contained ~ 53 g/l copper, a high enough concentration that solvent extraction may not be necessary. formerly with Martin Mariett Laboratories     

Dissolution kinetics of malachite in ammonia/ammonium carbonate leaching

The leaching of oxide copper ore containing malachite, which is the unique copper mineral in the ore, by aqueous ammonia solution has been studied. The effect of leaching time, ammonium hydroxide, and ammonium carbonate concentration, pH, [NH₃]/[NH₄⁺] ratio, stirring speed, solid/liquid ratio, particle size, and temperature were investigated. The main important parameters in ammonia leaching of malachite ore are determined as leaching time, ammonia/ammonium concentration ratio, pH, solid/liquid ratio, leaching temperature, and particle size. Optimum leaching conditions from malachite ore by ammonia/ammonium carbonate solution are found as ammonia/ammonium carbonate concentrations: 5 M NH₄OH+0.3 M (NH₄)₂CO₃; solid/liquid ratio: 1:10 g/mL; leaching times: 120 min; stirring speed: 300 rpm; leaching temperature: 25 °C; particle size finer than 450 μm. More than 98% of copper was effectively recovered. During the leaching, copper dissolves as in the form of Cu(NH₃)₄⁺² complex ion, whereas gangue minerals do not react with ammonia. It was determined that interface transfer and diffusion across the product layer control the leaching process. The activation energy for dissolution was found to be 15 kJ mol⁻¹.     

Leaching of a low-grade chalcocite-covellite ore containing iron in sulphuric acid: The influence of pH and particle size on the kinetics of copper leaching

The kinetics of copper leaching from a chalcocite-covellite ore have been investigated. Leaching was carried out in sulphuric acid solution at different pH values and particle sizes, in the presence of ferric sulphate leached from the ore. Three stages of leaching were established. The apparent activation energy shows a dependence on the period of leaching and the pH value of the solution. An attempt was made to determine the rate-determining steps according to both the activation energies and the analytical expressions for each period.     

某离子型稀土矿浸出试验

以中国南方地区某离子型稀土矿为研究对象,采用搅拌浸出和柱浸的方式,研究不同条件下矿样中稀土及杂质元素的浸出情况,为离子型稀土矿产资源的绿色高效开采提供参考。结果表明,浸出液固比对离子相稀土浸出率影响较大,浸出时间影响较小,离子相稀土浸出过程时间短,反应迅速;柱浸过程中离子相稀土流出速率最快,达到平衡时间短,杂质元素前期浸出浓度高,后续拖尾严重;离子相稀土浸出率随着样品深度的增加不断降低,符合南方离子型稀土成矿规律;硫酸铵浸出过程中铵根离子损失量较大,最低损失率超过11.31%,硫酸根不参与金属离子的交换反应过程,回收率最高可达99.22%。     

Reaction mechanism for the ferric chloride leaching of sphalerite

Reaction mechanisms for the ferric chloride leaching of sphalerite are proposed based on data obtained in leaching and dual cell experiments presented in this work and in a previous study. The results from the leaching experiments show that at low concentrations the rate is proportional to [Fe³⁺]_T ^0.5 and [Cl^-]_T ^0.43 but at higher concentrations the reaction order with respect to both [Fe³⁺]_T and [Cl^-]_T decreases. Using dual cell experiments which allow the half cell reactions to be separated, increased rates are observed when NaCl is added to the anolyte and to the catholyte. The increase in rate is attributed to a direct, anodic electrochemical reaction of Cl^- with the mineral. When NaCl is added only to the catholyte, a decrease in the rate is observed due to a decrease in theE ⁰ of the cathode which is attributed to the formation of ferric-chloro complexes. Several possible electrochemical mechanisms and mathematical models based on the Butler-Volmer relation are delineated, and of these, one model is selected which accounts for the experimentally observed changes in reaction order for both Fe³⁺ and Cl^-. This analysis incorporates a charge transfer process for each ion and an adsorption step for ferric and chloride ions. The inhibiting effect of Fe²⁺ noted by previous investigators is also accounted for through a similar model which includes back reaction kinetics for Fe²⁺. The proposed models successfully provide a theoretical basis for describing the role of Cl^-, Fe³⁺, and Fe²⁺ as well as their interrelationship in zinc sulfide leaching reactions. Possible applications of these results to chloride leaching systems involving other sulfides or complex sulfides are considered.