镍红土矿高压酸浸过程的金属元素浸出行为

以镍、钴的提取为目的,研究褐铁矿型镍红土矿高压酸浸过程中各金属元素的浸出行为,探讨硫酸加入量、浸出温度、浸出时间及液固比对各金属元素浸出率的影响.实验结果表明,在优化条件下Ni、Co、Mn和Mg的浸出率分别达到97%、96%、93%和95%以上,则Fe的浸出率小于1%.对高压浸出渣的分析表明,渣中的铁和硫主要分别以赤铁...     

Mechanism and kinetics of iron extraction from high silica boehmite–kaolinite bauxite by hydrochloric acid leaching

The chemical and mineral compositions of bauxite recovered from the Severoonezhsk Bauxite Mine (Arkhangelsk region, Russia) were studied by XRD, ICP-OES, TG/DSC, SEM, TEM, and Mössbauer spectroscopy. The iron-containing minerals of the bauxites were found to comprise alumogoethite (α-Fe_1–xAl_xOOH), alumohematite (α-(Fe_1–xAl_x)₂O₃), alumoakaganeite (β-Fe_1–xAl_xO(OH,Cl)), and chromite (FeCr₂O₄). The efficiency of Fe extraction from the bauxite by HCl leaching was 82.5% at 100 °C, HCl concentration of 10%, solid/liquid ratio of 1:10, and the process duration of 60 min, with aluminum loss from the bauxites below 4.5% of the total Al contents in the bauxite. Analysis of the kinetics of the iron leaching process proved diffusion to be the limiting stage of the process at 90–100 °C. Bauxite residue after leaching presented traces of α-Fe_1–xAl_xOOH and β-Fe_1–xAl_xO(OH,Cl), and most of the iron content was in the FeCr₂O₄. In bauxite residue after HCl leaching, in addition to iron oxide, the contents of chromium and calcium oxides significantly decreased. The iron chloride liquor after leaching contained the rare earth elements (REE) of 6.8 mg/L Sc, 4.1 mg/L Ce and 2.3 mg/L Ga.     

Characterization and kinetic study on ammonia leaching of complex copper ore

Ammonia leaching kinetics of a complex Cu-ore assaying 8.8% Cu and 36.1% Fe was examined. Mineralogical characterization indicated that the major phase of the ore was siderite with chalcopyrite as the major sulfide mineral. The effects of parameters such as agitation, temperature, NH₃ concentration, particle size and oxygen partial pressure (p_O2) were investigated. Under the standard leaching conditions of 125–212 µm particle size, 120 °C, 1.29 mol/L NH₃ and 202 kPa of p_O2, about 83% Cu could be selectively extracted in 2.5 h. However, when using higher NH₃ concentration and lower particle size, more than 95% extraction was achieved. The leaching process was found to be surface reaction controlling. The estimated activation energy was (37.6±1.9) kJ/mol and empirical orders of reaction with respect to p_O2 and [NH₃] were about 0.2 and 1, respectively.     

Microwave assisted atmospheric acid leaching of nickel from laterite ore

The recovery of nickel from laterite ore with sulphuric acid under the effect of microwave irradiation was studied.The experimental results indicated that the extraction rate of nickel was influenced by reaction time,sulphuric acid concentration,and temperature,especially by microwave power.The results obtained from the experiments of orthogonal arrays showed that the optimum conditions of sulphuric acid concentration,reaction time,microwave power,and temperature were 25 vol.%,1.5 h,600 W,and 90°C,respectively.Under the optimal conditions,the nickel recovery could reach approximately 90.8%,which was higher than that obtained by conventional water bath heating.Kinetic experiments showed that the leaching of nickel in a sulphuric acid medium was controlled by chemical reaction occurring on the surface of laterite ore.The apparent activation energy was 38.9 kJ/mol.Microwave heating technology is efficient,clean,and easy to control and facilitate continuous processing of materials.     

Extraction of molybdenum and nickel from roasted Ni–Mo ore by hydrochloric acid leaching,sulphation roasting and water leaching

To extract molybdenum and nickel from the roasted Ni–Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leaching rate of Mo and Ni. Under the optimum conditions of hydrochloric acid leaching (roasted Ni–Mo ore leached with 0.219 mL/g hydrochloric acid addition at 65 °C for 30 min with a L/S ratio of 3 mL/g), sulphation roasting (51.9% sulfuric acid addition, roasting temperature 240 °C for 1 h), followed by leaching with the first stage hydrochloric acid leaching solution at 95 °C for 2 h, the leaching rates of Mo and Ni reached 95.8% and 91.3%, respectively.     

Stepwise recovery of magnesium from low-grade ludwigite ore based on innovative and clean technological route

A novel and clean technological route for the comprehensive utilization of low-grade ludwigite ore was proposed, in which magnesium was extracted by metallizing reduction-magnetic separation, sulfuric acid leaching and ethanol precipitation operation. Meanwhile, iron-rich product, silicon-rich product and boron-rich product were obtained, respectively. In the process of metallizing reduction-magnetic separation, 94.6% of magnesium was enriched in the non-magnetic substance from the ore reduced at 1250 °C for 60 min with the ore size of 0.50–2.00 mm and coal size of 0.50–1.50 mm. When the non-magnetic substance was leached at 90 °C for 15 min with the liquid-to-solid ratio of 7:1, 87.4% of magnesium was leached into the liquor separated from silicon gathering in leaching residue. The ethanol precipitation was conducted for 30 min with the ethanol-to-original liquid volume ratio of 1.5:1 at room temperature. 97.2% of magnesium was precipitated out with the initial concentration of 0.8 mol/L in the form of MgSO₄·7H₂O.     

Kinetics of leaching lithium from α-spodumene in enhanced acid treatment using HF/H2SO4 as medium

An enhanced leaching of Li from α-spodumene was carried out using a mixture of hydrofluoric and sulfuric acid (HF/H₂SO₄) as the medium. Based on the optimized leaching conditions, the leaching kinetics of Li was investigated in an ore/HF/H₂SO₄ ratio of 1:3:2 g:mL:mL with leaching temperature ranging from 50 to 100 °C. The results indicate that the leaching kinetics of Li fitted well with a model based on the shrinking core model. In addition, the leaching rate of Li was controlled by chemical reactions and diffusion through the product layers. The apparent activation energy E_a was calculated to be 32.68 kJ/mol. Solid films were formed because of the generation of insoluble products such as cryolithionite (Na₃Li₂Al₂F₁₂), cryolite (Na₃AlF₆), calcium fluoride (CaF₂), potassium cryolite (K₂AlF₅), aluminum fluoride (AlF₃), and fluorosilicates (Na₂SiF₆ or KNaSiF₆). Furthermore, the effects of the ore/HF ratio and leaching temperature on the leaching behavior of Li, Al and Si were investigated. The results indicate that the ore/HF ratio and leaching temperature could clearly affect the distribution of HF molecules on the leaching of Li, Al and Si, which are important for the selective leaching of Li over Al and Si with this fluorine-based chemical method.     

焙烧氟碳铈矿硫酸浸出稀土的动力学(英文)

研究了硫酸浸出德昌稀土与天青石共伴生矿的焙烧矿过程。考查粒度、搅拌速度、硫酸浓度和温度对稀土浸出率的影响,并对稀土的浸出动力学进行分析。在选定的浸出条件下:粒径0.074～0.100mm、硫酸浓度1.5mol/L、液固比8:1、搅拌速度500r/min,稀土浸出反应受内扩散控制,表观活化能为9.977kJ/mol。     

Dechromization and dealumination kinetics in process of Na2CO3-roasting pretreatment of laterite ores

A novel process was proposed for the activation pretreatment of limonitic laterite ores by Na₂CO₃ roasting. Dechromization and dealumination kinetics of the laterite ores and the effect of particle size, Na₂CO₃-ore mass ratio, and roasting temperature on Cr and Al extraction were studied. Experimental results indicate that the extraction rates of Cr and Al are up to 99% and 82%, respectively, under the optimal particle size of 44–74 μm, Na₂CO₃-to-ore mass ratio of 0.6:1, and temperature of 1000 °C. Dechromization within the range of 600–800 °C is controlled by the diffusion through the product layer with an apparent activation energy of 3.9 kJ/mol, and that it is controlled by the chemical reaction at the surface within the range of 900–1100 °C with an apparent activation energy of 54.3 kJ/mol. Besides, the Avrami diffusion controlled model with on apparent activation energy of 16.4 kJ/mol is most applicable for dealumination. Furthermore, 96.8% Ni and 95.6% Co could be extracted from the alkali-roasting residues in the subsequent pressure acid leaching process.     

Pressure nitric acid leaching of alkali-pretreated low-grade limonitic laterite

The pressure nitric acid leaching of alkali-pretreated low-grade limonitic laterite, as well as removing impurity Al(III) and preparing intermediate product of nickel/cobalt sulphide from leaching liquor were investigated. After pretreatment, iron exists in the form of amorphous iron oxides, while nickel is adsorbed on the surface of iron oxides in the form of nickel oxide. The preferable pressure leaching conditions are determined as follows: leaching temperature of 458 K, leaching duration of 60 min, initial acidity of nitric acid of 1.90 mol L-1and liquid to solid ratio of 3:1(volume to mass ratio). Under these conditions, the leaching efficiencies of Ni, Co and Al are 95 %, 88 % and 55 %, respectively, and that of Fe is less than 1 %. The loss rates of Ni and Co are 1.8 % and1.5 %, respectively, during the step of removing impurity Al(III). The sulphide precipitation process produces the interim production of nickel/cobalt sulphides, recovering greater than 99 % of Ni and Co in the purified solution.The iron-rich([60 %) pressure leaching residue with low Cr, S can be further reclaimed as the raw materials for iron making.     

高压酸浸镍钼矿提取钼和镍

研究一项针对镍钼矿用高压酸浸的方法回收镍和钼的全湿法工艺。采用该工艺避免了传统上艺焙烧镍钼矿（15％～25％s）带来的人量S02和As2O3排放,减小了对环境的污染;与现有的湿法碱浸回收钼工艺相比,本工艺存酸浸过程中回收了儿乎全部的镍和人部分的钼。在氧压环境下,几乎全部的镍和大部分的钼都进入溶液,少部分的钼留在酸浸渣中,睃浸渣进一步用碱（NaOH）浸出。在最佳的实验条件下,97％的镍和96％的钼分别被浸出。     

Dynamics in simultaneous electro-generative leaching for sphalerite-MnO2

The principle for the electro-generative leaching was applied to simultaneous leaching of sphalerite-MnO2. A galvanic system for the bio-electro-generative leaching was set up. The effects of grain size and temperature on rate of zinc extraction fi＇om sphalerite under the conditions of presence and absence ofAcidithiobacillusferrooxidans （A.ferrooxidans） were studied, respectively. The results show that with bacteria, the maximum extraction of zinc fi＇om the ores with grain size of 16.6 μm can reach 32.01% after leaching for 12 h, while to obtain the same extraction ratio in the traditional bio-leaching route （i.e. not electro-generative one）10 d is needed to ore granules with same size. The unreacted shrinking core model was used for describing the reaction-relative and diffusion-relative phenomena presented in the process of the electro-generative leaching with and without bacteria, which is considered to be diffusion controlled. The activation energies of the anodic reaction for leaching system in the presence and absence of bacteria are 11.97 and 14.39 kj/mol, respectively, indicating that leaching rate can be decreased by A. ferrooxidans. SEM was used to study the effect ofA. ferrooxidans on the ores in the simultaneous electro-generative leaching, which indicates that the produced sulfur on the surface of the sulfides can be oxidized by A. ferrooxidans after bio-electro-generative leaching for 24 h, and the transferred charge due to the bacterial oxidation is up to 17.86%, which is an important part of the output electric quantity.     

Studies on the leaching of tungsten from composite barite–scheelite concentrate

Studies on the recovery of tungsten from a composite barite–scheelite concentrate ore were conducted utilizing a soda–silicon roast-Na₂CO₃ solution leach technique. Scheelite and barite mainly compose this ore with a slight trace of other impure components. A high tungsten recovery ratio is not achievable by physical separation by beneficiation or caustic autoclaving procedures as a significant amount of scheelite and barite occur in an interspersed state. Soda–silicon roast-Na₂CO₃ solution leach is then proposed to extract tungsten from this special ore with effects of the roasting and leaching process tested under variables such as roasting time, temperature, Na₂CO₃ amount and leaching conditions. Approximately 95.93% of W was dissolved into the solution under optimum roasting conditions of temperature 850 °C, reactive time 2 h, molar ratio of Na₂CO₃/W 3 and 10% SiO₂ dosage of feed mass with leaching conditions of temperature 60 °C, stirrer speed 300 rpm, leaching time 1 h, liquid/solid ratio 4 and in 10 g/l Na₂CO₃ solution.     

Recovery of waste rare earth fluorescent powders by two steps acid leaching

The effects of the acid leaching and alkali fusion on the leaching efficiency of Y,Eu,Ce,and Tb from the waste rare earth fluorescent powders were investigated in this paper.The results show that hydrochloric acid is better than sulfuric acid in the first acid leaching,and NaOH is better than Na2CO3in the alkali fusion.In the first acid leaching,the Wloss is 20.94%when the waste rare earth fluorescent powders are acid leached in H?concentration 3 mol L-1and S/L ratio 1:3 for 4 h due to red powders dissolved.The better results of the alkali fusion can be got at 800℃ for 2 h when the NaOH is used.The blue powders and the green powders can be dissolved into NaAlO2and oxides such as rare earth oxide(REO).The REO can be dissolved in H?concentration 5 mol L-1,S/L1:10 for 3 h in the second acid leaching.The leaching rates of the Y,Eu,Ce,and Tb are 99.06%,97.38%,98.22%,and 98.15%,respectively.The leaching rate of the total rare earth is 98.60%.     

Leaching of cuprite through NH4OH in basic systems

Cuprite is a difficult oxide to leach under acidic conditions (for the maximum extraction of 50%). In this research, the feasibility of leaching cuprite in an ammoniacal medium was studied. The working conditions addressed here were the liquid/solid ratio (120:1–400:1 mL/g), stirring speed (0–950 r/min), temperature (10–45 °C) and NH₄OH concentration (0.05–0.15 mol/L). In addition, different ammoniacal reagents (NH₄F and (NH₄)₂SO₄) were analyzed. The experiments were performed in a 2 L reactor with a heating mantle and a condenser. The most important results were that the maximum leaching rate was obtained at pH 10.5, 0.10 mol/L NH₄OH, 45 °C, 4 h, 850 r/min and a liquid/solid ratio of 400:1, reaching a copper extraction rate of 82%. This result was related to the non-precipitation of copper in solution by the formation of copper tetra-amine. The liquid/solid ratio and stirring speed were essential for increasing the cuprite leaching. The maximum leaching rate was achieved at higher temperatures; however, significant copper leaching rate occurred at temperatures near the freezing point of water (17.9% over 4 h). Increasing NH₄OH concentration and decreasing particle size increased the cuprite leaching rate. The two ammoniacal reagents (NH₄F and (NH₄)₂SO₄) had low extraction rate of copper compared with NH₄OH. The kinetic model representing cuprite leaching was a chemical reaction on the surface. The order of the reaction with respect to the NH₄OH concentration was 1.8, and it was inversely proportional to the radius of the ore particles. The calculated activation energy was 44.36 kJ/mol in the temperature range of 10–45 °C.     

Copper and elemental sulphur from chalcopyrite by pressure leaching

Processes employing direct oxidation under an over-pressure of air or oxygen in an aqueous sulphuric acid medium have been developed in the Sherritt Gordon Laboratories for iron, nickel, cobalt, zinc and lead sulphide concentrates. This study has recently also been extended to chalcocite, Cu₂S, concentrates. The rising interest in processes employing direct aqueous oxidation is stimulated by the fact that elemental sulphur can be produced as a by-product rather than sulphur dioxide or sulphuric acid.The present paper outlines a process which features the direct pressure oxidation of the most abundant copper sulphide mineral, chalcopyrite, CuFeS₂. The optimum conditions for a practical pressure leaching step have now been developed in the laboratory which results in the production of copper sulphate solution suitable for copper winninq by electrolysis, hydrogen reduction, solvent extraction combined with electrolysis, or other means. The leach residue yields pure elemental sulphur by-product. Copper and elemental sulphur recoveries of 98 and 85% respectively have been recorded. The fastest oxidation rate, corresponding to a leach retention time of 2.5 hr, was obtained when the copper concentrate was ground to 99.5% — 325 mesh, when a 50% stoichiometric excess of concentrate over the amount of available sulphuric acid for copper was used and when the oxygen partial pressure and temperature were maintained at 500 psi and 240°F, respectively. In an idealized form, the pressure leaching reaction can be expressed as follows:—CuFeS₂ + H₂SO₄ + 1 1/40₂ + 1/2H₂O → CuSO₄ + Fe(OH)₃ + 2S°After separation of the copper sulphate solution by filtration, elemental sulphur and excess concentrate ore recovered from the iron oxide tailing by flotation. The tailing, containing iron oxide and insolubles, is rejected. The elemental sulphur is separated from the concentrate by hot filtration, solvent extraction, distillation, or other means, and the unleached chalcopyrite is recycled to the leaching step.     

采用Na2SO3溶液从硒渣中选择性浸出Se及其动力学

采用SO2还原沉金后液制得硒渣,再用Na2SO3选择性浸出硒渣,使Se得到有效分离;通过研究浸出过程中Se浸出率随时间的变化,建立该反应的动力学方程,确定Na2SO3溶液浓度、液固比、搅拌速度及反应温度对Se浸出率的影响,并计算相应的表观活化能。结果表明:增加Na2SO3溶液浓度和升高反应温度可以较大幅度提高Se的浸出率,液固比和搅拌速度对浸出Se的影响较小;Na2SO3浸出Se过程为Avrami模型混合控制,其特征参数n和表观活化能E分别为0.235和20.847 kJ/mol,Se的浸出率受反应温度的影响较大。     

Extraction of manganese and zinc from their compound ore by reductive acid leaching

Comprehensive utilization of low grade manganese–zinc compound ore containing lead and silver with a method of reductive acid leaching was studied. According to the ?–pH diagram of Mn–Zn–H₂O system, Mn and Zn can be leached simultaneously in the pH range of –2 to 5.61. The results showed that both hydrogen peroxide and sucrose were effective reductants which could intensify the simultaneous leaching of Mn and Zn into leachate as well as enrich Pb and Ag in the residue. 95.88% of Mn and 99.23% of Zn were extracted when the compound ore was leached with hydrogen peroxide in sulfuric acid media, meanwhile the contents of Pb and Ag in the residue were enriched to 13.21% and 489.36 g/t, respectively. When sucrose was used as the reductant, the leaching efficiencies of Mn and Zn separately achieved 98.26% and 99.62%, and contents of Pb and Ag in the residue were as high as 13.92% and 517.87 g/t, respectively.     

Kinetics of simultaneous leaching of Ag and Pb from hydrometallurgical zinc residues by chloride

The leaching kinetics of silver and lead simultaneously from zinc residue by chloride was investigated.The effects of stirring speed,temperature,sodium chloride concentration,particle size and liquid/solid ratio on Ag and Pb dissolution in sodium chloride were studied.It was determined that the dissolution rates increased with increasing sodium chloride concentration,temperature and decreasing particle size.The dissolution kinetics followed a shrinking core model,with inter-diffusion through gangue layer as the rate determining step.This finding is in accordance with the apparent activation energy(Ea)of 26.8 kJ.mol-1(Ag)and 26.5 kJ.mo1-1(Pb),and a linear relationship between the rate constant and the reciprocal of squared particle size.The orders of reaction with respect to sodium chloride concentration,temperature and particle size were also achieved.The rate of reaction based on diffusion-controlled process can be expressed by semi-empirical equations.     

Modification of the SOI-like structures by annealing under high hydrostatic pressure

The effect of enhanced hydrostatic pressure during annealing (HP-HT treatment) on creation/transformation of SiO₂ in silicon implanted with oxygen (Si:O, oxygen doses 1×10¹⁴–1×10¹⁶ cm⁻²) was studied by infrared absorption, photoluminescence and X-ray (synchrotron) methods. The redistribution of oxygen atoms at HP-HT in the Si:O structures, prepared by implantation with low doses (down to 1×10¹⁴ cm⁻²) of oxygen ions, leads to generation of separated clusters having the structure of a SiO_2−x network. Enhanced HP stimulates the creation of these clusters. A qualitative explanation of presented results, stressing the role of vacancies in the process of oxygen atoms agglomeration and the influence of compressive stress on the concentration of vacancies, is proposed.