铟的还原浸出动力学及分步沉淀选择性分离（英文）

开展硫化锌精矿还原浸出高铁锌浸出渣高效浸铟及浸出液中铟选择性分离的研究。结果表明：在固体物料粒度74～105μm、反应温度90℃、浸出时间300 min、硫酸浓度1.4 mol/L的条件下,铟的浸出率达95%以上。采用收缩核模型对还原浸出动力学进行分析,不同条件下的浸出实验结果表明反应受穿过固体产物层的扩散控制,活化能为17.96 k J/mol,相对于硫酸浓度的反应级数为2.41。铁粉置换沉铜过程铜和砷的沉淀率均达99%以上。98%以上的铟从含高亚铁离子浓度的硫酸锌溶液中选择性分离,获得铟含量约为2.4%的富铟渣,经酸浸-萃取-电积工艺流程进一步处理后可得到纯铟。     

硫化锌银精矿富集银的工艺研究

基于国内外硫化锌矿处理的火、湿法研究进展,对含锌银精矿采用硫酸化焙烧、稀硫酸浸出工艺脱除锌、富集银,考察了焙烧和浸出过程中的主要影响因素。结果表明,硫酸配比为150%,在300℃焙烧90 min,以5%稀硫酸为浸出剂,液固比8:1,搅拌转速200~300 r/min,85℃浸出120 min,最终锌的浸出率可达到98%以上,浸出渣中银含量为7.24%,银被富集7倍。     

Kinetic study on pressure leaching of high iron sphalerite concentrate

The kinetics of pressure leaching high iron sphalerite concentrate was studied.The effects of agitation rate,temperature, oxygen partial pressure,initial acid concentration,particle size,iron content in the concentrate and concentration of Fe2 added into the solution on the leaching rate of zinc were examined.The experiment results indicate that if the agitation rate is greater than 600 r/min,its influence on Zn leaching rate is not substantial.A suitable rise in temperature can facilitate the leaching reaction,and the temperature should be controlled at 140-150℃.The increase trend of Zn leaching rate becomes slow when pressure is greater than 1.2 MPa,so the pressure is controlled at 1.2-1.4 MPa.Under the conditions of this study,Zn leaching rate decreases with a rise in the initial sulfuric acid concentration;and Zn leaching rate increases with a rise of iron content in the concentrate and Fe 2 concentration in the solution.Moreover,the experiment demonstrates that the leaching process follows the surface chemical reaction control kinetic law of“shrinking of unreacted core”.The activation energy for pressure leaching high iron sphalerite concentrate is calculated,and a mathematical model for this pressure leaching is obtained.The model is promising to guide the practical operation of pressure leaching high iron sphalerite concentrate.     

湿法炼锌净化渣酸浸的动力学

对湿法炼锌净化渣的浸出动力学进行了研究,并探讨了硫酸浓度、反应温度、粒度等对钴、锌浸出率的影响规律。从动力学的角度分析了整个浸出过程,得到优化条件：液固比50：1(mL/g),硫酸浓度100 g/L,反应温度70°C,粒度75~80μm,反应时间20 min。在此优化条件下钴的浸出率为99.8%,锌的浸出率为91.97%。结果表明：在硫酸体系中钴的浸出符合不生成固体产物层的“未反应收缩核”模型。通过 Arrhenius 经验公式求得钴和锌表观反应活化能分别为11.693 kJ/mol和6.6894 kJ/mol,这表明浸出过程受边界层扩散控制。     

Recovery of gallium from zinc concentrate by pressure oxygen leaching

Zinc concentrate with high gallium content is one of the main resources of gallium.The gallium presents in the form of isomorphism in tetrahedron coordination with sulfur in sphalerite.The research was to investigate the amenability of zinc concentrate with high gallium to pressure oxygen leaching.The particle size,sulfuric acid concentration,oxygen partial pressure,additive amount,and time of reaction were studied.The extraction yields of gallium and zinc are 86%and 98%,respectively.The optimal condition is 100 g of zinc concentrate with particle size smaller than 38 lm,sulfuric acid concentration150 g L-1,leaching temperature 150℃,leaching time120 min,oxygen partial pressure 0.7 MPa,additive amount of 0.2 wt%.     

APPLICATIN OF ATTRITION GRINDING IN ACID LEACHING OF NICKEL SULFIDE CONCENTRATE

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复杂铜铅锌银混合精矿两段逆流氧压浸出工艺

通过对呷村铜铅锌银多金属混合精矿的矿物分析可知,铅、锌主要赋存于方铅矿、闪锌矿中,而大部分铜、银均为难浸出的黝铜矿所载.采用两段逆流氧压浸出工艺进行处理该矿,条件试验研究得出一、二段最佳浸出条件分别为硫酸浓度150 g/L、液固比3:1、反应温度135℃、氧分压O.75 MPa、浸出时间2.0 h;硫酸浓度80 g/L、液固比3:1、反应温度180℃、氧分压1.0 MPa、浸出时间2.5 h.9轮扩大验证循环浸出试验运行顺利,Cu和Zn的平均两段总浸出率分别为93.23%和99.47%,而杂质元素Fe和As的浸出率仅15.77%和6.9%,元素硫的硫磺转化率为54.26%.Pb和Ag大部分转化为铅矾、铅铁矾和硫化银而留在浸出渣中,铜锌与铅银彻底分离.     

An investigation of oxygen pressure acid leaching of Gacun complex Cu-Pb bulk concentrate

The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The primary copper and lead minerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in the mineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.％ and 94.9 wt.％,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.     

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。     

Kinetics of indium dissolution from marmatite with high indium content in pressure acid leaching

The kinetics of oxygen pressure acid leaching marmatite with high indium content was studied. The effects of particle size, agitation rate, temperature, H_2SO_4 concentration, and oxygen partial pressure on leaching rate of indium were investigated. The results show that when the agitation rate is above 600 r-min~(-1), its influence on indium leaching rate is insignificant. It is determined that the leaching rates increase with the increase in sulfuric acid concentration, temperature, partial oxygen pressure, and the decrease in particle size. Moreover, the results demonstrate that the process of indium leaching is controlled by interface chemical reaction, with apparent activation energy of 65.7 k J-mol~(-1). The apparent reaction orders of sulfuric acid and oxygen partial pressure are determined to be 0.749 and 1.260, respectively. The leaching reaction process follows shrinking unreacted core model. And finally, the kinetics model equation is established for indium.     

The hydrometallurgical extraction of lithium from egyptian montmorillonite-type clay

The processing of El-Fayoum montmorillonite-type clay deposits is attained through leaching with commercial sulfuric acid using a ball-mill-type autoclave. This process yields lithium sulfate, which can be used either for the production of lithium carbonate or to produce lithium metal. The effects of temperature, grain size, and sulfuric acid concentration and leaching on lithium recovery as well as the kinetics of the leaching process have been studied.     

Pressure leaching of zinc silicate ore in sulfuric acid medium

Zinc silicate ore was characterized mineralogically and the results showed that zinc exists mainly as hemimorphite and smithsonite in the sample. Sulfuric acid pressure leaching of zinc silicate ore was carried out to assess the effect of particle size, sulfuric acid concentration, pressure, reaction time and temperature on the extraction of zinc and the dissolution of silica. Under the optimum conditions employed, up to 99.25% of zinc extraction and 0.20% silica dissolution are obtained. The main minerals in leaching residue are quartz and small amounts of undissolved oxide minerals of iron, lead and aluminum are associated with quartz.     

The sulfuric acid leaching of Bayer electrofilter fines: A practical kinetical approach

Electrofilter fines, which are by-products of the Bayer process for the production of alumina from bauxite, were characterized to evaluate the alumina that was potentially extractable with sulfuric acid. Acid leaching is carried out at different concentrations of sulfuric acid, at different temperatures, pulp densities, and times, to dissolve gibbsite and transition aluminas. The result is an aluminum sulfate solution. This article reports on a study of the kinetics of the leaching reaction at 90°C with two pulp densities: 10% and 30%.     

氟碳铈矿盐酸浸出过程的动力学

考察盐酸浓度、浸出温度、液固比和矿物粒度对浸出速度的影响.结果表明:在盐酸浓度6 mol/L、浸出温度90C、液固比15:1、矿物粒度25 μm的条件下浸出90 min后,氟碳铈矿中稀土碳酸盐的浸出率达到89.6％,而稀土氟化物的浸出率仪为1.5％.氟碳铈矿浸出过程符合产物层扩散模型,稀土碳酸盐和氟化物浸出过程的表观活...     

Reductive acid leaching of valuable metals from spent lithium-ion batteries using hydrazine sulfate as reductant

Hydrazine sulfate was used as a reducing agent for the leaching of Li, Ni, Co and Mn from spent lithium-ion batteries. The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and examined. 97% of the available Li, 96% of the available Ni, 95% of the available Co, and 86% of the available Mn are extracted under the following optimized conditions: sulfuric acid concentration of 2.0 mol/L, hydrazine sulfate dosage of 30 g/L, solid-to-liquid ratio of 50 g/L, temperature of 80 °C, and leaching time of 60 min. The activation energies of the leaching are determined to be 44.32, 59.37 and 55.62 kJ/mol for Li, Ni and Co, respectively. By performing X-ray diffraction and scanning electron microscopy in conjunction with energy dispersive X-ray spectroscopy, it is confirmed that the main phase in the leaching residue is MnO₂. The results show that hydrazine sulfate is an effective reducing agent in the acid leaching process for spent lithium-ion batteries.     

DESIGN OF AN EXPERT CONTROL SYSTEM FOR LEACHING PROCESS

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Wolframite Conversion in Treating a Mixed Wolframite–Scheelite Concentrate by Sulfuric Acid

Complete wolframite conversion in sulfuric acid is significant for expanding the applicability of the sulfuric acid method for producing ammonium paratungstate. In this paper, the conversion of wolframite in treating a mixed wolframite–scheelite concentrate by sulfuric acid was studied systematically. The results show that the conversion of wolframite in sulfuric acid is more difficult than that of scheelite, requiring rigorous reaction conditions. A solid H₂WO₄ layer forms on the surfaces of the wolframite particles and becomes denser with increasing H₂SO₄ concentration, thus hindering the conversion. Furthermore, the difficulty in wolframite conversion can be mainly attributed to the accumulation of Fe²⁺ (and/or Mn²⁺) in the H₂SO₄ solution, which can be solved by reducing Fe²⁺ (and/or Mn²⁺) concentration through oxidization and/or a two-stage process. Additionally, the solid converted product of the mixed wolframite–scheelite concentrate has an excellent leachability of tungsten in an aqueous ammonium carbonate solution at ambient temperature, with approximately 99% WO₃ recovery. This work presents a route for manufacturing ammonium paratungstate by treating the mixed concentrate in sulfuric acid followed by leaching in ammonium carbonate solution.     

硫酸铵焙烧法浸出镍磁黄铁矿中有价金属

将镍磁黄铁矿与(NH4)2SO4混合后在高温下焙烧,考察(NH4)2SO4用量、浸出温度、浸出时间和稀硫酸浸出液的pH值对焙烧产物中金属元素浸出率的影响,并在氨水-(NH4)2SO4混合溶液中浸出焙烧产物。结果表明:在不同情况下,Ni和Cu的浸出率较高,Mg和Fe的浸出率较低;氨性溶液有利于Ni和Cu的浸出,总氨浓度为7 mol/L时,Ni和Cu的浸出率分别为89.56%和79.35%;低pH值的稀硫酸溶液有利于Mg和Fe的浸出,pH值为0.5时,Mg和Fe的浸出率分别为61.39%和62.56%。由扫描电镜-能谱分析和XRD分析可知,矿样中Ni和Cu大部分被浸出;由于焙烧产物中部分Mg和Fe以铁酸钙和硅酸镁等形态存在,Mg和Fe的浸出率较低。     

Recovering indium with sulfating roasting from copper-smelting ash

A technology for recovering indium from Jinchuan copper-smelting ash was developed. Indium in the ash was first enriched to the leaching-slag in leaching process,and then recovered by sulfating roasting. The method included mixing the leaching-slag with sulfuric acid,making them into particles,roasting the mixture,and then leaching the calcine with hot water. Above 90% of indium in calcine could be dissolved into the leaching solution. The optimized conditions were determined as follows： the mass ratio of sulfuric acid to leaching slag was 0.1,the roasting time was about 1 to 1.5 h in the temperature range of 200-250℃,and the calcine was leached for 1 h with 5：1 of liquid/solid ratio at 60℃. Over 99% of indium in leaching solution was finally enriched by Zn substitution or sulfide precipitation.     

失效双氧水催化剂中钯的浸出-沉淀富集效率研究

研究了从失效双氧水催化剂中富集钯的工艺,采用“焙烧浸出-硫化沉钯”的工艺实现了钯的高效富集。重点考察在浸出中硫酸浓度、浸出温度、浸出时间、盐酸浓度、NaClO3用量对富集倍数的影响。研究了富集钯的最佳工艺条件,结果表明,钯平均富集12.09倍,钯的平均回收率为99.55%。