Mechanism of influence of chloride ions on electrogenerative leaching of sulfide minerals

1 INTRODUCTIONElectrogenerative leaching process is a newtechnique in hydrometallurgy .ZHANG et al[1]in-troduced the principle and technique of electrogen-erative process to metallurgy field by the leachingof synthetic Ni3S2with FeCl3.In order to utilizethe chemical energy in leaching process reasonablyand si mplify the purifying process , Wang et al[2 8]studied the electrogenerative leaching of a series ofsulfide minerals through a dual cell system withFeCl3and acidic MnO2as oxidant…     

Simultaneous electrogenerative leaching of chaicopyrite concentrate and MnO2

In order to enhance the electrogenerative leaching rate of chalcopyrite concentrate reasonably, the principle of generative process was applied to simultaneous leaching of chalcopyrite concentrate and MnO2. The results show that Cu^2＋ and Mn^2＋ in addition to electrical energy could be acquired in the simultaneous electrogenerative leaching process. The leaching cell has the open circuit potential of about 1.0 V and gains quantity of electricity of about 700 C. The optimum leaching rates of Cu^2＋ and Mn^2＋ are 23.10% and 22.1%, respectively after electrogenera- tive leaching for about 10 h under the present conditions.     

Sodium sulfide leaching of low-grade jamesonite concentrate in production of sodium pyroantimoniate

Sodium sulfide leaching of a low-grade jamesonite concentrate in the production of sodium pyroantimoniate through the air oxidation process and the influencing factors on the leaching rate of antimony were investigated.In order to decrease the consumption of sodium sulfide and increase the concentration of antimony in the leaching solution, two-stage leaching of jamesonite concentrate and combination leaching of high-grade stibnite concentrate and jamesonite concentrate were used. The experimental results show that the consumptions of sodium sulfide for the two-stage leaching process and the combination leaching process are decreased by 20% and 60% compared to those of one-stage leaching process respectively. The final concentrations of antimony in the leaching solutions of both processes are above 100g/L.     

氨络合物体系中镍阴极电沉积的电化学行为

通过极化曲线测量，对氨络合物体系中镍阴极电沉积电化学行为进行研究，系统探讨了溶液中总镍离子浓度、氨水浓度、氯化铵浓度、阴离子及温度等工艺条件对镍阴极还原的影响。研究结果表明：镍放电电流随着总镍离子浓度的上升而上升，随氨水浓度的升高而降低；在1—4mol／L氯化铵浓度范围内，镍放电电流随其浓度的降低而升高，而当氯化铵浓度低于1mol／L时，镍放电电流出现下降的现象；氯盐氨络合物体系中镍阴极放电电流明显高于硫酸盐氨络合物体系镍放电电流，镍放电电流随温度的升高而升高。根据实验现象，进一步分析了镍阴极电沉积电化学行为变化的原因。     

中等嗜热菌浸出高砷铜精矿研究

高砷铜精矿因含砷较高存在砷害问题,限制了其利用.论文针对云南某高砷硫化铜精矿,采用某中等嗜热菌S.P进行浸出,对比研究了精矿粒度、浸出方式、矿浆浓度、浸出时间和菌液初始Fe3+等因素对浸矿过程的影响.在最佳浸矿条件下中等嗜热菌S.P浸矿时Cu,As和Fe浸出率分别为82.39%,78.21%和40.38%.此外,试验表明高浓度的初始Fe3+显著促进铜精矿中铜、砷的浸出,在初始Fe3+浓度为0.08~0.32 mol/L时,铜浸出率为86.34%~97.06%,As浸出率为89.22%~94.13%.浸渣的X射线衍射结果表明中等嗜热菌S.P浸矿过程中生成单质硫和少量砷酸铜.研究为该类矿的生物冶金处理提供了一定的研究基础,对高砷硫化铜精矿资源的开发利用具有重要意义.     

Extraction of valuable metals from low-grade nickeliferous laterite ore by reduction roasting-ammonia leaching method

Nickel and cobalt were extracted from low-grade nickeliferous laterite ore using a reduction roasting-ammonia leaching method. The reduction roasting-ammonia leaching experimental tests were chiefly introduced, by which fine coal was used as a reductant. The results show that the optimum process conditions are confirmed as follows: in reduction roasting process, the mass fraction of reductant in the ore is 10%, roasting time is 120 min, roasting temperature is 1 023–1 073 K; in ammonia leaching process, the liquid-to-solid ratio is 4:1(mL/g), leaching temperature is 313 K, leaching time is 120 min, and concentration ratio of NH₃ to CO₂ is 90 g/L:60 g/L. Under the optimum conditions, leaching efficiencies of nickel and cobalt are 86.25% and 60.84%, respectively. Therefore, nickel and cobalt can be effectively reclaimed, and the leaching agent can be also recycled at room temperature and normal pressure.     

Electric-oxidation extraction of molybdenite concentrate in alkaline NaCl electrolyte

The dissolution of molybdenite concentrate in NaCl electrolyte was investigated. The results show that the dissolution rate increases with the increase in liquid-to-solid ratio, stirring speed, NaCl concentration and temperature. When the liquid-to-solid ratio is 30:1, stirring speed is 400 r/min, concentration of NaCl is 4 mol/L at pH=9 and room temperature, the leaching efficiency of molybdenite concentrate will reach 99.5% in 240 min. Molybdenite concentrate cannot be electro-oxidized directly on the anode. The kinetic studies show that the dissolution of molybdenite concentrate is represented by shrinking core model with diffusion through a porous product layer of element sulfur, and the apparent activation energy for the dissolution reaction is 8.56 kJ/mol.     

Electric-oxidation extraction of molybdenite concentrate in alkaline NaCl electrolyte

The dissolution of molybdenite concentrate in NaCl electrolyte was investigated. The results show that the dissolution rate increases with the increase in liquid-to-solid ratio, stirring speed, NaCl concentration and temperature. When the liquid-to-solid ratio is 30:1, stirring speed is 400 r/min, concentration of NaCl is 4 mol/L at pH=9 and room temperature, the leaching efficiency of molybdenite concentrate will reach 99.5% in 240 min. Molybdenite concentrate cannot be electro-oxidized directly on the anode. The kinetic studies show that the dissolution of molybdenite concentrate is represented by shrinking core model with diffusion through a porous product layer of element sulfur, and the apparent activation energy for the dissolution reaction is 8.56 kJ/mol.     

Kinetics of nickel leaching from roasting-dissolving residue of spent catalyst with sulfuric acid

Sulfuric acid leaching process was applied to extract nickel from roasting-dissolving residue of a spent catalyst, the effect of different parameters on nickel extraction was investigated by leaching experiments, and the leaching kinetics of nickel was analyzed. The experimental results indicate that the effects of particle size and sulfuric acid concentration on the nickel extraction are remarkable; the effect of reaction temperature is mild; while the effect of stirring speed in the range of 400–1 200 r/min is negligible. Decreasing particle size or increasing sulfuric acid concentration and reaction temperature, the nickel extraction efficiency is improved. 93.5% of nickel in residue is extracted under suitable leaching conditions, including particle size (0.074–0.100) mm, sulfuric acid concentration 30% (mass fraction), temperature 80 °C, reaction time 180 min, mass ratio of liquid to solid 10 and stirring speed 800 r/min. The leaching kinetics analyses shows that the reaction rate of leaching process is controlled by diffusion through the product layer, and the calculated activation energy of 15.8 kJ/mol is characteristic for a diffusion controlled process. Foundation item: Project (50574101) supported by the National Natural Science Foundation of China; Project (2003UDBEA00C020) supported by the Collaborative Project of School and Province of Yunnan Province, China     

氯化浸出铅阳极泥回收金的研究

试验对湿法处理铅阳极泥回收金进行了研疣．实验结果表明：铜的浸出是酸浸除杂的主要影响因素,在温度为65℃,盐酸浓度为2．9moL／L,氯化钠浓度为1．3mol／L,硫酸浓度为0．3mol／L时,铜的浸出率达到92．03％,杂质能有效去除．在氯化浸金的过程中,温度与氯酸钠的用量对金的浸出率影响最大．随着两者的增加,金的浸出率将明显地增加．当温度大于80℃,氯酸钠用量大于料重的12．5％,硫酸浓度为1．5mol／L时,金的浸出率大于98％．     

Techniques of copper recovery from Mexican copper oxide ore

Mexican copper ore is a mixed ore containing mainly copper oxide and some copper sulfide that responds well to flotation. The joint techniques of flotation and leaching were studied. The results indicate that an ore containing 19.01% copper could be obtained at a recovery ratio of 35.02% by using sodium sulfide and butyl xanthate flotation. Over 83.33% of the copper oxide can be recovered from the tailings by leaching in suitable conditions, such as 1 h stirring at a temperature around 25 ℃ with a mixing speed of 500 r/min, an H2SO4 concentration of 1.0 mol/L and a mass ratio of the ore-slurry-liquid to solid (mL/mS) of 3. The overall yield of refined ore after flotation and leaching is over 89.18% of the copper, which is much better than sole flotation or leaching. A copper product containing more than 99.9% copper was obtained by using the process: flotation-agitation leaching-solvent extraction-electro-winning.     

电沉积Ni La P合金上析氢的研究

在１ ｍｏｌ·Ｌ－ １ 　ＫＯＨ 水溶液中测量了Ｎｉ 电极和Ｎｉ Ｌａ Ｐ合金电极上析氢反应的极化曲线．实验结果表明,与镍电极相比,Ｎｉ Ｌａ Ｐ合金电极上析氢的速度比Ｎｉ 电极约大１０ 倍,析氢电势正移２００ ～３００ ｍＶ（ 在ｉｃ ＝ １００～１５０ｍＡ·ｃｍ － ２） ．显示出Ｎｉ Ｌａ Ｐ合金具有较高的析氢催化活性,有利于降低槽电压,减少能耗．     

Zinc bioleaching from an iron concentrate using Acidithiobacillusferrooxidans strain from Hercules Mine of Coahuila, Mexico

The iron concentrate from Hercules Mine of Coahuila,Mexico,which mainly contained pyrite and pyrrhotite,was treated by the bioleaching process using native strain Acidithiobacillus ferrooxidans(A.ferrooxidans)to determime the ability of these bacteria on the leaching of zinc.The native bacteria were isolated from the iron concentrate of the mine.The bioleaching experiments were carried out in shake flasks to analyze the effects of pH values,pulp density,and the ferrous sulfate concentration on the bioleaching process.The results obtained by microbial kinetic analyses for the evaluation of some aspects of zinc leaching show that the native bacteria A.ferrooxidans,which is enriched with a 9K Silverman medium under the optimum conditions ofpH 2.0,20 g/L pulp density,and 40 g/L FeSO4,increases the zinc extraction considerably observed by monitoring duringl 5 d,i.e.,the zinc concentration has a decrease of about 95％ in the iron concentrate.     

Leaching kinetics of low grade zinc oxide ore in NH3-NH4Cl-H2O system

The leaching kinetics of low grade zinc oxide ore in NH3-NH4Cl-H2O system was studied. The effects of ore particle size,reaction temperature and the sum concentration of ammonium ion and ammonia on the leaching efficiency of zinc were examined.The leaching kinetics of low-grade zinc oxide ore in NH3-NH4Cl-H2O system follows the kinetic law of shrinking-core model. The results show that diffusion through the inert particle pores is the leaching kinetics rate controlling step. The calculated apparent activation energy of the process is about 7.057kJ/mol. The leaching efficiency of zinc is 92.1% under the conditions of ore particle size of 69μm, holding at 80℃ for 60min, sum ammonia concentration of 7.5mol/L, the molar ratio of ammonium to ammonia being 2:1, and the ratio (g/mL) of solid to liquid being 1:10.     

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching ternperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h: sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50℃; and,agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.     

Dissolution kinetics of malachite in ammonia/ammonium sulphate solution

The dissolution kinetics of malachite was investigated in ammonia/ammonium sulphate solution. The effects of ammonia and ammonium sulphate concentration, pH, leaching time, reaction temperature, and particle size were determined. The results show that the optimum leaching conditions for malachite ore with a copper extraction more than 96.8% are ammonia/ammonium concentration 3.0 mol/L NH₄OH + 1.5 mol/L (NH₄)₂SO₄, liquid-to-solid ratio 25:1 mL/g, leaching time 120 min, stirring speed 500 r/min, reaction temperature 25 °C and particle size finer than 0.045 mm. The dissolution process of malachite with an activation energy of 26.75 kJ/mol is controlled by the interface transfer and diffusion across the product layer. A semi-empirical rate equation is obtained to describe the leaching process and the reaction orders with respect to concentration of ammonia and ammonium sulphate are 2.983 0 and 0.941 1, respectively.     

Extraction of nickel from molybdenum leaching residue of metalliferous black shale by segregation roasting and acid leaching

The recovery of nickel from molybdenum leach residue by the process of segregation roasting-sulfuric acid leaching-solvent extraction was investigated. The residue was characterized by microscopic investigations, using X-ray fluorescence spectrometry (XRF) and X-ray diffractometry (XRD) techniques and the residue after segregation roasting was characterized by chemical phase analysis method. A series of experiments were conducted to examine the mass ratio of activated carbon (AC) to the residue, segregation roasting time and temperature, sulfuric acid concentration, liquid-to-solid ratio, leaching time, leaching temperature, addition amount of 30% H₂O₂, stirring speed (a constant) on the leaching efficiency of nickel. A maximum nickel leaching efficiency of 90.5% is achieved with the mass ratio of AC to the residue of 1:2.5, segregation roasting time of 2 h, segregation roasting temperature of 850 °C, sulfuric acid concentration of 4.5 mol/L, liquid-to-solid ratio of 6:1, leaching time of 5 h, leaching temperature of 80 °C, addition of 30% H₂O₂ of 0.6 mL for 1 g dry residue. Under these optimized conditions, the average leaching efficiency of nickel is 89.3%. The nickel extraction efficiency in the examined conditions is about 99.6%, and the nickel stripping efficiency in the examined conditions is about 99.2%.     

Zinc leaching from electric arc furnace dust in alkaline medium

Physical and chemical properties of electric arc furnace (EAF) dust from Tianjin seamless Pipe Company were measured and analyzed. The zinc leaching tests in alkaline medium were carried out under variation of leaching agent concentration, leaching temperature, leaching cumulative time and solid-to-liquid ratio. The thermodynamics and kinetics of the zinc leaching process were also analyzed. The results show that the EAF dust contains 10% (mass fraction) zinc and the median particle size is 0.69 μm. The zinc recovery of 73.4% is obtained under the condition of 90 °C, 6 mol/L NaOH, and 60 min leaching time. With the increase of concentration of NaOH and the cumulative time, zinc leaching will be significantly increased. The kinetics study demonstrates that the leaching reaction is chemically controlled and the reaction activation energy is 15.73 kJ/mol.     

常压酸浸高冰镍的动力学研究

本文是关于常压酸浸高冰镍的动力学研究.实验考察了粒度、温度、硫酸浓度和氯离子浓度对镍浸出速率的影响.结果表明,在硫酸体系中选择性地浸出镍不仅在热力学上是可能的,而且反应速率快,完全能达到一次分离镍铜的目的.本文采用特殊的处理方法,将总镍浸出率R转换为Ni_3S_2中镍浸出率R~(?),从而导出实验条件下的动力学方程式.而浸出过程受通过产物层的固膜扩散控制.     

硫化铜精矿焙烧的非等温动力学研究

采用DTA，TG结合X-射线衍射分析，对硫化铜精矿的焙烧的动力学进行了研究．发现硫化铜精矿焙烧可以分成两个阶段．在第一个阶段(673～857K)，铜的硫化物被氧化成硫酸盐；在第三个阶段(857～1073K)，生成的硫酸盐将分解为氧化物．其两个阶段都受界面化学反应所控制，表观活化能分别为159．5kJ／mol和242．2H／mol．