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%.     

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.     

采用酸浸法从废旧锂离子电池中回收金属钴

分别采用盐酸-30%过氧化氢、硫酸-30%过氧化氢和硝酸-30%过氧化氢为浸取液,回收锂离子电池正极材料中的金属钴,研究了浓度、温度、30%过氧化氢的量、反应时间、固-液比等对钴浸取率的影响.经响应面分析实验结果表明,盐酸-30%过氧化氢最适宜做浸取液,最佳工艺条件为:盐酸浓度为4.0 mol/L,盐酸-30%过氧化氢...     

Electrogenerative leaching of nickel sulfide concentrate with ferric chloride

In order to utilize the chemical energy in hydrometallurgical process of sulfide minerals reasonably and to simplify the purifying process, the electrogenerative process was applied and a dual cell system was introduced to investigate FeCl3 leaching of nickel sulfide concentrate. Some factors influencing the electrogenerative leaching, such as electrode structure, temperature and solution concentration were studied. The results show that a certain quantity of electrical energy accompanied with the leached products can be acquired in the electrogenerative leaching process.The output current and power increase with the addition of acetylene black to the electrode. Varying the components of electrode just affects the polarization degree of anode. Increasing FeCl3 concentration results in a sharp increase in the output of the leaching cell when c(FeCl3) is less than 0.1mol/L. The optimum value of NaCl concentration for electrogenerative leaching nickel sulfide concentrate with FeCl3 is 3.0 mol/L. Temperature influences electrogeneratire leaching by affecting anodic and cathodic polarization simultaneously. The apparent activation energy is determined to be 34.63 kJ/mol in the range of 298 K to 322 K. The leaching rate of Ni^2 is 29.3% after FeCl3 electrogenerative leaching of nickel sulfide concentrate for 620 min with a filter bag electrode.     

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 7.057 kJ/mol. The leaching efficiency of zinc is 92.1% under the conditions of ore particle size of 69μm, holding at 80℃ for 60 min, sum ammonia concentration of 7.5 mol/L, the molar ratio of ammonium to ammonia being 2：1, and the ratio （g/mL） of solid to liquid being 1：10.     

New technique of comprehensive utilization of spent Al2O3 -based catalyst

A new technology was developed to recover multiple valuable elements from the spent Al2O3-based catalyst by X-ray phase analysis and exploratory experiments. The experimental results show that in the condition of roasting temperature of 750 ℃ and roasting time of 30 min, molar ratio of Na2O to Al2O3 of 1.2, the leaching rates of alumina, vanadium and molybdenum in the spent catalyst are 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by precipitators A and B, and the precipitation rates of vanadium and molybdenum are 94. 8% and 92. 6%. Al（OH）3 was prepared from sodium aluminate solution in the carbonation decomposition process, and the purity of Al2O3 is 99. 9% after calcination, the recovery of alumina reaches 90. 6% in the whole process; the Ni-Co concentrate was leached by sulfuric acid, a nickel recovery of 98. 2% and cobalt recovery over 98.5% can be obtained under the experimental condition of 30% H2SO4, 80 ℃, reaction time 4 h, mass ratio of liquid to solid 8, stirring rate 800 r/min.     

Extraction of vanadium from stone coal by modified salt-roasting process

In order to reduce the pollution of Cl₂ and HCl released during extracting vanadium from stone coal by sodium chloride roasting, a modified salt-roasting process was proposed by adding calcined lime in roasting process followed by H₂SO₄ leaching. The effects of parameters including roasting temperature, roasting time, addition mass ratio of NaCl, calcined lime upon leaching rate of vanadium and curing rate of chlorine were investigated, and the effects of leaching time and leaching temperature on leaching rate of vanadium were also studied. The results show that the vanadium leaching rate and the curing rate of chlorine are 67.3% and 51.5% (mass fraction), respectively, at roasting temperature of 750 °C, roasting time of 4 h, 15% sodium chloride and 8% (mass fraction) calcined lime, leaching temperature of room temperature, and leaching time of 4 h.     

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.     

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.     

Application of multi-stage dynamic magnetizing roasting technology on the utilization of cryptocrystalline oolitic hematite:A reviewOA

A large number of studies have shown that oolitic hematite is an iron ore that is extremely difficult to utilize because of its fine disseminated particle size, high harmful impurity content and oolitic structure.To recover iron from oolitic hematite, we developed a novel multistage dynamic magnetizing roasting technology. Compared with traditional magnetizing roasting technologies, this novel technology has the following advantages: firstly, the oolitic hematite is dynamically reduced in a mult...     

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.     

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     

氨基磺酸法剥离废印刷线路板表面镀金层的研究

含有贵金属金的废弃印刷线路板有较高回收价值。采用氨基磺酸腐蚀法,浸出镀金层下的铜和镍,回收得到金箔。探索氨基磺酸法回收金的最佳工艺条件,研究结果表明:常温条件下,取内存条与氨基磺酸溶液的固液比(g/mL) 1:5,双氧水体积分数15%,氨基磺酸浓度70 g/L,废旧内存条浸出时间120 min (不同类型板因镀层差异剥落时间稍有差异),此工艺条件下金的回收率可以达到96%以上。氨基磺酸腐蚀法对金的回收选择性好,剥离的金箔可直接熔炼,脱金后的线路板可用于回收其他金属和非金属;溶液回收铜镍后,可循环使用,实现回收过程中的闭路循环,减轻环境污染。     

Nickel extraction from low grade laterite by agitation leaching at atmospheric pressure

This research work deals with the extraction of nickel from a low grade nickel laterite ore, taken from a deposit located in southwestern of Iran, through agitation leaching at atmospheric pressure. The assaying and mineralogical studies carried out on the nickel laterite sample, showed the 0.88% Ni, and principally consisted of oxide and silicate crystalline phases i.e. dolomite, quartz, magnetite, and goethite. Among numerous factors affecting such process, four major parameters i.e. temperature, agitator speed (r/min), leaching agents and their concentration were considered in a two-level full factorial experimental design. The agitation leach tests showed that the ore could be leached at atmospheric pressure with sulfuric acid while citric acid was almost unpromising. Analysis of variance (ANOVA) using DX7 software was employed to identify effective parameters. Sulfuric acid concentration and temperature were the most effective parameters on Ni extraction. Furthermore, the factorial models for experiment responses were developed. The results showed 83% Ni extraction after 4 h leaching, under optimized conditions i.e. temperature at 95 °C, acid concentration at 5 N and agitator speed at 1000 r/min. This study revealed that factorial experimental design can be implemented to identify effective parameters on the agitation leaching process of nickel laterite.     

Mechanism for suspension magnetization roasting of iron ore using straw-type biomass reductant

As an alternative reductant for fossil fuel in the future, straw-type biomass contributes to emission reduction and green utilization in the suspension roasting process. In this study, the influences of the roasting time, roasting temperature and dose of straw-type biomass after suspension magnetization roasting (SMR) and separation were investigated. The optimal conditions were determined to be a roasting time of 7.5 min with a straw-type biomass dose of 20 wt% and a roasting temperature of 800 ℃, in which an iron grade of 71.07% and recovery of 94.17% were obtained for the iron concentrate. The maximum saturation magnetization under optimal conditions was 35.05 A·m²·g⁻¹, and the gaseous regulation of the biomass revealed that cumulative reducing gas volume was 293.93 mL at the optimal roasting time of 450 s. The transformation of hematite to magnetite was detected by X-ray diffraction (XRD). During microstructure evolution, the outer layer consisting of fissures and tiny holes continuously deepened toward the core.     

除铜渣氯气全浸探索

利用氯气对除铜渣进行了全浸,浸出条件为：除铜渣粒度约为－200目,电位315－460mV,温度102－110℃,搅拌强度330－350r/min,液固比2－2．5：1,氯气压力0．1－0．4MP,多级浸出结果表明,浸出液含铜90．23g/l,含镍55．10g/l,浸出渣含铜3．2％,含镍1．8％,铜浸出率为94．3％,镍浸出率为95．20％,本研究项目已投入工业运行。     

高硅氧化锌矿浸出工艺的研究

针对高硅氧化锌矿的处理进行了研究,提出了二段循环酸浸的湿法冶金工艺.通过条件优化实验确定一、二段浸出条件分别为:始酸浓度75g/L、温度70℃、液固比6∶1、浸出时间1.5 h;始酸浓度18g/L、温度60℃、液固比3∶1、浸出时间0.5 h.本试验采用一段六轮循环浸出及二段三轮循环浸出,运行稳定,锌的总浸出率达到93.4%.     

复配还原剂体系用于还原蓝RSN染色工艺研究

以保险粉-葡萄糖复配体系作为还原蓝RSN染纯棉织物的还原剂,通过单因子试验考察了葡萄糖与保险粉浓度比例、氢氧化钠浓度、染色温度等对上染率、固色率、K/S等的影响.通过正交实验得到染色优化工艺为:染料2%(owf),保险粉13.5 g/L,葡萄糖1.5 g/L,氢氧化钠16 g/L,染色还原温度50℃.通过优化工艺与传统还原染色工艺对比实验得出,复配体系优化工艺染色效果高于传统染色工艺.     

Process optimization of reaction of acid leaching residue of asbestos tailing and sodium hydroxide aqueous solution

Silica is the major component of the acid leaching residue of asbestos tailing. The waterglass solution can be prepared by the reaction of the residue with sodium hydroxide aqueous solution. Compared to the high temperature reaction method, this process is environmental friendly and low cost. In this paper, the reaction process of the residue and the sodium hydroxide aqueous solution is optimized. The optimum reaction process parameters are as follows: the usage of sodium hydroxide is 26.4 g/100 g acid leaching residue, the reaction temperature is 90℃, the reaction time is 1 h, and the ratio of the liquid/solid is 2.0. The significance sequence of the process parameters to the alkali leaching reaction effect is the usage of sodium hydroxide > the ratio of the liquid/solid > the reaction time > the reaction temperature. The significance sequence to the leaching ratio of SiO2 is the ratio of the liquid/solid > the usage of sodium hydroxide > the reaction time > the reaction temperature. The significance sequence to the modulus of the sodium silicate is the ratio of the liquid/solid > the usage of sodium hydroxide > the reaction time > the reaction temperature. Under the optimum conditions, the leaching ratio of the SiO2 is 77.5%, and the modulus of the sodium silicate is 3.15. The XRD analysis result indicates that the major components of the alkali leaching residue are serpentine, talc, quartz and some albite.     

磁化矿石颗粒模型及磁选过程分析

基于磁选过程中颗粒尺寸、磁场强度和磁选精矿品位三者之间的关系,建立磁化矿石颗粒模型,对其进行理论分析与计算,确定最佳磁场强度,并进行磁化矿石的磁选研究。结果表明：在配煤量4%（质量分数）,焙烧温度850℃,焙烧时间60 min,磨矿细度-0.074 mm占60%（质量分数）,磁场强度为40 mT的条件下,得到铁品位57.7%（质量分数）,铁回收率90.3%（质量分数）的铁精矿,较好地实现了铁精矿的富集和回收。