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 cobalt from laterite ores by citric acid in presence of ammonium bifluoride

Citric acid was used to selectively extract cobalt from limonite-type laterite ores in the presence of ammonium bifluoride. The results show that ammonium bifluoride enhances the leaching of cobalt by citric acid, and 84.5% cobalt is extracted from a laterite ore containing 0.13% Co when leached at ambient temperature for 2 h with 30 g/L citric acid and 10 g/L ammonium bifluoride. Pyrolusite is reduced by citric acid during leaching, cobalt intergrown with which is liberated and subsequently chelated by the citric acid. The extraction of cobalt is enhanced in the presence of ammonium bifluoride because the matrix of silicate minerals is destroyed by ammonium bifluoride and the adsorbed cobalt is subsequently liberated.     

镍基合金电解加要渣泥中镍和钴的回收

根据镍基合金电解加工渣泥的组成和热力学分析,用水洗-浸出-水争沉淀工艺处理这种渣泥,试验结果表明：氯化钠、硝酸钠和部分铬盐等可溶性化合物可以从渣泥中洗出;继而的硫酸浸出过程中,可将洗后渣泥中的镍、钴和铬浸出,而钨和钼留在浸出渣中,在浸出过程中,镍和钴的浸出率达98%,浸出液中的铬可通过水解沉淀法以氢氧化铬的形态除去,铬的脱除率达94.44%。     

Recovery of Mn^2＋, Co^2＋ and Ni^2＋ from manganese nodules by redox leaching and solvent extraction

The recovery of Mn, Co and Ni from deep-sea manganese nodules was conducted by acid oxidative leaching and solvent extraction. The results indicate that pyrrhotite used during leaching can effectively facilitate the leaching out of manganese, cobalt and nickel. The leaching behaviors of Mn, Ni and Co were determined and the influences of temperature, leaching time and sulphuric acid concentration on leaching rate were also investigated. Co and Ni are precipitated from the leaching liquor by adding sodium sulfide into solution with agitation for 2 h at 50 ℃, and the manganese sulphate is obtained by concentrating the resulting solution. By re-dissolving the precipitates of cobalt and nickel, the separation of cobalt and nickel is performed using di（2-ethylhexyl） phosphoric acid （D2EHPA） for impurities elimination with 8 stages at organic-to-aqueous（O/A） volume ratio of 3：5, and 2- ethylhexyl phosphonic acid mono-2-ethylhexyl ester （known as PC88A or P507） for cobalt extraction with 3 stages counter-current operations at O/A volume ratio of 2：3 followed by their scrubbings and strippings, respectively. The final maximum recovery rates for manganese, cobalt and nickel are 85%, 75% and 78%, respectively.     

Process for cobalt separation and recovery in the presence of nickel from sulphate solutions by Cyanex 272

The present paper deals with the extraction of cobalt from a solution containing cobalt and nickel in a sulphate medium similar to the leach liquor obtained by the dilute sulphuric acid pressure leaching of the Pacific Ocean nodules matte followed by copper extraction. The commercial extractant Cyanex 272 (bis (2, 4, 4-trimethylpentyl) phosphinic acid) is used for this purpose. The leach liquor used for the present study contains Co =1.78 g/L and Ni=16.78 g/L. Before cobalt extraction, impurities, such as copper and iron, are removed from the leach liquor by the precipitation method. Increasing the concentration of Cyanex 272 increased the extraction percentage of cobalt due to the increase of equilibrium pH. Cobalt extraction efficiency of >99.9 % is achieved with 0.20 M Cyanex 272 in two counter-current stages at an aqueous: organic (A:O) phase ratio of 1.5∶1. Complete stripping of cobalt from the loaded organic containing 2.73 g/L Co was carried out at pH 1.4 by a synthetic cobalt spent electrolyte in two stages at an A:O ratio of 1∶2. The enrichment of cobalt during extraction and stripping operations was about 3.5 times. A complete process flowsheet for the separation and recovery of cobalt is presented.     

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...     

Recovery of Co（Ⅱ） and Ni（ Ⅱ） from hydrochloric acid solution of alloy scrap

A hydrometallurgical process was developed for recovery of nickel and cobalt from the hydrochloric acid leaching solution of alloy scraps. The process consists of five maj or unit operations： 1） leaching with 6 mol/L hydrochloric acid under the L/S ratio of 10：1 at 95 ℃ for 3 h; 2） copper replacement by iron scraps under pH value of 2.0 at 80 ℃, and stirring for 1 h, 3） removal of iron and chromium by chemical precipitation： iron removal under pH value of 2.0 at 90 ℃ by dropwise addition of sodium chlorate and 18% sodium carbonate solution, then chromium removal under pH value of 4.0 at 70 ℃ by addition of nickel carbonate solution, stirred by air flow for 2 h; 4） selective separation of cobalt from nickel by extraction using 30% trialkyl amine＋50% kerosene （volume fraction） and tri-n-butylphosphate （TBP） as a phase modifier with the O/A ratio of 2：1, and stripping of cobalt with 0.01 mol/L HCl; 5） crystallization of nickel chloride and electrodeposition of cobalt. It is found that the nickel recovery of 95% and the cobalt recovery of approximately 60% with purity over 99.9% are obtained by this process.     

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。     

Recovery of valuable metals from anode material of hydrogen-nickel battery

Simultaneous recovery of rare earth, nickel and cobalt resources from the anode material of hydrogen-nickel battery was performed through a hydrometallurgical process. Most of rare earth elements are separated from nickel and cobalt in the form of sulfates when the anode material is firstly leached with sulfuric acid. Then, the precipitated rare earth sulfates are dissolved with sodium hydroxide to form rare earth hydroxides. The rare earth element, zinc and manganese ions in the lixivium are also separated from nickel and cobalt by using PC-88A extractant system, and the organic phase loaded rare earth is stripped with hydrochloric acid. By neutralizing the stripping solution with rare earth hydroxide, the rare earth chloride is obtained. Under the suitable leaching conditions of sulfuric acid 3 mol/L, leaching time 4 h and temperature 95 ℃, 94.5% of rare earth in the anode material is transformed into the sulfate precipitates, and the leaching ratios of nickel and cobalt can approach 99.5%. When the pH value of the extractive system is controlled in the range of 3.0-3.5, the rare earth elements in the lixivium can be extracted completely into the organic phase, and the stripping recovery of the rare earth can reach 98% in the extraction stage. The total recoveries of rare earth, nickel and cobalt are 98.9%, 98.4% and 98.5%, respectively.     

加压浸出脱除钴白合金中金属铜和钴的研究

阐述了采用加压浸出方法对金属铜和钴进行有效脱除。实验结果表明：加压浸出过程的较佳反应条件为反应温度150℃、硫酸浓度30 g/L、反应压力1.5 MPa、反应时间6 h、液固比6：1。在此反应条件下,金属Cu、Co的浸出效率分别达到了90％和88％。     

Use of solutions of organic acids to produce low friction anodised surfaces

The properties of anodic films produced by anodising aluminium in aqueous solutions of organic acids at room temperature at commercially interesting current densities were investigated with particular emphasis on the wear index, micro hardness, surface roughness and pore structure. Voltage time curves obtained using 10% w./v. solutions of the pure acids showed a rapid rise in voltage at commercially realistic current densities of 1–2 A dm^?2 but this problem could be overcome by the addition of 1–2% v./v. sulphuric acid. The films obtained under these conditions using sulphosuccinic acid (SSA), sulphosalicylic (SCA) or sulphophthalic acids (SPHA) had properties closely resembling those obtained under hard anodising conditions in 10% sulphuric acid at 0–4°C. Films were not formed using para-toluenesulphonic acid (PTSA) and the use of maleic acid (MA) resulted in the formation of substantial quantities of fumaric acid. The oxide films exhibited a structure made up of densely packed columns suffused by narrow pores. The films fall into two categories. On the one hand films produced using sulphur containing organic acids have very narrow pores of less than 10 nm diameter resembling those produced by the conventional low temperature sulphuric acid process, and on the other hand films produced using maleic or citric acids exhibit much larger pores with diameters of the order of 20 nm. Attempts to incorporate PTFE particles from aqeous dispersions of this polymer into the films were unsuccessful due to their large size (~200 nm). Duplex films with a highly porous outer layer (~100 nm diameter) on top of a dense hard layer were produced and evidence is presented which shows that some of the smaller PTFE particles can enter the mouths of these pores.     

从镍钼矿中提取镍钼的工艺

针对现行镍钼矿处理工艺存在的钼镍需要分别提取的缺陷,提出镍钼矿加钙氧化焙烧-低温硫酸化焙烧-水浸提取镍钼的新工艺。以贵州遵义镍钼矿为原料,对CaO加入量、氧化焙烧温度、氧化焙烧时间、硫酸加入量、硫酸化焙烧温度、硫酸化焙烧时间以及焙砂水浸工艺参数对镍钼浸出率的影响进行研究。结果表明:在最佳工艺条件下,钼的浸出率为97.33%,镍的浸出率为93.16%,且最佳工艺参数为100 g镍钼矿加入35 g CaO,700℃氧化焙烧2 h,得到的焙砂加入70 mL浓硫酸,再经250℃硫酸化焙烧2 h;硫酸化焙烧得到的焙砂按液固比2:1加水搅拌,经98℃浸出2 h。加入CaO不仅能有效减少镍钼矿氧化焙烧烟气对环境造成的污染,而且能显著提高镍的浸出率。     

通过选择性浸出、结晶和沉积回收废重整催化剂中的镍和铝

研究和优化不同工艺条件下硫酸(3.0~5.5 mol/L)浸出回收Ni和Al的工艺。浸出实验表明,在H₂SO₄浓度5.5 mol/L、反应时间4 h、固液比0.2 g/mL、温度358 K、粒径<100μm、搅拌速度200~250 r/min、催化剂用量5.0 g的条件下,可提取98.5%的NiO和40.7%的Al₂O₃。浸出液中的Al用1.4 mol/L KOH选择性结晶分离,Ni用0.3 mol/L H₂C₂O₄选择性沉淀分离。此法可回收约97.9%的NiO,纯度达98.3%;约25%的Al₂O₃以明矾(纯度为99%)形式回收,14.7%的Al₂O₃以Al-K-C₂O₄-SO₄盐的形式回收。研究结果表明,硫酸是一种合适的选择性浸出溶剂,而且可以从硫酸盐溶液中选择性结晶出明矾。TG-DTA/DTG和XRD表征研究证明,本工艺可以有效地提取和回收镍和铝。     

Study of decomposing carbonyl slag

1 Introduction Carbonyl slag is a kind of solid from the material with nickel by carbonylation. It usually contains Cu, Ni, Co, Fe, S, As, Sb, Au, Ag and platinum metals, and holds great recovery value. At present, with the development of the technology o…     

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.     

Pressure leaching of metals from waste printed circuit boards using sulfuric acid

Printed circuit boards (PCBs) are essential components of electronic equipments which contain various metallic values. This paper reports a hydrometallurgical recycling process for waste PCBs, which consists of the novel pretreatment consisting of organic swelling of PCBs followed by sulfuric acid leaching of metals from waste PCBs. To recycle the waste PCBs, experiments were carried out for the recovery of copper from the crushed and organic swelled materials of waste PCBs using sulfuric acid leaching in presence of hydrogen peroxide under atmospheric and pressure condition. The leaching of PCBs at 90°C, pulp density 100 g/L under atmospheric condition, using 6M sulfuric acid resulted in the dissolution of a minor amount of copper due to the presence of plastic coating on the surface of metallic layers. On the other hand, when the liberated metal sheets from organic swelled PCBs were treated with dilute sulfuric acid of concentration 2M along with hydrogen peroxide in an autoclave under oxygen atmosphere, the percentage recovery of copper was found to increase from 59.63% to 97.01% with an increase in hydrogen peroxide concentration from 5 to 15% (v/v) keeping constant pulp density 30 g/L.     

Photogenerated-hole scavenger for enhancing photocatalytic chalcopyrite bioleaching

The effects of photogenerated-hole scavengers (ascorbic acid, oxalic acid, humic acid and citric acid) on chalcopyrite bioleaching in the presence of visible light were studied using Acidithiobacillus ferrooxidans (A. ferrooxidans). Four sets of bioleaching experiments were performed: (1) visible light + 0 g/L scavenger, (2) visible light + 0.1 g/L of different scavenger (ascorbic acid, oxalic acid, humic acid and citric acid), (3) dark + 0.1 g/L of different scavenger (ascorbic acid, oxalic acid, humic acid and citric acid), and (4) dark + 0 g/L scavenger (control group). The results showed that ascorbic acid and oxalic acid could act as photogenerated-hole scavengers and significantly enhance chalcopyrite bioleaching under visible light. The dissolved copper in the light group without scavenger was only 18.7% higher than that of the control group. The copper extraction rates of the light groups with oxalic acid and ascorbic acid were respectively 30.1% and 32.5% higher than those of the control group. Scanning electron microscopy (SEM), X-ray diffraction (XRD) and Fourier transform infrared spectroscopy (FT-IR) analyses indicated that ascorbic acid and oxalic acid as photogenerated-hole scavenger could capture photo-generated holes and inhibit jarosite formation on the chalcopyrite surface, thereby enhancing bioleaching of chalcopyrite under visible light.     

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.     

Efficient recycling of WC-Co hardmetal sludge by oxidation followed by alkali and sulfuric acid treatments

We present a process to recycle strategic metals, viz. tungsten and cobalt, from a WC-Co hardmetal sludge (WCHS) via oxidation followed by a two-step hydrometallurgical treatment with alkali and acid solutions. The oxidation of WCHS was investigated in the temperature range of 500 to 1000 °C and optimized at 600 °C to transform the maximum WC into an alkali-soluble WO₃. The conditions for the selective dissolution of WO3 in stage-I were optimized as follows: 4.0 M NaOH, pulp density of 175 g/L, and temperature of 100 °C for 1 h, yielding maximum efficacy. Subsequently, in the second step, the optimal conditions for cobalt leaching from the alkali-treated residue were established as follows: 2.0 M H₂SO₄, 25 g/L pulp density, and 75 °C temperature for 30 min. Downstream processing of the obtained metal ions in solutions was also easier, as the only impurity of dicobaltite ions with the Na₂WO₄ solution was precipitated as Co(OH)₃ under atmospheric O₂; meanwhile, the CoSO₄ solution obtained through the second step of processing can be treated via electrolysis to recover the metallic cobalt. The present process is simpler in operation, and the efficient use of eco-friendly lixiviants eliminates the previously reported disadvantage.     

化学镀镍工艺条件的计算机优化

为研究化学镀镍最佳工艺条件,以LM-4铝为基体,讨论了均匀设计及计算机优化在化学镀镍中的应用.试验按5因子、10水平、10次试验构成均匀试验表,用综合指数对试样进行量化测评,试验结果用SPSS软件进行多项式回归分析,得到了综合指数与诸因素的多项式回归方程,用MATLAB语言对该方程进行单纯性寻优处理,得到的最佳镀液组成为:硫酸镍30g/L、次亚磷酸钠30g/L、糖精0.5g/L、醋酸钠8g/L、柠檬酸钠19g/L、丁二酸7.5g/L、DL-苹果酸9.5g/L,pH值为5.2,综合指数为98.82.经过试验验证的最后结果与计算机优化的结果基本相符.