RECOVERING RE FROM LEACHING LIQUOR OF RARE EARTH ORE BY EXTRACTION

ＲＥＣＯＶＥＲＩＮＧＲＥＦＲＯＭＬＥＡＣＨＩＮＧＬＩＱＵＯＲＯＦＲＡＲＥＥＡＲＴＨＯＲＥＢＹＥＸＴＲＡＣＴＩＯＮＣｈｉＲｕ＇ａｎ；ＸｕＪｉｎｇｍｉｎｇ；ＨｅＰｅｉｊｉｏｎｇ；ＺｈｕＹｏｎｇｊｕｎ（ＩｎｓｔｉｔｕｔｅｏｆＮｕｃｌｅａｒＥｎｅｒｇｙＴｅｃ...     

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.     

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.     

Removing aluminum from a low-concentration lixivium of weathered crust elution-deposited rare earth ore with neutralizing hydrolysis

Aluminum,the main impurity in the lixivium of weathered crust elution deposited rare earth ore,not only results in an increasing consumption of precipitant in the rare earth precipitation process,but also lowers the purity of final rare earth product.Aluminum in the weathered crust elution-deposited rare earth ore lixivium should be removed.Neutralizing hydrolysis method was employed to remove aluminum from the lixivium.Hexamethylenetetramine was found to be the optimum pH regulator for the removal of aluminum in the low concentration.When used to adjust the pH value of the lixivium to 5.0,aluminum in the lixivium can be effectively removed in the form of aluminum hydroxide precipitation with removal rate of 97.60%.It shows that hexamethylenetetramine has a good effect on the removing of aluminum ions from the lowconcentration lixivium.Moreover,hexamethylenetetramine in removing aluminum from lixivium has little adverse effect on the RE precipitation process.     

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%。     

铜阳极泥加压酸浸预处理脱铜富集贵金属

铜阳极泥是铜电解精炼中的一种副产品,是回收贵金属的重要原料。以空气代替氧气为加恪气体,系统悁究了铜阳极泥加恪酸浸预处理工艺。悁究的影响因素包括硫酸浓度、空气分恪、浸出温度、浸出时间和液固比等。结果表明,在最优工艺条件下,铜的浸出率高达98%;碲、硒、银的浸出率分别为49%、13%、1%。试验数据表明：铜阳极泥经过加恪酸浸预处理,几乎所有的铜和部分的碲,能有效的从贵金属中分离出来,使贵金属得以富集。机理悁究表明：充分利用阳极泥中水溶性铜离子的自催化氧化作用,能有效提高铜的溶解速度。     

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.     

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...     

Dissolution of lateritic nickel ore using ascorbic acid as synergistic reagent in sulphuric acid solution

The dissolution of nickel and cobalt from Caldag lateritic nickel ore using the combination of sulphuric and ascorbic acids was investigated. The use of other organic acids, namely citric, maleic and stearic acids, as synergistic reagents was studied for comparison. The results revealed that the use of ascorbic and citric acids markedly improved the dissolution of cobalt compared to the other two organic acids that only showed slight synergistic effect on the leaching rate. In terms of nickel dissolution, ascorbic acid is the most effective synergist, followed by citric, maleic and stearic acids in descending order. Under the most optimized conditions found in this study, i.e., using 1 mol/L of sulphuric acid with the presence of 4 g/L of ascorbic acid at 80 °C and solid-to-liquid ratio of 1/10, more than 99% and 98% leaching rates of cobalt and nickel, respectively, can be achieved within 4 h of leaching. In addition, the leaching performance is relatively insensitive to the change of ascorbic acid concentration from 2 to 4 g/L which is highly desirable from operational perspective.     

稀土Y掺杂非晶态纳米Ni(OH)2的结构及其电化学性能研究

以Tween-80/n-C4H9OH/c-C6H12/NiSO4水溶液体系,采用微乳液快速冷冻沉淀法制备出稀土Y掺杂非晶态纳米级氢氧化镍粉体材料.采用XRD、SAED、SEM、TEM、EDS、Raman、IR,粒度分析和比表面等测试方法对所制备的粉体进行了结构形态表征,并对其充放电性能和交流阻抗谱进行测试.结果发现,适量稀土元素Y的掺入使非晶态纳米氢氧化镍的结构缺陷增多、无序性增强,平均粒度减小、比表面积增大,有利于降低其溶液电阻、电荷转移电阻和Warburg阻抗,从而提高其放电比容量.样品作为MH-Ni电池正极材料以0.2 C充放电,终止电压为1.0 V,当掺杂Y的质量分数为4%时,放电比容量达到333.3 mAh/g.     

从废旧锂离子电池中分离回收钴镍锰

提出一种新型的从废旧锂离子电池中分离回收钴镍锰的工艺.该工艺采用物理擦洗-稀酸搅拌浸出的方法分离集流体与活性物质,采用H2SO4+H2O2为浸出剂对活性物质进行浸出,然后采用黄钠铁矾法去除浸出液中的铁,再采用N902萃取分离铜,通过水解沉淀法除铝,最后采用碳酸盐共沉淀法制备镍钴锰碳酸盐前躯体.结果表明:最优浸出条件为液固比10:1、H2SO4浓度2.5 mol/L、H2O2加入量2.0 mL/g(粉料)、温度85℃、浸出时间120 min;在此条件下,钴、镍和锰的浸出率分别达到97%、98%和96%;除去浸出液中的铁、铜和铝后,钴、镍和锰的损失率分别为1.5%、0.57%和4.56%;总体来说,废旧锂离子电池中钴、镍和锰的回收率均可以达到95%.     

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%.     

废镍催化剂回收制备纤维状氢氧化镍

研究从废旧镍催化剂中回收镍以制备氢氧化镍。采用酸浸法,在90℃下用1mol／L硫酸浸没废镍催化剂,从中提取镍。向净化后的含镍溶液中加入NaOH,分别采用3种不同的方法,即尿素水解、传统方法和水热方法制得3种不同的氢氧化镍,分别命名为Ni（OH）2-U,Ni（OH）2-C和Ni（OH）2-H。与传统方法相比,采用水热方法制得的氢氧化镍具有更好的结晶度。Ni（OH）2-C和Ni（OH）2·H都含有β-Ni（OH）2与a-Ni（OH）2·0．75H20的混合相,而Ni（OH）2．U只含有a-Ni（OH）2．0．75H20相。TEM观察显示Ni（OH）2-U样品具有棒状结构。在这3种样品中,Ni（OH）2-U表现出最好的电化学活性。     

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.     

Phase transformation in reductive roasting of laterite ore with microwave heating

Selective reduction of laterite ores followed by acid leaching is a promising method to recover nickel and cobalt metal, leaving leaching residue as a suitable iron resource. The phase transformation in reduction process with microwave heating was investigated by XRD and the reduction degree of iron was analyzed by chemical method. The results show that the laterite samples mixed with active carbon couple well with microwave and the temperature can reach approximate 1000 ℃ in 6.5 min. The reduction degree of iron is controlled by both the reductive agent content and the microwave heating time, and the reduction follows Fe2O3→Fe3O4→FeO→Fe sequence. Sulphuric acid leaching test reveals that the recoveries of nickel and iron increase with the iron reduction degree. By properly controlling the reduction degree of iron at 60% around, the nickel recovery can reach about 90% and iron recovery is less than 30%.     

Recovery of copper and tin from stripping tin solution by electrodeposition

In this paper, the method of recycling copper and tin by constant-current and constant-voltage electrolysis from the stripping tin solution of tinned copper wastes was studied. The experimental results show that the elements could be deposited on the cathode in turn by different deposition potentials, therefore, the copper and tin were separated by constant-voltage electrolysis but not constant-current electrolysis. In this study, the influence of anode materials was also investigated. Graphite anode is stable without impurities dissolved into the stripping tin solution, while 316 stainless steel anode is dissolved into Fe2？and Fe3？as anodic corrosion, which could decrease the deposition efficiency of tin. The copper and tin in the stripping tin solution are separated orderly by electrodeposition at different voltages using the graphite anode. The recovery rate of copper is up to 100 % at 2.00 V, while that of tin is 80 % at 3.00 V.     

Leaching isomorphism rare earths from phosphorite ore by sulfuric acid and phosphoric acid

Most of the phosphorite deposits in the world contain isomorphism rare earths(RE) which are considerably difficult to be leached into solution in the wet phosphoric acid process. In this work, a systematic study of leaching RE using sulfuric acid, phosphoric acid, mixed acid and two-step leaching of phosphoric acid and sulfuric acid was performed. The aims are to illuminate the main factors that inhibit RE leaching and to provide insights into the further enrichment of RE in the wet phosphoric acid solution. The results indicate that H_2SO_4 is not an effective acid for leaching isomorphism RE from phosphorite ore.The low RE leaching efficiency attributes to the RE cocrystallized and encapsulated by phosphogypsum(PG) as well as the precipitation of RE by RE sulfates or phosphates. High concentration of H_3PO_4 can enhance the dissolution and diffusion of RE ions. Hence, the optimized leaching mode of improving RE leaching efficiency is to adequately dissolve phosphorite ores in high concentration of H_3PO_4 solution and then add H_2SO_4 to crystallize PG.The effect of co-crystallization or encapsulation of PG on RE can be decreased due to the crystallizing mode of PG in the bulk solution instead of on the interface of solid reactants. RE leaching efficiency can be high up to 65% by the optimized leaching mode.     

还原酸浸法从低品位水钴矿中提取铜和钴

以Na2SO3为还原剂从水钴矿还原酸浸液中提取铜和钴,研究了还原剂种类及用量、浸出温度、硫酸浓度等因素对水钴矿还原酸浸过程中有价金属铜和钴浸出率的影响。结果表明,Na2SO3是较适宜的还原剂;在还原剂用量为水钴矿原矿质量的10%、硫酸浓度为3 mol/L、浸出温度为60℃、液固比为2-1、浸出时间为60 min的条件下,铜和钴的浸出率分别达99.06%和98.87%。并提出了"M5640萃铜→黄钠铁矾法除铁→碳酸钠除铝→氟化钠除钙、镁→蒸发结晶得钴产品"的后续分离净化流程,能有望应用于水钴矿及类似物料中有价金属的提取与分离的工业生产。     

EFFECT OF ELECTROLESS PLATING COBALT UPON CHARACTERISTICS OF NICKEL HYDROXIDE ELECTRODE FOR RECHARGEABLE ALKALINE BATTERIES

ＩＮＴＲＯＤＵＣＴＩＯＮＮｉｃｋｅｌｈｙｄｒｏｘｉｄｅｅｌｅｃｔｒｏｄｅｓｈａｖｅｂｅｅｎｗｉｄｅｌｙｕｓｅｄａｓｐｏｓｉｔｉｖｅｅｌｅｃｔｒｏｄｅｉｎＮｉＣｄ,ＮｉＨ２ａｎｄＮｉＭＨｓｅｃｏｎｄａｒｙｂａｔｅｒｉｅｓ．Ｉｍｐｒｏｖｅｍｅｎｔ...     

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。