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.     

Processing of rare earth concentrates

The paper describes process details for extraction of rare earths from an intermediate grade concentrate of Madhya Pradesh region in India and a South African slag. The xenodme concentrate obtained from the former place was an intermediate grade （47%） rare earth phosphate containing both monazite and xenotime. The South African slag was a low-grade waste product typically containing only 4% of rare earths. The rare earth resource concentrates have been treated individually by different methods such as alkali fusion and alkali leaching to convert them into their mixed oxides. Both types of materials have been processed and greater than 98% solubilization of metal values has been achieved in the intermediate grade xenotime and 80% from the South African slag. The residue of xenotime hydroxide has been washed thoroughly to collect the sodium phosphate, as by-product and the slurry pH have been adjusted to separate rare earths from thorium effectively. Other impurities such as uranium and iron have been removed by precipitation of rare earths by oxalic acid. It has been possible to recover 〉95% yttrium along with other rare earth oxides.     

Separation and recovery of rare earths via a dry chemical vapour transport based on halide gaseous complexes

The progress in the field of mutual separation of rare earths from the rare earth ores and their recovery from rare earth containing industrial waste by the dry chemical vapour transport (CVT) process mediated by the corresponding gaseous rare earth complexes with aluminum or alkali chlorides, RAI_nCl_3+3n or KRCl₄, is reviewed and discussed on the basis of their thermodynamic characteristics. The experimental results for the mutual separation and recovery of rare earth elements show the potential of this simple dry CVT process to substitute the conventional energy consuming complicated wet methods used at present for the costly preparation of the rare earths.     

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.     

3,5-二硝基水杨酸改性硅胶脱除稀土料液中的铝

针对离子吸附型稀土矿浸出液中铝元素含量高、对稀土后续分离影响大的问题,本研究以3,5-二硝基水杨酸(DNS)作为有机配体,以硅胶作为基体材料,通过化学合成3,5-二硝基水杨酸化学键合硅胶进行吸附除铝研究。结果表明:当3,5-二硝基水杨酸与Al³⁺的配位比为1:1、溶液pH=3.5、3,5-二硝基水杨酸改性硅胶用量为2g时,溶液中Al³⁺去除率达90%以上,稀土几乎无损失。采用紫外吸收的方法对机理进行研究,结果显示稀土料液中的Al³⁺通过取代3,5-二硝基水杨酸分子中羟基上的H+与其形成络合物,进而实现稀土与铝的分离。由吸附动力学研究可知,3,5-二硝基水杨酸改性硅胶对Al³⁺的吸附符合二级吸附动力学模型,其R²为0.98813。     

Process optimization and kinetics for leaching of cerium,lanthanum and neodymium elements from iron ore waste's apatite by nitric acid

The leaching of rare earth elements (REEs) including cerium, lanthanum and neodymium from apatite concentrate obtained from iron ore wastes by nitric acid was studied. The effects of nitric acid concentration, solid to liquid ratio and leaching time on the recoveries of Ce, La and Nd were investigated using response surface methodology. The results showed that the acid concentration and solid to liquid ratio have significant effect on the leaching recoveries while the time has a little effect. The maximum REE leaching recoveries of 66.1%, 56.8% and 51.7% for Ce, La and Nd, respectively were achieved at the optimum leaching condition with 18% nitric acid concentration, 0.06 solid to liquid ratio and 38 min leaching time. The kinetics of cerium leaching was investigated using shrinking core model. It was observed that the leaching is composed of two stages. In the first stage a sharp increase in cerium leaching recovery was observed and at the longer time the leaching became slower. It was found that in the first stage the diffusion of reactants from ash layer is the rate controlling mechanism with an apparent activation energy of 6.54 kJ/mol, while in the second stage the mass transfer in the solution is the controlling mechanism.     

Leaching process of rare earths from weathered crust elution-deposited rare earth ore

In order to strengthen the leaching procedure, the chemical processes of leaching rare earths (RE) from the weathered crust elution-deposited rare earth ore were investigated frow the viewpoints of kinetics, hydrodynamic and mass transfer. The results show that the leaching hydrodynamics follows the Darcy law. The leaching kinetics can be described by the shrinking core model; the leaching process is controlled by diffusion of porous solid layer; and the mass transfer can be described with Van Deemter equation. This provides a theoretic basis and a scientific approach with high efficiency and optimized extraction conditions in industrial practice.     

La-Ce混合稀土对Mg-Al-Mn合金力学性能及耐蚀性能的影响

研究了La-Ce混合稀土对Mg-Al-Mn合金组织形貌、力学性能及耐蚀性的影响。采用T-1200CB坩埚炉冶炼稀土含量(质量分数)分别为4.63%、5.81%、6.18%的Mg-Al-Mn合金。在箱式电阻炉中对研究试样进行430 ℃保温24 h的固溶处理,然后进行200 ℃保温24 h时效处理。对不同热处理状态的试样进行组织观察,对固溶时效后的试样进行拉伸、硬度及盐雾腐蚀试验,从而分析La-Ce混合稀土对Mg-Al-Mn合金显微组织、力学性能及耐蚀性的影响。研究表明,随着合金中的La-Ce混合稀土含量的增加,Mg₁₇Al₁₂相逐渐被Al₄(La, Ce)相代替;硬度、抗拉强度和伸长率都逐渐减小,力学性能下降;合金的腐蚀速率逐渐下降,耐蚀性提高。     

RECOVERING RE FROM LEACHING LIQUOR OF RARE EARTH ORE BY EXTRACTION

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

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%.     

Recovery of rare earth elements from phosphorites in the USSR

Data on the main rare earth compositions in the USSR are presented, showing that the most promising sources are phosphorites and apatites. The rate earth oxide content in the above ore concentrates reaches 1%. Technological schemes for recovering the total and individual rare earth oxides as byproducts have been developed. Extraction methods are used for recovering rare earths from nitrate solutions, with individual oxides more than 99.9% pure being obtained. Apatites are decomposed by nitric acid, yielding apart from fertilizers, strontium carbonate, sodium fluorosilicate, chalk and oxides of rare earth elements.     

基于二元平衡解吸模型的离子型稀土浸矿剂用量计算方法

以硫酸铵作为浸矿剂浸取稀土时,其在矿土表面赋存形式分为可逆和不可逆两部分;根据铵在浸矿中的作用,铵消耗分为三个部分:离子交换消耗、专性吸附消耗和维持浸矿剂浓度的消耗.本文结合铵的赋存形式和浸矿作用,采用二元平衡解吸模型(DED模型)刻画浸矿,通过数值拟合确定相应参数,提出一种针对离子型稀土的浸矿剂用量计算方法.结果表明...     

Commercial production of rare earth metals by fused salt electrolysis

The commercial production of rare earth metals by fused salt electrolytic methods is described. These methods are used to make mischmetal, cerium, lanthanum, and didymium (Nd + Pr). The feed materials consist essentially of anhydrous chlorides of the metal to be produced, augmented by additions of nonrare earth salts to yield an electrolyte with satisfactory properties for reduction. The rare earths are derived either from monazite or from bastnasite ores. The anhydrous chlorides are manufactured from the hydrated chlorides by methods which minimize oxidation or hydrolysis. Alternatively, anhydrous chlorides may be used which result from the direct chlorination of rare earth ores by the Gold-schmidt process. These are particularly suitable for electrolysis due to their very low oxychloride content. Fluorides and oxides of the rare earths are produced by wet chemical methods to provide the relatively small quantities of these compounds now used for commercial fused salt electrolysis. Reduction cells in use today are constructed mainly from a) ceramics, b) graphite, or c) iron. The advantages and disadvantages of these types of cells are described together with the typical products which result from their use. Primary attention is given to production of mischmetal (mixed rare earth metal), which is the rare earth metal produced on the largest scale today. Brief reference is made also to production of cerium, lanthanum, and didymium (Nd + Pr) metal. Finally, a resume of the current uses of these metals is presented with reference to recent trends in the industrial applications of the rare earth metals.     

Copper and elemental sulphur from chalcopyrite by pressure leaching

Processes employing direct oxidation under an over-pressure of air or oxygen in an aqueous sulphuric acid medium have been developed in the Sherritt Gordon Laboratories for iron, nickel, cobalt, zinc and lead sulphide concentrates. This study has recently also been extended to chalcocite, Cu₂S, concentrates. The rising interest in processes employing direct aqueous oxidation is stimulated by the fact that elemental sulphur can be produced as a by-product rather than sulphur dioxide or sulphuric acid.The present paper outlines a process which features the direct pressure oxidation of the most abundant copper sulphide mineral, chalcopyrite, CuFeS₂. The optimum conditions for a practical pressure leaching step have now been developed in the laboratory which results in the production of copper sulphate solution suitable for copper winninq by electrolysis, hydrogen reduction, solvent extraction combined with electrolysis, or other means. The leach residue yields pure elemental sulphur by-product. Copper and elemental sulphur recoveries of 98 and 85% respectively have been recorded. The fastest oxidation rate, corresponding to a leach retention time of 2.5 hr, was obtained when the copper concentrate was ground to 99.5% — 325 mesh, when a 50% stoichiometric excess of concentrate over the amount of available sulphuric acid for copper was used and when the oxygen partial pressure and temperature were maintained at 500 psi and 240°F, respectively. In an idealized form, the pressure leaching reaction can be expressed as follows:—CuFeS₂ + H₂SO₄ + 1 1/40₂ + 1/2H₂O → CuSO₄ + Fe(OH)₃ + 2S°After separation of the copper sulphate solution by filtration, elemental sulphur and excess concentrate ore recovered from the iron oxide tailing by flotation. The tailing, containing iron oxide and insolubles, is rejected. The elemental sulphur is separated from the concentrate by hot filtration, solvent extraction, distillation, or other means, and the unleached chalcopyrite is recycled to the leaching step.     

Technology development for rare earth cleaner hydrometallurgy in China

China is the global leader in rare earth(RE)production using hydrometallurgical processes. Advantageous extraction techniques of rare earths from Baotou mixed rare earth minerals, bastnaesite, and ion-adsorption clays of rare earth deposits have been developed in China.The separation and purification technologies have also achieved rapid development. The industrial application processes for rare earth hydrometallurgy were summarized in the present paper. With the large demands and rapid development of rare earths, the issues of resources and environment are more prominent. This review gives an overview of the main processes that were developed in the past toward greener hydrometallurgy of rare earths in China. Based on the development of the rare earth industry,comprehensive utilization and cleaner production should still be focused on in the future, to support the sustainable development.     

Partitioning properties of rare earth ores in China

1 IntroductionChinese rare earth resources are dominant in theworld with a total amount up to 80% of the wholeworld resources. Rare earth resources in China notonly have high grade and various types, but also dis-tribute reasonably. There are many rare ea…     

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。     

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%。     

钒钛磁铁矿提钒尾渣浸取钒

采用硫酸氢氟酸次氯酸钠组合浸出体系浸取钒钛磁铁矿提钒尾渣中的钒,研究浸出过程中试剂浓度、浸出液固比、浸出温度、浸出时间、物料粒度对钒浸出率的影响。结果表明:钒的浸出率随试剂浓度、液固比、温度和时间的升高而增大;当矿物粒度小于0.20 mm时,钒浸出率有随矿物粒度变小而减小的趋势。在物料粒度0.15~0.25 mm、初始硫酸浓度150 g/L、初始氢氟酸浓度30 g/L、次氯酸钠加入量为矿量1.5%、矿浆液固比6:1、浸出温度90℃、浸出时间6 h、搅拌速度500 r/min的条件下,钒的浸出率可达85%以上。