南方稀土矿浸出液提取稀土评述

本文综述了我国南方稀土矿浸出液提取稀土的方法,并从经济和技术上加以评价和论证。作者认为液膜分离法和溶剂萃取法进一步完善后将取代传统的草酸沉淀法。     

专利文摘

二、稀土的提取和分离 1.C22B 3/00 CN85102529A 1985.4.1 赣州有色冶金研究所 离子型稀土矿提取稀土及综合利用工艺 本工艺是根据离子型稀土矿的特点,在机械分级机中同时进行浸取稀土、脱粗和除铝作业,不含稀土的粗尾砂通过筛分作业得到石英砂和钾长石粗产品,含稀土的矿浆通过水力旋     

离子型稀土矿渗流溶浸行为研究

离子型稀土矿性质复杂,当前所采用的原地溶浸工艺适应性不强、影响因素较多,为揭示离子型稀土矿原地溶浸过程影响机制,探寻溶浸过程优化调控方法,本文以江西龙南足洞风化离子型稀土矿为研究对象,采用室内柱式溶浸法模拟原地溶浸工艺过程,开展离子型稀土矿溶浸过程影响规律研究,考察浸出过程各工艺因素对离子型稀土矿浸出效率的影响以及稀土原矿浸出前后的物质变化。研究结果表明,调控离子型稀土矿浸出过程原矿含水量4%、矿层高度43 mm,浸出剂浓度4%、浸出剂溶液p H值5.0、浸取过程固液比1.0∶0.8、浸取剂流速5 ml·min-1为最佳浸出条件参数,在原矿REO总量为0.078%的情况下,可获得浸出率为92.87%的稀土母液,母液中稀土平均浓度为1.73g·L-1,效果良好。稀土离子在浸出过程中与浸出剂的电解质NH+4发生了离子交换反应,稀土离子被交换浸出,而阳离子电解质NH+4则保留在浸渣中,整个浸出过程未生成其他新的物质,没有改变离子型稀土矿原有的晶型结构,不会影响稀土离子的渗流溶浸,有利于离子型稀土矿的绿色高效提取。     

赣南地区离子型稀土矿山水环境成本量化研究

矿山项目环境成本核算对投资决策具有重要影响,这使得环境成本量化的规范性和可靠性具有更高的要求。水环境成本是离子型稀土矿山生产过程中的主要环境成本。本文以离子型稀土矿山水环境成本为研究对象,建立了水环境污染损失模型,并对赣南地区12个离子型稀土矿山的水环境成本进行了量化;分析了离子型稀土矿山生产规模与水环境成本之间的关系,给出了两个计算公式,编制了不同生产规模下水环境成本速查表,根据矿山生产规模可以快速查找其水环境成本,为离子型稀土矿山水环境成本量化提供了极大便利。研究表明,12个离子型稀土矿山中,水环境成本最高的是A1稀土矿(6015.93万元/年),最低的是A7稀土矿(390.36万元/年);在相同单位生产规模下,全覆式稀土矿水环境成本约为裸脚式稀土矿的1.62倍。本文研究成果可以为离子型稀土矿山的成本核算提供依据,对稀土矿山投资决策及生产管理具有重要意义。     

离子吸附型稀土矿再吸附试验研究

针对离子型稀土矿原地浸矿中的再吸附问题,进行了实验室条件试验,探讨了离子型稀土矿再吸附存在的条件,提出避免离子型稀土矿浸出再吸附的措施。结果表明稀土母液与稀土矿相互接触越充分,再吸附效果就越好;母液稀土离子浓度升高或母液液固比增大,会使稀土矿再吸附量增大;母液中阳离子离子交换能力大于稀土离子,会抑制稀土矿再吸附,交换能力小于稀土离子,会促进稀土矿再吸附;pH值升高再吸附量先增大后减小。同时,为避免浸出过程再吸附,应该选择合适的浸取剂浓度与液固比。     

从铁矾土矿中提取镍和钴的方法

本发明涉及从含镍和钴的铁矾土矿中提取镍和钴的方法,所述方法包括：（a）在回转窑中在还原气氛中对原料矿石进行焙烧以选择性地还原镍和钴,其中在焙烧前向所述原料矿石中加入少于2．5％（重量比）的还原剂,或不加还原剂;（b）用充气的含氨的碳酸铵溶液对还原后的矿石进行浸取,将镍和钴提取到浸取液中;（c）将该浸取液与矿石残渣分离,并通过选自含氨溶剂萃取法、沉淀法或离子交换法的方法对镍和钴进行提取。     

风化壳淋积型稀土矿浸出渗流和传质过程研究现状及展望

风化壳淋积型稀土矿富含中重稀土,主要分布在中国的南方七省,目前主要用硫酸铵作为浸取剂的化学浸出法回收稀土。浸取剂在矿体中的渗流、传质过程直接影响着稀土浸出效率,强化其渗流、传质过程可以提高稀土浸出的有效渗流速度,加快稀土离子交换速率,降低浸取剂用量,缩短生产周期,实现稀土绿色高效回收。本文分析了风化壳淋积型稀土矿渗流、传质过程的研究现状,总结了风化壳淋积型稀土矿渗流、传质过程的强化方法。结果表明,继续探索浸取液在风化壳淋积型稀土矿矿体中的有效渗流扩散规律和找寻促进稀土传质的高效传质规律,并以此为指导筛选浸取剂和助浸剂、改进注液方式、增加外力场将是强化风化壳淋积型稀土矿渗流、传质过程的主要研究方向。     

风化壳淋积型稀土矿稀土浸取剂的研究进展与展望

风化壳淋积型稀土矿中的稀土主要以离子的形式吸附在黏土矿物上,可采用盐类电解质溶液作为浸取液提取稀土.为实现稀土的绿色高效浸取,稀土浸取剂不断改进.本文主要阐述了浸取剂在助渗、抑杂、防膨以及少铵或无铵浸取等方面的研究进展,指出了现有浸取剂研究中存在的一些不足,提出了进一步加强助渗、抑杂、防膨等机理的研究,加强浸取剂残留对...     

浅析原地浸矿开采的离子型稀土矿山安全问题

南方离子型稀土矿是我国所特有的稀土矿床,多采用原地浸矿工艺开采。近年来,矿山安全生产事故频发,给企业和个人造成了巨大伤害。本文结合南方离子型稀土矿山多年开采经验,对生产中可能存在的一些安全隐患做出分析,并提出部分整改建议,为离子型稀土矿山安全生产提供借鉴。     

超声波强化浸取离子型稀土矿中稀土

利用超声波的空化作用可有效强化南方离子型稀土矿中稀土的浸出,从而提高稀土浸出率并缩短矿物中稀土总量的分析时间。在20 g/L的硫酸铵浸矿液中超声浸矿30 min,可使离子型稀土的浸出率达99%以上,而传统搅拌法需4 h,甚至浸取过夜。超声法与搅拌法对干扰杂质铁、铝的浸出率相近,加入乙酰丙酮及磺基水杨酸等掩蔽剂后不影响EDTA滴定稀土时的终点判断,且测定结果与电感耦合等离子体发射光谱法测定结果一致性好。     

电感耦合等离子质谱法测定离子型稀土原矿中离子相稀土总量

采用电感耦合等离子质谱法(ICP-MS)直接测定离子型稀土原矿中离子相稀土总量,实验对浸取溶液、浸取溶液的浓度及用量、被测元素的同位素、样品酸度、内标元素、仪器的最佳工作参数进行了选择,建立了优化分析条件.单一元素标准加入的回收率为95%~105%,离子相稀土总量回收率为98.4%,相对标准偏差小于5%,测定范围为:0.010%~0.50%.方法准确可靠,操作简单,精密度和准确度都能满足分析要求.     

从抛光废料中回收稀土的研究

探索了从抛光废料中回收稀土并确保稀土回收率在80%以上的工艺.通过分析现有抛光废料成分,有针对地提出了初步回收其中稀土成分的方法.整个试验过程分为两部分,小试主要探索原料前处理方式、酸的种类、酸的浓度、浸出温度、浸出时间和添加剂种类等因素对抛光废料中稀土回收率的影响.试验最佳条件为:加入添加剂B,用8 mol/L的盐酸在92℃下直接浸出2 h.经过除杂、萃取分离、沉淀、灼烧工序后,稀土回收率最高可达85.94%.然后根据小试确定的最佳条件进行综合扩试,所得稀土回收率能稳定在81%以上.初步概算,处理1 t该废料收益可达2000元左右.      

Migration behavior,separation and recycling of rare earth during thermal decomposition of Zhijin phosphorus ore

Zhijin phosphorus ore is a moderate and low-grade phosphorus rare-earth ore contained in mines. The separation and extraction of associated rare earth are important research topics. In this study,the migration behavior of rare earth during the thermal decomposition of Zhijin phosphorus ore and the separation and extraction of rare earth in phosphorus slag are discussed systematically. During the thermal decomposition process of phosphorus ore, almost all of the associated rare earth enters into the phosphorus slag phase but does not enter into the ferrophosphorus or gas phases. Amorphous calcium metasilicate and calcium fluosilicate are major components of phosphorus slag, and rare earth mainly exists as a calsil solid solution. Hydrochloric acid was used for acidolysis of the phosphorous slag.Under the following conditions, 96% of the rare earth in the phosphorous slag can be dissolved in the acidolysis solution: acid excess coefficient of 1.5, reaction time of 50 min and reaction temperature of 50℃. The rare earth in the acidolysis solution was separated and recycled using oxalic acid as a precipitator and NaOH as a pH modifier. At pH of 1.7, rare-earth-enriched matter with rare-earth content of 2.1 wt% was obtained, and the recovery of the rare earth was 88%.     

Separation of rare earth elements from sulfate leach liquor by heterocyclic nitrogen compound

Separation of rare earth dements by solvent extraction has actually been widely used in various fields from analytical chemistry to hydrometallurgy. A representative ore sample obtained from Kadabora Batholiths-Eastern Desert of Egypt, containing the multiple oxides rare earth minerals： Samarskite, Fergusonite, Betafite, and Pyrochlore, was subjected to sulfuric acid leaching. Different sets of equilibrium loading experiments were carded out on a bench scale for the extraction of rare earths （cerium and yttrium） from the sulfate leach liquor using 8,9-dihydro[1,2,4]triazolo[1,5-a]quinazolin-6（TH）-one {TQ} dissolved in methylene chloride. Stripping was carded out by 20% sodium hydroxide. A rare earth cake was produced by oxalic acid precipitation. Its purity reached 87.3%.     

废旧镍氢电池负极板中稀土的回收

采用湿法冶金工艺,回收废旧镍氢电池负极板中的稀土（RE）元素,用硫酸浸出负极板中的有价金属,分析硫酸浓度、浸出温度、浸出时间等因素对稀土元素浸出率的影响,在硫酸浓度为2.0 mol/L、浸出温度为60℃、浸出时间120 min下,RE的浸出率为92.31%.采用磷酸二异辛酯（P204）为萃取剂萃取浸出液中的稀土,当P204在煤油中的比率为20%时,萃取率为92.86%.用硫酸钠沉淀溶液中的稀土,浸出液中稀土元素回收率可达98.78%.采用XRD和SEM分析表征回收的稀土氧化物的物相和表面形貌,结果表明,回收产物为铈系稀土氧化物,为立方晶系,呈面心立方结构,表面形貌为棱柱形.     

稀土冶炼氨氮废水的处理技术现状

分析了稀土冶炼废水产生原因和组成成分,阐述了国内外稀土冶炼氨氮废水主要处理技术及研究现状,分别介绍了生化法、物化法及蒸馏法等处理技术原理,讨论了这些稀土冶炼氨氮废水处理技术实际运用中存在的问题.并着重针对目前风化壳淋积型稀土矿提取和分离过程中面临的氨氮达不到废水排放新标准等问题,提出了今后加强对风化壳淋积型稀土矿提取分离新技术工艺及新型药剂等的研究,探索氨氮废水循环利用新途径,从源头消除氨氮废水并从根本上治理氨氮污染.      

Non-ammonia enrichment of rare earth by magnesium oxide from rare earth leaching liquor in magnesium salt system

In order to solve the problem of ammonia-nitrogen pollution in the enrichment process of the ionadsorption type rare earth ore,the technology of non-ammonia precipitation with magnesium oxide precipitant was carried out.It is determined that the rare earth precipitation efficiency is 99.6% and the purity of rare earth concentrates is only 85.89 wt%under the optimum precipitation conditions.And the contents of MgO,SO_3 and Al_2O_3 in the rare earth concentrates are 5.12 wt%,6.77 wt%and 1.78 wt%,respectively.Furthermore,the thermo-decomposition process of precipitates was investigated by TGDSC,XRD and FI-IR.The thermal decomposition process consists of two stages:the dehydration of rare earth hydroxide and alkaline rare earth sulfate within 900 ℃ and the thermal decomposition of RE_2O_2SO_4 at 900-1300 ℃.Therefore,a high-temperature calcinations method for removing SO_3 from precipitates is proposed.When the precipitates are calcined at 1300 ℃ for 2 h,the rare earth concentrates with a purity of 92.03 wt%can be acquired.Moreover,the content of SO_3 in the concentrate is only 0.46 wt%.In the MgO precipitation and high-temperature calcinations process,the raw material cost is low and the quality of rare earth concentrates is acceptable.It could have great significance for nonammonia enrichment of rare earth from the rare earth leaching liquor,and finally solve the problem of ammonia nitrogen in the extraction process of the ion-adsorption type rare earth ore within magnesium salt system.     

Reduction leaching of rare earth from ion-adsorption type rare earths ore with ferrous sulfate

The practice of in-situ leaching of the ion-adsorption type rare earths ore with ammonium sulfate could only leach most of rare earth in ion-exchangeable phase,but not the colloidal sediment phase.Therefore,the reduction leaching of rare earth from the ion-adsorption type rare earths ore with ferrous sulfate was innovatively put forward.The soak leaching process and the column leaching process were investigated in the present study.It was determined that ion-exchangeable phase could be released,and part of colloidal sediment phase rare earth could be reduction leached by the cations with reduction properties.The mechanism of reduction leaching was discussed with the Eh-pH diagram of cerium.Moreover,the stronger reduction of reductive ions,the greater acidity of leaching agent solution,and the higher reductive ion concentration,could result in the higher rare earth efficiency and the bigger cerium partition in the leaching liquor.In the ferrous sulfate column leaching process,the rare earth leaching rate and the rare earth efficiency were a little higher than with(NH_4)_2SO_4 agent,and the rare earth efficiency and the partitioning of cerium in leaching liquor could be about 102% and 5.31%,respectively.However,the ferrous sulfate leaching process revealed some problems,so compound leaching with magnesium sulfate and a small amount of ferrous sulfate was proposed to an excellent alternative leaching agent for further studies,which may realize efficiency extraction and be environment-friendly.     

High-efficiency simultaneous extraction of rare earth elements and iron from NdFeB waste by oxalic acid leaching

Iron can not be recovered at high value because only rare earth elements are effectively recovered from NdFeB waste via oxidation roasting-hydrochloric acid leaching process.In this study,a new method for leaching NdFeB waste with oxalic acid was developed.The high-efficiency,simultaneous and high-value recovery of rare earth elements and iron was realized to simplify the process and improve the economic benefit.Results of the oxalic acid leaching experiments show that under the optimum leaching conditions at 90℃ for 6 h in the aqueous solution of oxalic acid(2 mol/L) with a liquid-solid ratio of60 mL/g,the iron leaching efficiency and precipitation rate of rare earth oxalate reach 93.89% and 93.17%,respectively.Rare earth oxalate and Fe(C₂O₄)3^3- were left in the residue and the leaching solution,respectively.The leaching mechanism was further analyzed by characterising the leach residues obtained through X-ray powder diffraction(XRD) and scanning electron microscopy-energy dispersive X-ray spectroscopy(SEM-EDS).Results of the leaching kinetics study indicate that the process of oxalic acid leaching follows the shrinking nucleus model,and the leaching kinetics model is controlled by the mixed factors of diffusion and chemical reaction.The leaching residue was calcined at 850℃ for 3 h and then decomposed into rare earth oxide,which can be directly used to prepare rare earth alloy via molten salt electrolysis.For the leaching solution,ferric oxalate solution was reduced using Fe powder to prepare the ferrous oxalate(FeC₂O₄-2H₂O).