从蒸馏渣中回收稀土工艺研究

以稀土电解熔盐渣经矿相重构—真空蒸馏处理后的蒸馏渣为原料,采用盐酸酸浸提取蒸馏渣中的稀土,研究了酸浸时间、酸浸温度、盐酸浓度、液固比(L/S)对稀土和铁浸出率的影响。结果表明,在酸浸温度50℃、盐酸浓度4mol/L、酸浸时间1h、液固比4的较优工艺条件下,稀土浸出率高达99.88%,铁浸出率为44.43%,达到了稀土优先溶出的目的。     

微波加热低品位稀土精矿酸浸实验研究

针对稀土精矿高温酸浸焙烧钍难回收、成本高而低温酸浸焙烧又效率低的问题,采用"微波加热低温酸浸"新工艺,研究了低品位稀土精矿硫酸焙烧浸出的过程。实验首先考察了微波加热稀土精矿硫酸焙烧的升温特性,重点探讨了微波加热的焙烧温度、酸矿比、焙烧时间对酸浸矿稀土浸出率的影响,同时考察了不同焙烧温度下水浸渣中钍的残留率。实验结果表明:稀土精矿微波酸浸焙烧的升温速率随着酸矿比和微波功率的增加而加快;而且随着温度的升高、酸矿比和焙烧时间的增加,微波加热酸浸稀土精矿的浸出率提高,其浸出的最佳条件为:焙烧温度220℃,酸矿比1.5,焙烧时间8 min;此条件下的稀土浸出率为92.55%,且水浸渣中的钍未生成焦磷酸钍,可用于下一步提取。与现行的稀土精矿硫酸高温焙烧生产工艺和常规的低温酸浸焙烧工艺相比,微波焙烧低温酸浸工艺更具优势,在保证稀土较高浸出率和后续工艺能回收钍的基础上,将焙烧时间缩短为常规低温酸浸工艺浸出时间的1/15,从而提高了浸出效率。     

氢氧化钠焙烧法提取稀土电解渣中稀土的工艺研究

针对目前从氟盐体系稀土熔盐电解渣中回收稀土效率低的问题,提出了一种NaOH焙烧-盐酸优溶浸出法。系统考察了焙烧温度、焙烧时间、NaOH添加量,以及盐酸浓度、液固比、浸出温度、浸出时间对渣中稀土提取效果的影响。结果表明:在焙烧温度600℃、焙烧时间1.5h、NaOH与稀土熔盐电解渣质量比0.8∶1、盐酸浓度2mol/L、液固比8∶1、浸出温度40℃、浸出时间15min的工艺条件下,稀土浸出率为99.22%。     

氟化稀土熔盐电解渣硫酸浸出脱氟研究

氟化稀土熔盐电解渣不仅是稀土回收的重要二次资源,而且氟的回收利用也非常重要。本研究采用硫酸浸出法处理氟化稀土熔盐电解渣,使氟与稀土分离,并通过多级吸收将生成的氟化氢回收。研究了浸出温度、液固比、浸出时间、硫酸浓度对脱氟率的影响。结果表明:在浸出温度为360 ℃、液固比（体积与质量之比,单位为mL/g,下同）为2:1、粒度58~75 μm、搅拌转速恒定为300 r/min,反应3 h的条件下,氟的脱除率可达到95.28%,达到氟回收的目的。最后,通过对实验数据进行因次分析,得出氟脱除率的准数方程。      

稀土熔盐渣中有价金属综合回收技术研究现状及进展

中国是稀土生产大国，在每年的工业生产中都会产生大量的稀土固废，其中熔盐电解法制备稀土金属过程中产生的熔盐电解冶炼渣成分复杂、稀土含量较高(20%～80%,以稀土氧化物计)。结合国家稀土资源的可持续发展战略，本文针对稀土熔盐渣中稀土元素等有价金属的综合回收方法进行了综述总结，包括酸浸回收法、碱转酸浸法、盐转焙烧法等工艺，分析对比了各回收工艺的原理、特点和优劣，总结概述了各工艺综合回收稀土等有价金属时的最佳条件，为后续探究高效回收稀土等有价金属的工艺、解决稀土资源短缺问题提供了参考与技术支持。     

还原酸浸对废旧稀土荧光粉中稀土提取的影响

为解决废旧稀土荧光粉中高价态Ce、Tb氧化物浸出困难的难题, 在碱熔焙烧预处理的研究基础上, 采用还原酸浸的方法, 以抗坏血酸作为还原剂, 来提高废旧稀土荧光粉的稀土浸出率。实验表明:在HCl浓度为4 mol/L, HCl与水洗渣液固比(单位为:mL/g, 下同)为10:1、抗坏血酸用量为10%、温度为338 K、时间为120 min、搅拌转速为650 r/min的条件下, 稀土浸出率可以达到98%以上。      

从废稀土荧光粉中酸浸回收稀土的研究

从稀土荧光灯生产工艺过程产生的废稀土荧光粉中酸浸出稀土的实验结果表明,酸浸出法能够浸出废稀土荧光粉中的稀土。与用盐酸和硝酸浸出相比,用硫酸浸出废稀土荧光粉中稀土的浸出率较高,从技术、经济及环保角度考虑,优选用硫酸作为从废稀土荧光粉中浸出回收稀土的浸出剂。提高浸出反应温度、增加硫酸浓度和提升浸出器转速,都能提高稀土的浸出率。在温度45℃条件下,用2 mol.L-1硫酸浸出工艺废稀土荧光粉8 h,4种稀土Y,Eu,Ce,Tb的浸出率分别为67.9%,73.1%,66.4%,67.9%,非稀土成分Al的浸出率为39.2%。当升高温度到接近100℃进行硫酸浸出时,4种稀土Y,Eu,Ce,Tb的浸出率分别上升到80.4%,82.2%,81.4%,80.0%,非稀土成分Al的浸出率则增高到86.1%。扫描电镜图像显示废稀土荧光粉浸出前表面较平整,而其浸出渣的表面则有微小的絮状物和粒度变细,表明硫酸浸蚀废荧光粉而使稀土进入溶液中。浸出前后能谱分析显示,废稀土荧光粉浸出渣中稀土的相对含量已大大降低,表明稀土大部分已被硫酸浸出,浸出渣中的不溶物主要是C。     

采用硫酸化焙烧 — 水浸工艺从锂云母精矿中提取锂

研究了采用硫酸化焙烧—水浸工艺从Li_2O品位3.23%的锂云母浮选精矿中回收锂,考察了焙烧过程中硫酸质量浓度、酸矿体积质量比、焙烧温度、焙烧时间,浸出过程中液固体积质量比、浸出温度、浸出时间对Li_2O浸出率的影响。结果表明:在硫酸质量浓度1 127 g/L、酸矿体积质量比1.5/1、焙烧温度150℃条件下焙烧12 h后,对焙烧渣在液固体积质量比3/1、室温下浸出40 min,Li_2O浸出率达98.39%,浸出效果较好。     

从黏土型锂矿中高效浸出锂的研究

采用焙烧—酸浸的方法从某Li_2O品位为0.64%的黏土型锂矿中浸出锂,考察了焙烧时间和焙烧温度对Li_2O浸出率的影响,利用正交试验研究了酸浸工艺中浸出温度和时间、硫酸浓度和液固比对Li_2O浸出率的影响。结果表明,黏土型锂矿在600℃焙烧30min后,锂焙烧渣在浸出温度90℃、浸出时间30min、硫酸浓度1.5mol/L、浸出液固比为6的条件下搅拌浸出,Li_2O浸出率最高达92.97%,浸出效果良好。     

废镍氢电池中镍、钴和稀土金属回收工艺研究

介绍了湿法处理工艺对废镍氢电池中镍、钴、稀土(RE)金属的回收,考察了浸出时间、液固比、硫酸初始浓度及浸出温度等因素对镍、钴、稀土浸出率的影响;溶液pH、无水硫酸钠与浸出液中RE3+的摩尔比、反应温度等因素对稀土回收率的影响。通过正交试验确定了镍、钴、稀土在稀硫酸中的优化浸出条件以及产生稀土复盐沉淀的优化沉淀条件。实验结果表明,优化硫酸浸出条件为:浸出时间为3.8h,液固比为15,硫酸初始浓度为1.8mol·L-1,浸出温度80℃。在优化浸出条件下,镍的浸出率达96.8%,钴的浸出率达97.3%,稀土的浸出率达94.6%。稀土复盐的优化沉淀条件为:溶液pH为2.0,无水硫酸钠与浸出液中RE3+的摩尔比为4,反应温度为60℃。在此条件下,RE回收率为96.7%。     

Effect of roasting activation of rare earth molten salt slag on extraction of rare earth,lithium and fluorine

A new process was proposed to extract rare earth elements(REEs),Li and F from electrolytic slag of rare earth molten salt by synergistic roasting and acid leaching.Firstly,the thermodynamic analysis of roasting reaction was carried out,then the effects of roasting factors on leaching REEs,Li and F in slag were investigated.In additions,the mineral phase and morphology of molten salt slag,roasting slag and acid leaching slag were characterized,and the migration mechanism of REES,Li and F minerals...     

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.     

Rare earth double sulfates from pre-concentrated bastnasite

A mixture of rare earth double sulfates was produced from a Turkish bastnasite-containing pre-concentrate (low grade concentrate) by sulfuric acid baking, subsequent water leaching and precipitation with sodium sulfate. The results of acid baking and leaching indicated that recoveries of rare earth elements up to 90% were readily obtained and the recovery of hydrofluoric acid as a by-product was also possible. Reasonable decontamination of the rare earth double sulfate salt from impurities such as Th, Fe, Al and Mg was possible by rapid precipitation at 50 °C using 1.25 times the stoichiometric amount of Na₂SO_4. The total rare earth double sulfate content (TREDS) was > 90% and analysed 17.3% La, 15.6% Ce, 3.2% Nd, 1.1% Pr, 0.3% Sm, 0.03% Eu, 0.01% Yb and 0.02% Y together with about 0.7% Ca, Fe, Al and other impurities.     

Rare earth recovery from fluoride molten-salt electrolytic slag by sodium carbonate roasting-hydrochloric acid leaching

Fluorinated rare earth molten-salt electrolytic slag contains a considerable amount of rare earth elements,as well as a variety of heavy metals and fluorides that cause environmental pollution.Therefore,it is of great importance to fully utilise this resource.In this study,the transformation mechanism of fluorinated rare earth molten-salt electrolytic slag roasted with sodium carbonate,and the regulation mechanism of rare earth leaching under different roasting conditions were investigated with ...     

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%.     

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.     

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.     

硫酸法从锂磷铝石中提取锂工艺研究及优化

研究了硫酸法从锂磷铝石中提取锂的工艺。研究结果表明, 当矿物与硫酸质量比为1:0.4、焙烧温度为780~820℃、浸出液固质量比为1.6:1时, 锂提取率达96%以上; 将硫酸锂溶液用NaOH调节pH值为12, 可彻底去除溶液中的Al3+、Fe3+、PO₄3-杂质, 所得硫酸锂溶液用EDTA络合Ca2+后, 与Na₂CO₃溶液反应可获得电池级碳酸锂。针对混酸料呈稀糊状和物料中的氟元素难处理两大问题展开工艺优化工作。在混酸料中加入吸水性物质, 可改善物料的稀糊状态, 有利于后续工业化生产; 将锂磷铝石煅烧后再混酸焙烧, 可消除混酸料的稀糊状, 锂提取率达97%以上; 酸化时在280℃左右进行保温反应, 能驱氟、降低硫酸锂溶液中的氟离子含量, 氟可以回收; 尾渣中AlPO₄有较高的回收价值。      

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).