NdFeB废料浸出工艺研究及浸出设备设计

由于目前企业钕铁硼废料回收稀土的自动化程度低,难以保证生产的安全性和运行的可靠性、高效益,实际生产存在产品的纯度波动大,稀土回收率指标低等问题。提出以盐酸优溶法对钕铁硼废料回收为技术依托,首先对影响稀土流失较大的浸出除杂过程进行实验研究,并总结出提高浸出率的技术措施;然后对Nd Fe B废料浸出工艺过程进行研究;最后对浸出设备进行设计。本文的研究达到了提高和稳定产品质量、增加生产率、降低生产成本和改善生产条件等要求,对整个工艺的自动化及绿色生产的实现有积极的促进意义。     

钕铁硼废料综合回收技术的研究进展

我国是全球最大的钕铁硼(NdFeB)永磁材料生产基地和消费市场。在NdFeB永磁材料生产加工过程中以及含有NdFeB永磁材料的报废产品中产生大量的NdFeB废料。对NdFeB废料进行回收再利用有助于建设稀土资源高效的循环经济体系,对保持我国稀土资源优势和环境安全具有重要的战略意义。本文对现有的NdFeB废料回收技术进行了总结,综述了直接回用法、火法冶金、湿法冶金、电化学回收工艺等多种不同NdFeB废料回收技术的作用原理和研究进展,分析了各类NdFeB废料回收技术的优劣势,并提出了未来NdFeB废料绿色、高效、可持续回收技术的重点研究方向,为稀土二次资源的高效开发利用研究提供有益参考。     

自然氧化预处理钕铁硼废料浸出过程

为降低钕铁硼废料预处理成本,探讨利用盐酸润湿-空气自然氧化法对钕铁硼废料进行预处理,并对经盐酸润湿-空气自然氧化处理的钕铁硼废料中稀土的浸出工艺和浸出动力学进行研究.结果表明:以4 mol/L HCl润湿原料,在空气中放置20 d后铁的氧化率达到92.37 %,可满足铁硼废料中稀土回收的前期处理工艺要求,降低生产成本;在浸出的过程中,当反应温度为363 K,盐酸浓度为2 mol/L、粒度为0.055~0.088 mm、液固比V_L/W_S=8:1、搅拌速率500 r/min下,反应时间为60 min后经盐酸润湿-空气自然氧化Nd-Fe-B废料中稀土的浸出率可达89.36 %;研究表明,钕铁硼废料中稀土浸出过程主要是受扩散控制,其表观化学反应活化能E=17.49 kJ/mol.      

钕铁硼废料熔分渣盐酸浸出稀土的实验研究

以钕铁硼废料经H₂选择性还原-渣金熔分处理得到的多组元熔分渣为原料进行盐酸浸出,研究了低温常压和高温高压条件下各因素对稀土浸出率的影响,并对浸出过程的动力学进行了分析。实验结果表明：低温常压浸出最优条件为盐酸浓度2.84 mol/L、液固比10∶1、时间60 min和温度85℃,稀土浸出率达到96.04%;浸出过程受扩散和化学反应混合控制,表观活化能29.25 kJ/mol,指前因子2.020 9 s^-1,与盐酸浓度和粒度相关的反应级数分别为1.49和-0.55。高温高压浸出最优条件为盐酸浓度2.03 mol/L、液固比10∶1、时间30 min和温度110℃,稀土浸出率达到98.13%;回收得到的稀土氧化物主要为Pr₄O₇和Nd₂O₃,纯度达99.56%;浸出过程属于内扩散控制,表观活化能为9.63 kJ/mol,指前因子为3.57×10^-3 s^-1。     

独居石优溶渣浸出过程研究

独居石是典型伴生铀、钍的稀土矿资源,通过现有的碱溶转化、优溶等步骤提取稀土后,所得优溶渣是富含铀、钍、稀土的重要二次资源。为与稀土提取保持一致的盐酸体系,研究优溶渣的盐酸浸出过程对整体回收工艺十分关键。采取单因素试验考察浸出过程条件对铀、钍、稀土浸出率的影响。结果表明,使用下述优化参数：盐酸浓度6 mol/L、浸出时间1.5～2 h、浸出温度60℃、液固体积质量比3 mL/g时,优溶渣中铀、钍、稀土的浸出率分别可达90%～95%、92%～93%、>60%,实现了较高的资源回收率。浸出渣的工艺矿物学分析表明,其主要由锆石、钍化合物和石英等脉石矿物组成。剩余的稀土组分则主要集中在未分解的独居石中,其余为少量磷钇矿和褐钇铌矿。试验结果可为独居石优溶渣的综合回收技术提供基础数据和支撑。     

硫化铵提硫过程中汞、银、锌的浸出行为研究

硫化锌矿采用加压浸出技术处理后,得到的浸出渣经浮选和热过滤能获得纯度较高的硫磺,但硫的回收率低,且其中有价金属不易综合利用。利用硫化铵从热过滤渣中进一步回收硫,并对提取硫磺过程中汞、银、锌的浸出行为进行研究,分析了(NH4)2S浓度、液固比和浸出时间对浸出过程的影响。研究表明,在常温,(NH4)2S浓度为1.0 mol/L,液固比为6:1,浸出时间为60min的条件下,元素硫浸出率为95.36%,Hg、Ag、Zn的浸出率分别为4.71%、33.73%、0.32%。采用蒸馏热分解多硫化铵浸出母液,元素硫回收率为95%,获得的元素硫纯度达到99.5%以上。     

浸出条件对稀土和铁浸出率的影响试验

以稀土精矿浓硫酸焙烧工艺中焙烧矿水浸过程为对象,研究了焙烧矿浸出温度、浸出时间、焙烧矿粒度等条件对稀土、铁浸出率的影响,并对水浸渣中稀土赋存状态进行了研究。研究表明,浸出温度和焙烧矿粒度对稀土、铁的浸出速率有较大影响,但对其浸出率没有影响,延长浸出时间,焙烧矿中的可溶性稀土、铁均可被浸出。水浸渣中的稀土主要以磷酸盐和氟氧化稀土形式存在,铁主要以磷酸铁形式存在,并含有少量硫化铁。     

钕铁硼废料浸出前后物相转变行为研究

为了研究钕铁硼废料浸出前后的工艺矿物学,将钕铁硼废料在650 ℃下焙烧2 h,而后用4 mol/L的盐酸浸出,得到浸出渣。通过XRF、XRD、XPS和SEM-EDS对焙烧产物和浸出渣进行表征。实验结果表明:焙烧产物中主要由Fe₂O₃、Fe₃O₄、SiO₂、NdFeO₃和Nd₂O₃等物质组成,且焙烧产物中稀土含量为16.40%;浸出后,浸出渣中无NdFeO₃、Nd₂O₃两种物质,稀土含量仅为0.66%。在XPS检测中,Fe以Fe（Ⅱ）和Fe（Ⅲ）两种价态存在于焙烧产物中,说明此温度下Fe没有被完全氧化成Fe（Ⅲ）,仍有部分Fe（Ⅱ）存在;渣中除Fe（Ⅲ）外同样检测出Fe（Ⅱ）,说明浸出过程并没有将Fe（Ⅱ）完全除去。本实验进一步完善了钕铁硼废料浸出理论,对未来钕铁硼的回收具有一定的指导意义。      

某金矿选矿厂浸出前移工艺优化实践

某金矿选矿厂采用全泥氰化炭浆提金工艺，处理量为3 000 t/d,根据生产需要进行流程改造后出现浸出效率低等问题。为进一步提高选矿厂生产效率，在现场开展提高浸出率的工艺研究。考察了浸出前移对炭浆法提金工艺流程浸出效率的影响，结果表明：浸出前移后，磨矿分级段浸出率由8.87%提高至26.16%。实践结果对类似选矿厂高效回收金提供了参考和借鉴。     

从锌浸出渣中浮选银基础条件研究

研究了湿法炼锌常规浸出工艺下酸性浸出渣的浮选性质,提出并讨论了从锌常规浸出渣中浮选回收银的关键条件,认为锌常规浸出的酸浸矿浆直接作为原矿浆用于浮选回收银的工艺路线优势明显,银浮选回收可以达到较好的回收率、产率和精矿品位指标,并能在浮选中同时有效富集回收金。     

Oxidative Roasting–Selective Pressure Leaching Process for Rare Earth Recovery from NdFeB Magnet Scrap

To recover rare earths (RE) with low acid consumption and low environmental pollution, selective pressure leaching with hydrochloric acid from roasted NdFeB scrap was explored. The phase evolution of NdFeB scrap during roasting at 800 °C as a function of time was confirmed, and after complete oxidation, its phase components consisted of Fe2O3, NdFeO3, and NdBO3. In the selective pressure leaching procedure, the optimal leaching was achieved at 110 °C for 30 min, in which the leaching rate of rare earth was 96.27% along with 13.33% of Fe. Subsequently, the effects of the hydrochloric acid dosage, the hydrochloric acid concentration and the particle size of the roasted NdFeB powder on the leaching rate of rare earth were investigated. For leaching at 110 °C for 30 min, the leaching of 13.33% Fe2O3 was derived from the Fe2O3 and NdFeO3 phases in the fully oxidized NdFeB scrap. This phenomenon was verified by the leaching of Fe from Fe2O3 of analytical purity and synthetic NdFeO3. Moreover, the leaching of Nd and Fe from the NdFeO3 phase was found to occur simultaneously. The advantages of the selective pressure leaching process using hydrochloric acid for the oxidized NdFeB scrap were comprehensively evaluated.     

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。     

离子型稀土矿浸出过程优化与分析

针对赣南不同性质离子型稀土矿浸出率不同等问题,为确定不同性质离子型稀土矿浸出最佳工艺匹配参数,进行了两种不同性质离子型稀土矿的浸出试验研究,并对其进行了工艺优化和浸出过程分析.结果表明,不同性质的离子型稀土矿浸出的工艺条件应有所不同,同时各工艺条件中的影响因素的大小关系也不一样,应根据稀土矿石性质确定浸出工艺条件.      

某离子型稀土矿浸出试验

以中国南方地区某离子型稀土矿为研究对象,采用搅拌浸出和柱浸的方式,研究不同条件下矿样中稀土及杂质元素的浸出情况,为离子型稀土矿产资源的绿色高效开采提供参考。结果表明,浸出液固比对离子相稀土浸出率影响较大,浸出时间影响较小,离子相稀土浸出过程时间短,反应迅速;柱浸过程中离子相稀土流出速率最快,达到平衡时间短,杂质元素前期浸出浓度高,后续拖尾严重;离子相稀土浸出率随着样品深度的增加不断降低,符合南方离子型稀土成矿规律;硫酸铵浸出过程中铵根离子损失量较大,最低损失率超过11.31%,硫酸根不参与金属离子的交换反应过程,回收率最高可达99.22%。     

Kinetics study on leaching of rare earth and aluminum from polishing powder waste using hydrochloric acid

In this study,a novel hydrometallurgical process consisting of hydrochloric acid three-stage countercurrent leaching and solvent extraction was proposed to recover rare earth oxide(REO) from the rare earth polishing powder waste(REPPW).The effects of HCI concentration,liquid-solid ratio(L/S ratio),temperature and time on the leaching yields of rare earths(in REO) and aluminum(in Al_2O_3) were studied.The result shows that the leaching yields of REO and Al_2O_3 are 90.96% and 43.89% respectively under the optimum leaching parameters of HCl concentration=8.00 mol/L,L/S ratio=4 mL/g,leaching temperature=353 K and leaching time=180 min.Meanwhile,the leaching kinetics of REO and Al_2O_3 were investigated in this study.The leaching behaviors of REO and Al_2O_3 follow a shrinking sphere/core model and the general leaching process is controlled by the surface chemical reaction.The leaching activation energies of REO and Al_2O_3 are 9.86 and 13.68 kJ/mol,respectively.The leaching yield of each substance in three-stage countercurrent leaching is improved substantially compared with single-stage leaching,with a change from 90.96% to 95.38% for REO and from 43.89% to 46.22% for Al_2O_3,respectively.Especially,the total concentration of REO in three-stage countercurrent leaching solution is greatly increased to above 300 g/L,and the acidity of which is decreased to ca.pH=2,which is conducive to subsequent solvent extraction directly.High purity REO(99.92%) is obtained by solvent extraction separation,oxalate precipitation and calcination.The total recovery yield of REO is 85.13%.     

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).     

Study on leaching rare earths from bastnaesite treated by calcification transition

Extracting rare earths from bastnaesite concentrate treated by calcification transition was studied through the single factor test and XRD patterns of bastnaesite after calcification and slags after leaching in HCl solution. And the effects of the main calcified parameters such as temperature, liquid/solid and calcified time on transition performance of bastnaesite were investigated. It was found that under the optimal conditions of calcification temperature of 250 oC, liquid/solid of 20 mL/g, calcification time of 180 min, the highest leaching rate of rare earth were obtained, with the leaching ratio of rare earths 83.70% and Ce 77.01%, La 90.55%, Nd 92.03%, respectively; loss rates of fluorine with different calcification conditions were always less than 1% and XRD patterns of calcification slags and leaching slags showed that fluorine existed in the form of CaF2.     

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.     

聚乙二醇对风化壳淋积型稀土矿浸出的影响

针对氯化铵作为浸取剂浸取风化壳淋积型稀土时,稀土浸取速率较低和浸出周期长等问题,使用聚乙二醇(PEG)为添加剂,通过柱浸试验模拟工业生产中的原地浸出工艺,探究不同聚合度的聚乙二醇对稀土矿渗透性和稀土浸出的影响。结果表明,聚乙二醇400、聚乙二醇1000和聚乙二醇4000溶液在稀土浸出过程中表现出较好的渗透效果。采用质量分数2.00%的聚乙二醇400、聚乙二醇1000、聚乙二醇4000分别与0.20mol/L的氯化铵复配,复配浸取剂溶液在稀土矿样中的渗透速率随着水力梯度的增大呈线性增大,符合达西定律。PEG400+NH_4Cl、PEG1000+NH_4Cl、PEG4000+NH_4Cl以及NH_4Cl溶液的稀土浸出率分别为82.61%、89.12%、70.67%和88.06%,且当聚乙二醇1000作为助浸剂时,渗流速率最大,渗透效果最佳。添加了聚乙二醇1000的复配浸取剂溶液,在浸取风化壳淋积型稀土矿中稀土的反应过程符合内扩散动力学模型。研究结果对提升复配浸取剂溶液在稀土矿体中的渗流速率有重要意义。