钕铁硼磁材二次废料制备高纯硫酸亚铁

钕铁硼磁材废料经回收稀土后产生大量的二次废料,针对该废料含铁量高的特点,对废料中的铁元素提取进行了相关研究,铁元素以硫酸亚铁产品形式回收。酸浸提铁阶段考察了酸浓度、浸提温度、反应时间、液固比(酸体积/废渣质量)、浸取次数等因素对铁离子浸出效果的影响,通过单因素实验得到较优的酸浸工艺参数：硫酸浓度6 mol/L、浸提温度80℃、反应时间120 min、液固比4∶1(mL/g)和浸取次数2次,在此条件下铁的浸出率约为97.8%。还原阶段考察了温度、反应时间及废铁屑过量系数等因素对Fe³⁺转化为Fe²⁺效果的影响,得到较优的还原工艺参数：还原温度为80℃、反应时间为120 min和废铁屑过量系数为1.2,在此优化条件下,Fe³⁺转化为Fe²⁺的转化率约为97.69%。最终采用浓缩、冷却结晶、重结晶的方法制得硫酸亚铁产品,产品纯度99.92%。     

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.     

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。      

硫酸焙烧分解包头混合稀土精矿添加铁泥的研究

在硫酸焙烧分解包头混合稀土精矿工艺中,利用处理钕铁硼废料所产生铁泥替代铁粉,通过焙烧分解、P204转型得到氯化稀土溶液,并与原工艺得到的的溶液进行对比,结果表明,使用铁泥和使用铁粉达到了相同的工艺效果,铁泥中的稀土得到了很好的回收。     

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.      

Evaluating organic acids as alternative leaching reagents for rare earth elements recovery from NdFeB magnets

This study proposes an advanced leaching method using organic acids to recover rare earth elements (REEs) from NdFeB permanent magnets from end-of-life computers hard disk drives (HDDs). The end-of-life HDDs were first dismantled in order to recover NdFeB magnets, which were then thermally demagnetized at 350 °C during 30 min before crushing in a ball mill under inert atmosphere. Scanning electron microscopy/energy dispersive X-ray spectroscopy (SEM/EDS) analyses performed on the NdFeB magnets show the heterogeneous structure containing the major matric phase Nd₂Fe₁₄B and the REEs-rich phase containing Nd and Pr oxides. Additionally, X-ray diffraction (XRD) and Mössbauer spectroscopy (MS) analyses on the ground NdFeB magnet show that grinding NdFeB magnets under inert atmosphere helps to minimize its oxidation. Chemical analysis shows that the composition of the ground sample is Nd: 22.8 wt%, Pr: 3.3 wt%, Dy: 1.2 wt%, Fe: 62.6 wt%, Co: 1.5 wt%, B: 0.9 wt%, Ni: 0.6 wt%. Diagrams of speciation and equilibrium phases (E_h vs. pH) were calculated to determine the predominance of the formed species in the REEs–organic acids systems. The influence of the organic acid type (acetic acid, formic acid, citric acid and tartaric acid), the acid concentration (10 vol%, up to saturation), and the solid/liquid (S/L) ratio (0.5%–10%) on NdFeB magnets leaching was investigated employing an optimal experimental design conceived by the statistical software JMP. Acetic acid (CH₃COOH) shows the highest leaching performance of REEs, allowing leaching yields over 90% for Nd, Dy and Pr in the acid concentration range of 1.6–10 mol/L and the S/L ratio range of 0.5%–5% at a temperature of 60 °C. The results presented in this investigation suggest that REEs can be recovered from magnets of end-of-life HDDs using an eco-friendly method assisted by organic acids.     

Recovering REEs from NdFeB wastes with high purity and efficiency by leaching and selective precipitation process with modified agents

A leaching and selective precipitation approach is proposed in this work to recover rare earth elements (REEs) from NdFeB magnet wastes collected from industry. Hydrochloric acid and oxalic acid were employed as the leaching and precipitation agents, respectively. Hexamethylenetetramine (HMTA) or tartaric acid was used as the chelating agent during leaching. Both leaching and precipitation processes were optimized individually. For leaching process, the effects of two different chelating agents, the concentrations of leaching agent, chelating agent, and temperature on the extraction and recovery yields were investigated. The optimized process based on the factorial experiment was determined to be the hydrochloric acid concentration of 6 mol/L, the tartaric acid concentration of 50 g/L, and the temperature of 313 K, by which the extraction yields of Fe and REEs up to 67.99% and 99.27%, respectively, are obtained. For the precipitation process, the optimized oxalic acid dosage and pH value were also determined. The produced RE oxide products have the purity and recovery yield up to 95.83% and 90.18%, respectively. These results indicate that the present method with low acid consumption and high product purity has advantages over many other approaches for REE recovery.     

NdFeB废料浸出工艺研究及浸出设备设计

由于中国目前对NdFeB废料回收稀土的自动化程度低,产品的纯度波动大,难以保证生产的安全性和运行的可靠性,导致稀土的回收率指标低。以盐酸优溶法对NdFeB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NdFeB废料浸出工艺过程进行研究;最后对浸出设备进行设计。稳定了稀土回收的纯度,提高了稀土的回收率指标,降低了生产成本,对NdFeB废料浸出工艺自动化和绿色生产的实现有积极的促进意义。     

Preparation of Sm2O3 and Co3O4 from SmCo magnet swarf by hydrometallurgical processing in chloride media

The recycling of rare earth elements(REE) from end-of-life REE based products is an environment friendly proposition. Waste Sm-Co based permanent magnet generated during machining is a good source for both Sm and Co. In the present study a simpler process of acid leaching at 80 ℃ followed by solvent extraction, oxalate precipitation and calcination is described for producing pure Sm_2 O_3 and Co_3 O_4. With either 10 vol% H_2SO_4 or 15 vol% HCI at 80 ℃ more than 95% Sm and Co are leached in 1 h.Extraction of Sm from sulphate leach liquor with TBP or Aliquat 336 was poor. Although extraction with TOPS-99 is quantitative but Sm from sulphate leach liquor precipitated as Sm_2(SO_4)_3·8 H_2 O. The chloride leach liquor at an initial pH of 2.5 and with 1.2 mol/L TOPS-99 shows requirement of 4-stages at A:O = 3:2. Stripping with oxalic acid precipitates Sm-oxalate which is calcined at 800 ℃ to produce Sm_2 O_3. Co_3 O_4 is recovered from the raffinate through oxalate precipitation followed by calcination at450℃.     

Recovery of yttrium from deep-sea mud

Deep-sea mud rich in rare earth yttrium has received lots of attention from the international community as a new resource for Y. A novel process, which mainly includes acid leaching, solvent extraction, and oxalic acid precipitation-roasting, is proposed for recovery of Y from deep-sea mud. A series of experiments were conducted to inspect the impacts of various factors during the process and the optimum conditions were determined. The results show that the Y of deep-sea mud totally exists in apatite minerals which can be decomposed by hydrochloric acid and sulfuric acid solution. The highest leaching efficiency of Y is 94.53% using hydrochloric acid and 84.38% using sulfuric acid under the conditions of H~+concentration 2.0 mol/L, leaching time 60 min, liquid-solid ratio 4:1 and room temperature 25 ℃(only in case of sulfuric acid, when using hydrochloric acid, the leaching temperature should be 60 ℃). Because of the much lower leaching temperature, sulfuric acid leaching is preferred. The counter current extraction and stripping tests were simulated by a cascade centrifugal extraction tank device. Using 10 vol% P204,15 vol% TBP and 75 vol% sulfonated kerosene as extractant, 98.79% Y~(3+) and 42.60% Fe~(3+) are extracted from sulfuric acid leaching liquor(adjusted to pH = 2.0) after seven-stage counter current extraction with O/A ratio of 1:1 at room temperature, while other metals ions such as Al~(3+), Ca~(2+), Mg~(2+)and Mn~(2+) are almost not extracted. The Y~(3+) in loaded organic can be selectively stripped using 50 g/L sulfuric acid solution and the stripping efficiency reaches 99.86% after seven-stage counter current stripping with O/A ratio of 10:1 at room temperature, while only 2.26% co-extracted Fe~(3+) is stripped. The Y~(3+) of loaded strip liquor can be precipitated by oxalic acid to further separate Y~(3+) and Fe~(3+). The precipitation efficiency of Y~(3+) in loaded strip liquor can be 98.56% while Fe~(3+) is not precipitated under the conditions of oxalic acid solution concentration 200 g/L, quality ratio of oxalic acid to Y of 2, and precipitation time 0.5 h. And the precipitate was roasted at 850 ℃ for 3 h to obtain the oxide of Y in which the purity of Y_2 O_3/REO is 79.02% and the contents of major non-rare earth impurities are less than 0.21%.Over the whole process included acid leaching, solvent extraction, and oxalic acid precipitation-roasting,the yttrium yield is 82.04%.     

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。      

Extraction of scandium from red mud by acid leaching with CaF_2 and solvent extraction with P507

The extraction of Sc by acid leaching with CaF_2 and solvent extraction with P507 from red mud was proposed.The influence of acid leaching and solvent extraction on recovery of Sc was investigated.The CaF_2 can obviously improve the leaching efficiency of Sc and reduce the acid consumption.The leaching efficiency of Sc increases from 74% to 92% and the dosage of acid reduces under suitable conditions by adding 5% CaF_2.The minerals in red mud can easily be decomposed and leached into the acid solution with CaF_2 through analysis of XRD pattern.The particles of red mud become smaller and multihole.The Sc can be selectively extracted with 10% P507 at the pH value of 0.1 from the acid leaching solution.More than 98% of Sc and less than 10% of Al and Fe are extracted.The SC_2O_3 with purity of 99% is obtained after the process of reverse extraction with NaOH,H_2SO_4 dissolution,precipitation by oxalic acid and roasting at 750℃.     

Simultaneous recovery of rare earth elements and phosphorus from phosphate rock by phosphoric acid leaching and selective precipitation:Towards green process

Phosphate rock has been considered as one of the potential promising resources for rare earth elements(REEs). But the cost issues and the technical challenges caused by the low content of REEs in ores did hinder the further development of REEs recovery technologies. In order to explore a green process for the recovery of REEs from phosphate rock, this study investigates the effects of phosphoric acid concentration, liquid-to-solid ratio(L/S ratio), leaching time and temperature on the leaching efficiencies of the major components from phosphate rock. A REEs recovery of 94.3% and a phosphorus recovery of 95.3%are achieved under the optimal conditions of attacking phosphate rock using 30%P_2 O_5 acid with an L/S ratio of 10:1 and a stirring speed of 250 r/min at 25 ℃ for 4 h. Then,the selective precipitation of REEs with 81.3% REEs recovery is realized by heating up the leaching solution from 25 to 90 ℃ and keeping for4 h. Thereafter, more than 95% phosphoric acid is recovered by H_2 SO_4 and high purity gypsum, more than95% CaSO_4(tested by XRF), is also produced at the same time. Ultimately, a green process that leaches phosphate rock with H_3 PO_4, selectively precipitates REEs from leaching solution by heating up, recovers H_3 PO_4 with H2 SO4 is proposed. Compared with REE recovery in traditional processes, this process owns the merits of simple operation, energy saving and minimum wastes.     

复杂稀有金属矿综合利用新工艺

研究一种从复杂稀有金属矿中综合回收稀土、铌、钛的新工艺。按6∶5的酸矿质量比添加浓硫酸混匀后在400℃酸化,酸化渣浸出后,浸出液按1∶1的体积比加水在100℃水解60min得到水解沉淀,浸出渣采用强磁选分离得到磁性物及非磁性物。将水解沉淀与磁性物混匀在1 800℃还原熔炼,获得Nb2O5含量22.38%、铁品位52.32%的铌铁合金和TiO2含量35.12%的钛渣,铌、钛回收率分别为66.89%和50.38%。水解液在通入空气的条件下用氨水调节至pH=3进行固液分离,按理论量1.2倍添加草酸沉淀稀土,最后将该稀土沉淀在950℃煅烧60min,可得到REO含量92.4%的稀土氧化物,稀土总回收率71.32%。     

Recovery of rare earth and cobalt from Co-based magnetic scraps

The reuse of RE and cobalt in Co-based magnetic scraps was studied.The optimized feat lixiviated condition was:200 mesh,sulfuric acid dosage was of 1.4 times theoretic dosage,temperature was 80 oC and leaching time 1 h.The optimum technology conditions was:Na2S2O8 dosage was of 8 times theoretic dosage,oxidation temperature 80 oC,oxidation time 2 h and pH=4.5.Rare earth was precipitated by saturated(NH4)2C2O4 solution,after roasting of rare earth oxalate,rare earth oxide was received.Cobalt-iron residue was soaked by hydrochloric acid,the Fe(OH)3 was preferential solution,pH was adjusted to 1.4 by hydrochloric acid,Co(OH)3 did not dissolve,cobalt and iron were separated,after roasting of Co(OH)3,cobalt oxide was received.The total recovery of cobalt was found to be 97% and rare earths was 96%.     

Processing Technology of Eudialithe Concentrates

Kola eudialithe concentrates contain 7-10% ZrO₂, about 2% rare earth metals, and 0·5-0·7% Nb₂O₃, and about 50% SiO₂. The rare earth fraction contains 40% of yttrium group metals including 20% of yttrium. Eudialithe is easily dissolved in acids but silica precipitates in colloidal nonfilterable form and carries away up to 40% of Zr content. In sulphuric acid leaching there is additional problem of coprecipitation of sodium-lanthanides sulphates. Solutions to these problems were proposed by adding fluoride ions to sulphuric acid:

?Output of zirconium to solution grew up to 95%.

?The rate of filtration of silica increased 100 times, the precipitate became granular and white.

?It became possible to separate heavy, quickly sedimenting double Na-Ln sulphates from light suspension of silica.

The heavy precipitate contains Na-Ln sulphates, monazite (LnPO₄ + ThSiO₄), and Na-Fe silicate eguirine. Lanthanides are dissolved by alkali or Ca-Ba-Al nitrates mixtures solutions. Then they are treated with nitric acid and any of 15 rare earth metals can be separated by well-known methods. The residue contains Nb, Ti, Mg, Ca, Fe, Si admixtures.

The zirconium sulphate acid solution is extracted with quarternary ammonium base organic solutions. Fluoride ion addition into Zr-containing organic phase increases extraction coefficients 2?3 times. The high quality silica precipitate can be used in rubber, plastics or silica gel production. After reextraction by weak HC1 solutions we obtain pure ZrO₂ product.

A scheme is proposed based on the use of standard industrial equipment and technologies.     

Decomposition of mixed rare earth concentrate by NaOH roasting and kinetics of hydrochloric acid leaching process

A new clean extraction technology for the decomposition of Bayan Obo mixed rare earth concentrate by NaOH roasting is proposed.The process mainly includes NaOH roasting to decompose rare earth concentrate and HCl leaching roasted ore.The effects of roasting temperature,roasting time,NaOH addition amount on the extraction of rare earth and factors such as HCl concentration,liquid-solid ratio,leaching temperature and leaching time on the dissolution kinetics of roasted ore were studied.The experimental results show that when the roasting temperature is 550℃ and the roasting time is 60 min,the mass ratio of NaOH:rare earth concentrate is 0.60:1,the concentration of HCl is 6.0 mol/L,the ratio of liquid to solid(L/S) 6.0:1.0,and the leaching temperature 90℃,leaching time 45 min,stirring speed 200 r/min,and the extraction of rare earth can reach 92.5%.The relevant experimental data show that the process of HCl leaching roasted ore conforms to the shrinking core model,but the control mechanism of the che mical reaction process is different when the leaching temperature is different.When the leaching temperature is between 40 and 70℃,the chemical reaction process is controlled by the diffusion of the product through the residual layer of the inert material.The average surface activation energy of the rare earth element is E_a=9.96 kJ/mol.When the leaching temperature is 75-90℃,the chemical reaction process is controlled by the interface transfer across the product layer(product layer interface mass transfer) and diffusion.The average surface activation energy of rare earth elements is E_a=41.65 kJ/mol.The results of this study have certain significance for the green extraction of mixed rare earth ore.     

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...     

The use of computational thermodynamic for yttrium recovery from rare earth elements-bearing residue

The rare earth elements are considered critical metals,due to the risk of supply interruption.The recycling of waste electrical and electronic equipment can be an alternative to supply the rare earth market.Several processes have been developed,and by aqueous media is the most prominent,which makes possible the extraction and separation of elements even in low concentration(traces).As an example of thermal processing,the use of thermodynamic simulations might benefit the metal extraction in hydrometallurgical processing.For this reason,the goal of this work is to evaluate the use of FactSage 7.2 software for the leaching of fluorescent lamp powders by sulfuric acid.The effect of concentration and temperature was evaluated.Results comprise that the thermodynamic software wellpredicted the solid phase formed in all residues of leaching experiments-gypsum was predicted by the software and identified in XRD analyses.It demonstrates that FactSage software can be explored for metals extraction in aqueous media,being important for trace-elements extraction.Yttrium extraction reaches up to 95%at 45℃using H₂SO₄2 mol/L.     

废FCC催化剂吸附苯酚废水

将炼油厂排放的废FCC(fluid catalytic cracking,流化催化裂化)催化剂粉末(含稀土元素2%)进行酸浸处理,其中的稀土及有价金属进入浸出液后得以回收,粉末残渣经过水洗、干燥后,分析其结构,并进行吸附苯酚废水的实验,考察吸附时间、苯酚的初始浓度、吸附温度及吸附方式等对苯酚吸附量的影响,分析废FCC催化剂残渣对苯酚的等温吸附特性。结果表明,吸附达到平衡需要60 min;pH在7.0~8.0时,苯酚的吸附量较大;苯酚的初始浓度越高,吸附量越大;温度对苯酚吸附量的影响较显著;Freundlich方程和Langmuir方程都可用于描述废FCC催化剂残渣对苯酚的等温吸附特征,Langmuir方程更为准确,通过计算,理论饱和吸附量为36.751 mg/g,实验测得饱和吸附容量为35.540 mg/g。