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。      

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.     

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).     

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%。     

Dissolution Kinetics of Iron from Low-Grade Mn Ore and Optimization Using Response Surface Methodology

In this study an attempt has been made to increase Mn/Fe ratio in dump Manganese ore fines so that it can be used for the production of ferromanganese. For this purpose non-coking coal was used as reductant and dilute hydrochloric acid as leaching medium for the roasted ore. The effects of acid strength, leaching time, leaching temperature, stirring speed, ore particle size and pulp density have been studied. The dissolution of iron follows the kinetic model 1 ? 2x/3 ? (1 ? x)^2/3 = k_dt. Thus product layer diffusion is the controlling mechanism and the activation energy has been determined to be 26.23 kJ/mol at 40–95 °C. Another set of experiments have been conducted according to 2³ full factorial design, and regression equation for iron dissolution has been developed.     

A ferric chloride hydrometallurgical process for recovery of silver from electronic scrap materials

A hydrometallurgical method for the recovery of silver from electronic scrap materials is suggested. Electronic scrap materials, containing silver with the accompanying brass and beryllium bronze, are leached at 80°C with ferric chloride. The leaching should be carried out with the concentration of ferric chloride and phase ratio chosen so that in the final phase of leaching the concentration ratio of Fe(III) to Fe(II) is not lower than 1. Under such conditions silver is found to be only slightly solubilized. This was confirmed by electrochemical studies.Potentiodynamic studies of the behaviour of silver in hydrochloric acid demonstrated that at the electrode potential value of 0.77 V, corresponding to the redox potential at 25°C of the solution where the concentration ratio [Fe(III)]/[Fe(II)] = 1, the silver surface is passivated not only by silver chloride but also by silver oxide. This prevents solubilization of silver in chloride solutions owing to the formation of AgCl₂^? and AgCl₃^2? complexes. The lowering of the solution oxidation potential resulting from consumption of Fe³⁺ ions during the leaching process causes a sudden increase in silver concentration in solution.     

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 ...     

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%.     

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.     

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。     

从钕铁硼废料中提取稀土工艺研究

采用氧化焙烧-盐酸分解法,研究从钕铁硼废料中提取稀土的工艺条件,探讨了焙烧温度和时间对铁的氧化率的影响,在浸出过程中考察了盐酸浓度、反应时间、反应温度以及液固比对稀土浸出率的影响,并分析了pH值和陈化时间对浸出液除杂效果的影响.结果表明:在700℃焙烧1.5 h,铁的氧化率最高,铁基本完全氧化成三价铁,在最佳浸出条件下稀土浸出率高达到99.33%,浸出液中和除杂时,调节pH值为3.5,陈化时间大于2 h,料液中非稀土杂质含量低,特别是铁仅为0.0014 g/L,浸出液完全达到稀土萃取的要求.      

钕铁硼废料中钕、镝及钴的回收

研究了钕铁硼废料中钕、镝、钴的回收与分离,根据废料中所含元素的化学性质,选择了硫酸溶解、复盐沉淀稀土、碱转化、盐酸溶解、复盐沉淀铁及萃取分离等手段,成功的将钕铁硼废料中有价值的元素进行了提取,得到了纯度较高的氧化钕、氧化镝及氧化钴。     

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。     

从荧光粉废料中提取稀土工艺研究

采用4种方案从荧光粉废料中提取稀土元素,并考查了盐酸法提取稀土时盐酸和双氧水用量对稀土浸出率的影响,随后采用碳酸钠焙烧法提取渣中较难浸出的铈、铽,最后采用中和法对酸浸出液进行除杂。结果表明,100g物料盐酸最佳用量为150mL,双氧水用量为20mL,钇、铕浸出率可达99%。经碳酸钠焙烧—盐酸浸出后铽浸出率达到55%,除杂后铁、硅、铝含量分别降至11.47mg/L、15.93mg/L和150mg/L。     

钒钛铁精矿常压盐酸浸出分离铁钛试验研究

在常压条件下,采用盐酸浸出法对钒钛铁精矿进行选择性浸出试验,研究了液固比、浸出温度、浸出时间和盐酸浓度对铁和二氧化钛浸出率的影响,并对浸出渣结构、形貌、粒度及元素分布进行了研究。结果表明,盐酸浸出过程破坏了钒钛铁精矿中的磁铁矿物相,浸出渣表面出现了较为明显的粉化现象,铁元素进入浸出液;而钛铁矿未被破坏,仍以钛铁矿的形式存在酸浸渣中。最优浸出条件为:液固比为9∶1,浸出时间为60 min,盐酸浓度为18.6%,浸出温度为85℃。最优条件下铁的浸出率为85.41%,二氧化钛的浸出率为7.22%;酸浸渣的产率为27%,Ti O2品位约为34%。     

铝土矿盐酸浸出过程研究

以澳大利亚难处理三水铝土矿为对象、盐酸为浸出剂进行无焙烧浸出试验,考察浸出温度、矿物粒度以及浸出时间对氧化铝浸出率的影响。结果表明,优化工艺条件为:矿物粒度-55μm、浸出温度100~110℃、浸出时间120 min、盐酸浓度10%、浸出液固比100∶7。此条件下氧化铝的浸出率为95.49%,氧化铁的浸出率为96.72%。以该酸浸液为原料,使用TBP-苯体系进行铝铁分离萃取试验,在萃取温度25℃、相比O/A=1∶1、盐酸浓度1.5 mol/L,萃取时间10 min的条件下,经单级萃取,溶液中铁元素的萃取率可达95%,铝元素损失率为6%,铁铝萃取分离系数为408,经3级以上逆流萃取,铁铝分离系数可达800以上。以纯水作为反萃剂,在温度25℃、相比O/A=1∶1、反萃时间5 min、单级反萃,铁的反萃率达95%。     

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%.     

用铁鳞制备纳米Fe_2O_3的研究

以铁鳞在盐酸溶液中的浸出液为原料,采用强迫水解法制备不同形貌纳米Fe2O3光催化剂,研究了铁鳞浸出液纯度对于纳米Fe2O3的微观形貌和光催化性能的影响。结果表明:8g铁鳞与150ml物质的量浓度为3mol/L的盐酸溶液100℃回流反应2h,铁鳞的浸出率达到93%左右,且浸出液纯度较高;浸出液中的杂质离子改变了纳米Fe2O3的微观形貌;以铁鳞浸出液为原料制备出的纳米Fe2O3可见光光催化性能较好,光降解甲基蓝溶液50min后,其降解率可达82%左右。     

Announcements

A process for the recovery of uranium, vanadium and fluorine as byproducts of phosphoric acid obtained via hydrochloric acid (the cheapest locally produced acid) acidulation of phosphate ores was studied. At a redox potential of 180–200 mV, by introducing iron powder into the leaching circuit at 90°C, a relatively small amount of uranium and vanadium residue was obtained. The residue obtained was roasted with sodium chloride at 850°C to produce water-soluble sodium vanadate. Water leaching then removed the vanadium, leaving a uranium residue. The uranium was selectively leached by sodium carbonate at a redox potential of 480 mV. The product was then processed by alkaline decomposition to prepare the sodium diuranate concentrate necessary for uranium processing. Fluorine was recovered during the acid leaching process as sodium fluorosilicate by adding a stoichiometric amount of rock salt or sodium carbonate to the leaching circuit.