Selective leaching and recovery of V as iron vanadate from industrially generated Mo-V residue

A commercial process was developed to treat a Ca-based Mo-V residue generated in Mo processing plant. Vanadium was selectively leached using acetic acid and recovered as iron vanadate by hydro process. Process conditions for selective V leaching and iron vanadate precipitation were investigated. V recovery efficiency of 90.3% was achieved with a V content of 26.5% and an Fe content of 29% in the iron vanadate cake suitable for ferrovanadium industry. The overall Mo recovery in the whole process was 98.6%. The obtained leach residue containing 14.3% Mo with negligible impurities can be used as a feed material for the Mo production process or ferromolybdenum industry. This simple and economical process generates two product streams from a single operation and has the potential to solve a long standing problem of handling such a mixed Mo-V residue.     

Leaching of iron concentrate separated from kiln slag in zinc hydrometallurgy with hydrochloric acid and its mechanism

It is taken as a novel prospective process to treat iron concentrate from hydrometallurgical zinc kiln slag for comprehensive utilization of valuable metals by a hydrochloric acid leaching-spray pyrolysis method. The leaching mechanism of different valuable metals was studied. The results revealed that the leaching rates of Ag, Pb, Cu, Fe, As and Zn were 99.91%, 99.25%, 95.12%, 90.15%, 87.58% and 58.15%, respectively with 6 mol/L HCl and L/S ratio of 10:1 at 60 °C for 120 min. The action of SiO₂ in leaching solution was also studied. The results showed that the precipitation and settlement of SiO₂(amorphous) adsorbed part of metal ions in solution, which greatly inhibited the leaching of Cu, Fe, As and Zn, so it is crucial to control the precipitation of amorphous SiO₂.     

Phase transformation in reductive roasting of laterite ore with microwave heating

Selective reduction of laterite ores followed by acid leaching is a promising method to recover nickel and cobalt metal, leaving leaching residue as a suitable iron resource. The phase transformation in reduction process with microwave heating was investigated by XRD and the reduction degree of iron was analyzed by chemical method. The results show that the laterite samples mixed with active carbon couple well with microwave and the temperature can reach approximate 1000 ℃ in 6.5 min. The reduction degree of iron is controlled by both the reductive agent content and the microwave heating time, and the reduction follows Fe2O3→Fe3O4→FeO→Fe sequence. Sulphuric acid leaching test reveals that the recoveries of nickel and iron increase with the iron reduction degree. By properly controlling the reduction degree of iron at 60% around, the nickel recovery can reach about 90% and iron recovery is less than 30%.     

Application of suspension magnetization roasting as technology for high-efficiency separation of valuable iron minerals from high-iron bauxite

A technology for suspension magnetization roasting?magnetic separation was proposed to separate iron minerals for recovery. The optimum parameters were as follows: a roasting temperature of 650 °C, a roasting time of 20 min, a CO concentration of 20%, and particles with a size less than 37 μm accounting for 67.14% of the roasted product. The total iron content and iron recovery of the magnetic concentrate were 56.71% and 90.50%, respectively. The phase transformation, magnetic transition, and microstructure evolution were systematically characterized through iron chemical phase analysis, X-ray diffraction, vibrating sample magnetometry, X-ray photoelectron spectroscopy, and transmission electron microscopy. The results demonstrated the transformation of hematite to magnetite, with the iron content in magnetite increasing from 0.41% in the raw ore to 91.47% in the roasted product.     

Recovery of iron from waste ferrous sulphate by co-precipitation and magnetic separation

Magnetite concentrate was recovered from ferrous sulphate by co-precipitation and magnetic separation. In co-precipitation process, the effects of reaction conditions on iron recovery were studied, and the optimal reaction parameters are proposed as follows: n(CaO)/n(Fe²⁺) 1.4:1, reaction temperature 80 °C, ferrous ion concentration 0.4 mol/L, and the final mole ratio of Fe³⁺ to Fe²⁺ in the reaction solution 1.9–2.1. In magnetic separation process, the effects of milling time and magnetic induction intensity on iron recovery were investigated. Wet milling played an important part in breaking the encapsulated magnetic phases. The results showed that the mixed product was wet-milled for 20 min before magnetic separation, the grade and recovery rate of iron in magnetite concentrate were increased from 51.41% and 84.15% to 62.05% and 85.35%, respectively.     

Ida~（2-）-H_2O体系浸出低品位氧化锌矿

采用Ida2--H2O体系（亚氨二乙酸盐水溶液）处理高碱性脉石型低品位氧化锌矿,考察浸出时间、液固比、配体总浓度、温度及pH值对矿物中主金属Zn及杂质元素Ca、Mg、Cu、Ni、Fe、Pb、Cd的溶出影响。结果表明：在弱碱性Ida2--H2O体系中,Ca、Mg、Fe不会被大量溶出,有价金属Cu、Ni、Pb、Cd可部分随主金属Zn溶出而进入浸出液;在浸出时间4h、液固比5：1、配体总浓度0.9mol/L、温度70℃、pH8的优化条件下,锌浸出率为76.6%。     

离子型稀土矿中铝铁杂质的浸出与抑制规律

通过红外光光谱测定、XRD检测等测试方法分析了稀土矿浸出过程中各种矿物表面性质的变化,稀土离子及铝、铁杂质离子与浸出剂和抑制剂的浸出交换过程及规律。结果表明,抑制剂的添加会与稀土矿中的铝、铁等杂质离子反应,形成化合物,从而降低浸出母液中铝、铁杂质离子含量,但不会影响离子型稀土的交换浸出过程。在机理分析的基础上,采用对铝铁杂质有高效抑制效果的抑杂剂LG-01进行离子型稀土矿抑制铝铁杂质的浸出实验研究。结果表明,在不影响离子型稀土矿稀土离子浸出率的情况下,LG-01能有效降低离子型稀土矿浸出母液中铝、铁等杂质离子含量,去除率可达92%。     

铜渣中铁组分的直接还原与磁选回收

以褐煤为还原剂,采用直接还原-磁选方法对含铁39.96%(质量分数)的水淬铜渣进行回收铁的研究.在原料分析和机理探讨基础上,提出影响铜渣中铁回收效果的主要工艺参数,并进行试验确定.结果表明:在铜渣、褐煤和CaO质量比为100:30:10,还原温度为1 250℃,焙烧时间为50 min,再磨细至85%的焙烧产物粒径小于43μm的最佳条件下,可获得铁品位为92.05%、回收率为81.01%的直接还原铁粉;经直接还原后,铜渣中的铁橄榄石及磁铁矿已转变成金属铁,所得金属铁颗粒的粒度多数在30 μm以上,且与渣相呈现物理镶嵌关系,易于通过磨矿实现金属铁的单体解离,从而用磁选方法回收其中的金属铁.     

Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue (ZLR), produced from traditional zinc hydrometallurgy process, is not only a hazardous waste but also a potential valuable solid. The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR. The ZLR was initially roasted with ferric sulfate at 640 °C for 1 h with ferric sulfate/zinc ferrite mole ratio of 1.2. In this process, the valuable metals were efficiently transformed into water soluble sulfate, while iron remains as ferric oxide. Thereafter, water leaching was conducted to extract the valuable metals sulfate for recovery. The recovery rates of zinc, manganese, copper, cadmium and iron were 92.4%, 93.3%, 99.3%, 91.4% and 1.1%, respectively. A leaching toxicity test for ZLR was performed after water leaching. The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.     

草酸配位浸出二段焙砂中包裹金的赤铁矿

草酸根(ox^2-)对三价铁具有强的配位能力,可用草酸配位浸出二段焙砂中包裹金的赤铁矿,提高金的回收率。考察了草酸用量、液固比、浸出温度和时间对二段焙砂中铁浸出率的影响。结果表明,用1.17倍理论量的草酸在液固比为12 mL/g时于90℃浸出2 h,铁浸出率达到75.8%以上。除铁渣进一步氰化浸出,渣中金品位为8.8 g/t,低于直接氰化浸出渣12.3 g/t的金品位。草酸浸出液主要成分为具有光催化活性的Fe(ox)⁺和Fe(ox)2^-,可采用光催化法回收铁、再生草酸,再生的草酸可返回浸铁过程。     

Ida~(2-)-H_2O体系浸出低品位氧化锌矿(英文)

采用Ida2--H2O体系(亚氨二乙酸盐水溶液)处理高碱性脉石型低品位氧化锌矿,考察浸出时间、液固比、配体总浓度、温度及pH值对矿物中主金属Zn及杂质元素Ca、Mg、Cu、Ni、Fe、Pb、Cd的溶出影响。结果表明:在弱碱性Ida2--H2O体系中,Ca、Mg、Fe不会被大量溶出,有价金属Cu、Ni、Pb、Cd可部分随主金属Zn溶出而进入浸出液;在浸出时间4h、液固比5:1、配体总浓度0.9mol/L、温度70℃、pH8的优化条件下,锌浸出率为76.6%。     

改性含钛高炉渣中铁、钒和钛的回收

提出两段氧化—碱浸—酸浸工艺来回收改性含钛高炉渣中的铁、钒和钛.较佳的提铁实验条件为一段氧化时间40 s和保温时间8 min,铁的回收率为89.93％.较佳的提钒实验条件为总氧化时间126 s、NaOH浓度4.0 mol/L、浸出温度95℃、浸出时间90 min和碱浸循环次数4,钒的浸出率为92.13％.较佳的提钛实验...     

铂钯精矿湿法分离Bi、Fe富集贵金属工艺

以铜阳极泥分铜液还原所得铂钯精矿为原料,根据其矿物特性选择HCl作为浸出剂湿法脱除Bi、Fe等主要贱金属元素,富集Au、Ag、Pt、Pd等贵金属;通过热力学计算绘制Bi(Ⅲ)、Fe(Ⅲ)在盐酸体系的组分分布图,实验考察了HCl浓度、Cl-浓度、反应温度和时间等因素对Bi、Fe浸出率的影响.结果表明:在HCl浓度为2 m...     

铁锰多金属矿综合利用新工艺

以生物制剂KZSH01作为还原剂与铁锰多金属矿发生氧化还原反应,研究了铁锰矿的细度和还原剂的含量对还原效果的影响,考察了还原过程中温度和物相的变化,探讨了H2SO4用量对Mn和Zn浸出效果的影响.结果表明:生物还原剂KZSH01可使铁锰矿中93.0%MnO2转化为MnO,90.0?2O3转化为Fe3O4;Mn和Zn的浸出率均大于90.0%;Fe的磁选回收率大于85.0%,79.0%Pb和82.5%Ag富集在渣中.     

钢铁厂烧结机头灰中铅的配位浸出行为与综合回收

针对钢铁厂烧结机头灰中富含铅、铁、碳、钾、氯等多种有价元素的特点,根据氯离子与铅配位的特性,采用配位浸出的方式实现铅与铁、碳等元素的选择性分离回收。SEM-EDS、XRD等研究分析表明,烧结机头灰中铅主要以絮状的KPb2Cl5等物相吸附于铁氧化合物、硅铝酸盐和碳颗粒表面,铁主要以Fe2O3和Fe3O4物相存在。实验考察了溶液pH值、温度、氯离子浓度、浸出时间和液固比等因素对铅浸出率的影响。研究表明,在溶液pH值为3.0,浸出温度为80℃,氯离子浓度为6 mol/L,液固比(mL/g)为10:1,浸出时间为2 h的优化条件下,烧结机头灰中铅化合物与氯发生配位溶解反应生成PbCl2 i i-(i=1~4)等易溶解的络合离子,实现铅的浸出,铅浸出率为95.7%;而烧结机头灰中对钢铁冶炼有用的铁、碳、硅、铝等元素不被浸出,富集在浸出渣中,较好地实现了选择性浸出。浸出液中的铅经冷却结晶、洗涤纯化后,获得纯度为99%的氯化铅产品。     

Purification of metallurgical silicon using iron as impurity getter,part II: Extent of silicon purification

Purification of metallurgical grade silicon (MG-Si), using iron as the impurity getter has been investigated. The technique involves growing Si dendrites from an alloy of MG-Si with iron, followed by their separation using a gravity based technique and acid leaching. The effects of cooling rate of the alloy and the subsequent quenching temperature on the segregation of the impurities were studied. It was found that slow cooling of the alloy below the eutectic temperature causes an increase in the Si impurity concentration due to diffusion of the impurities from the alloy to the Si. Quenching the alloy from temperatures above the eutectic eliminated this effect, increasing the purity of the Si product. A significant reduction in the concentration of the major impurities was achieved, making the Si product a suitable feedstock for solar grade silicon generation. The concentrations, in ppmw, of some elements in the Si product are Al: 10, B: 2, Mn: 3, Ni: 3, Cr: 1, Fe: 1, P: 29. Other impurities including V, Ba, Li, Be, and Mg were all below 0.5 ppmw.     

从失效有机铑催化剂中富集铑的新工艺研究

针对失效有机铑催化剂,利用还原-磨选工艺富集铑。在还原过程中,添加剂促进铁晶粒长大并富集铑,再通过磁选分离富集含铑的铁精粉。研究表明最佳工艺参数为:还原温度1200℃,还原时间6 h,添加剂配比10%,煤粉配比5%,球磨时间45 min,磁场强度1.28×105 A/m。利用X射线衍射对磁选铁精粉进行了分析,磁选后主要物相为金属铁、铑,除去了大部分脉石。最终得到铁精矿品位88.67%,回收率92.74%,铑回收率为92.08%。本工艺具有还原温度低、收率高等特点,为失效有机铑催化剂富集提供一种途径。     

酸洗去除冶金硅中的典型杂质

以王水和氢氟酸为酸洗介质,研究酸洗对冶金硅中典型杂质元素的脱除效果。结果表明:王水和氢氟酸对金属类杂质均有明显的去除作用,但对非金属B和P去除效率差。Al杂质经王水和氢氟酸酸洗后去除率分别为80%和86.3%,Fe杂质的去除率分别为78.6%和89.4%。氢氟酸对总杂质去除率比王水提高12%,但其对Cu的去除率仅为4.8%。Cu的相态分析表明:Cu在冶金硅中主要是以Al-Si-Cu合金相存在,抗氢氟酸腐蚀性强。与氢氟酸不同,王水通过化学破碎作用使硅粉粒度变小,利于杂质相暴露。两种酸对硅粉的损失率均小于3.5%。     

Leaching and recycling of zinc from liquid waste sediments

The selective leaching and recovery of zinc in a zinciferous sediment from a synthetic wastewater treatment was investigated. The main composition of the sediment includes 6% zinc and other metal elements such as Ca, Fe, Cu, Mg. The effects of sulfuric acid concentration, temperature, leaching time and the liquid-to-solid ratio on the leaching rate of zinc were studied by single factor and orthogonal experiments. The maximum difference of leaching rate between zinc and iron, 89.85%, was obtained by leaching under 170 g/L H2SO4 in liquid-to-solid ratio 4.2 mL/g at 65 ℃ for 1 h, and the leaching rates of zinc and iron were 91.20% and 1.35%, respectively.     

Comprehensive Utilization of Iron and Phosphorus from High-Phosphorus Refractory Iron Ore

An innovative process of coal-based reduction followed by magnetic separation and dephosphorization was developed to simultaneously recover iron and phosphorus from one typical high-phosphorus refractory iron ore. The experimental results showed that the iron minerals in iron ore were reduced to metallic iron during the coal-based reduction and the phosphorus was enriched in the metallic iron phase. The CaO-SiO₂-FeO-Al₂O₃ slag system was used in the dephosphorization of metallic iron. A hot metal of 99.17% Fe and 0.10% P was produced with Fe recovery of 84.41%. Meanwhile, a dephosphorization slag of 5.72% P was obtained with P recovery of 67.23%. The contents of impurities in hot metal were very low, and it could be used as feedstock for steelmaking after a secondary refining. Phosphorus in the dephosphorization slag mainly existed in the form of a 5CaO·P₂O₅·SiO₂ solid solution where the P₂O₅ content is 13.10%. At a slag particle size of 20.7 μm (90% passing), 94.54% of the P₂O₅ could be solubilized in citric acid, indicating the slag met the feedstock requirements in phosphate fertilizer production. Consequently, the proposed process achieved simultaneous Fe and P recovery, paving the way to comprehensive utilization of high-phosphorus refractory iron ore.