石煤钒矿硫酸活化常压浸出提钒工艺

研究石煤钒矿的硫酸活化提钒方法。分别考察矿石粒度、硫酸浓度、活化剂用量、催化剂用量、反应温度、反应时间和浸出液固比等因素对钒浸出率的影响。结果表明:石煤提钒的优化条件为矿石粒度小于74μm的占80%、硫酸浓度150 g/L、活化剂CaF2用量(相对于矿石)60 kg/t、催化剂R用量20 g/L、反应温度90℃、反应时间6 h、液固比(体积/质量,mL/g)2:1,在此优化条件下,钒浸出率可达94%以上;在优化条件下,采用两段逆流浸出,可有效减少活化剂CaF2以及浸出剂硫酸的消耗量;经过两段逆流浸出萃取反萃氧化水解工艺,全流程钒资源总回收率可达86.9%;V2O5产品纯度高于99.5%。     

铜阳极泥处理过程中中和渣中碲的提取与制备

采用硫酸浸出二氧化硫还原方法从中和渣中制取单质碲。研究表明:采用硫酸浸出中和渣,当反应温度为30℃、反应时间为0.5 h、硫酸浓度为53.9 g/L、硫酸用量为理论用量的1.5倍时,碲浸出率为99.99%;采用亚硫酸钠还原酸浸液中碲时,碲(Ⅳ)发生水解生成二氧化碲;采用二氧化硫还原酸浸液中碲时,当反应温度为75℃、反应时间为2 h、盐酸浓度为3.2 mol/L、二氧化硫流量为0.4 L/min时,碲回收率达到99.84%。X射线衍射(XRD)分析表明二氧化硫还原得到的产物为单质碲,电感耦合等离子体发射光谱(ICP)分析表明,碲粉中碲含量为98.27%。扫描电子显微系统(SEM)分析表明,碲粉的形态为针形。     

还原酸浸法从低品位水钴矿中提取铜和钴

以Na2SO3为还原剂从水钴矿还原酸浸液中提取铜和钴,研究了还原剂种类及用量、浸出温度、硫酸浓度等因素对水钴矿还原酸浸过程中有价金属铜和钴浸出率的影响。结果表明,Na2SO3是较适宜的还原剂;在还原剂用量为水钴矿原矿质量的10%、硫酸浓度为3 mol/L、浸出温度为60℃、液固比为2-1、浸出时间为60 min的条件下,铜和钴的浸出率分别达99.06%和98.87%。并提出了"M5640萃铜→黄钠铁矾法除铁→碳酸钠除铝→氟化钠除钙、镁→蒸发结晶得钴产品"的后续分离净化流程,能有望应用于水钴矿及类似物料中有价金属的提取与分离的工业生产。     

硫酸铵焙烧法浸出镍磁黄铁矿中有价金属

将镍磁黄铁矿与(NH4)2SO4混合后在高温下焙烧,考察(NH4)2SO4用量、浸出温度、浸出时间和稀硫酸浸出液的pH值对焙烧产物中金属元素浸出率的影响,并在氨水-(NH4)2SO4混合溶液中浸出焙烧产物。结果表明:在不同情况下,Ni和Cu的浸出率较高,Mg和Fe的浸出率较低;氨性溶液有利于Ni和Cu的浸出,总氨浓度为7 mol/L时,Ni和Cu的浸出率分别为89.56%和79.35%;低pH值的稀硫酸溶液有利于Mg和Fe的浸出,pH值为0.5时,Mg和Fe的浸出率分别为61.39%和62.56%。由扫描电镜-能谱分析和XRD分析可知,矿样中Ni和Cu大部分被浸出;由于焙烧产物中部分Mg和Fe以铁酸钙和硅酸镁等形态存在,Mg和Fe的浸出率较低。     

Reductive acid leaching of valuable metals from spent lithium-ion batteries using hydrazine sulfate as reductant

Hydrazine sulfate was used as a reducing agent for the leaching of Li, Ni, Co and Mn from spent lithium-ion batteries. The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and examined. 97% of the available Li, 96% of the available Ni, 95% of the available Co, and 86% of the available Mn are extracted under the following optimized conditions: sulfuric acid concentration of 2.0 mol/L, hydrazine sulfate dosage of 30 g/L, solid-to-liquid ratio of 50 g/L, temperature of 80 °C, and leaching time of 60 min. The activation energies of the leaching are determined to be 44.32, 59.37 and 55.62 kJ/mol for Li, Ni and Co, respectively. By performing X-ray diffraction and scanning electron microscopy in conjunction with energy dispersive X-ray spectroscopy, it is confirmed that the main phase in the leaching residue is MnO₂. The results show that hydrazine sulfate is an effective reducing agent in the acid leaching process for spent lithium-ion batteries.     

Recovering indium with sulfating roasting from copper-smelting ash

A technology for recovering indium from Jinchuan copper-smelting ash was developed. Indium in the ash was first enriched to the leaching-slag in leaching process,and then recovered by sulfating roasting. The method included mixing the leaching-slag with sulfuric acid,making them into particles,roasting the mixture,and then leaching the calcine with hot water. Above 90% of indium in calcine could be dissolved into the leaching solution. The optimized conditions were determined as follows： the mass ratio of sulfuric acid to leaching slag was 0.1,the roasting time was about 1 to 1.5 h in the temperature range of 200-250℃,and the calcine was leached for 1 h with 5：1 of liquid/solid ratio at 60℃. Over 99% of indium in leaching solution was finally enriched by Zn substitution or sulfide precipitation.     

离子交换膜电解槽中电沉积金属镍(英文)

对离子交换膜电解槽中电沉积金属镍的条件进行优化。研究镍(II)及硼酸浓度、pH及温度对电沉积镍的电流效率和能耗的影响。电解槽阴极液为含硼酸的硫酸镍溶液,阳极液为硫酸溶液。采用阴离子交换膜将阴、阳极室隔开,同时维持极室间的导电性。结果表明:阴极电流效率随镍和硼酸浓度以及pH值的增加而提高,随电流密度和搅拌速率的增大而降低。得到的优化电解条件为:Ni40g/L、硼酸40g/L、温度42oC、pH6、阴极电流密度300A/m2。在该条件下的电流效率为97.15%。     

铟的还原浸出动力学及分步沉淀选择性分离（英文）

开展硫化锌精矿还原浸出高铁锌浸出渣高效浸铟及浸出液中铟选择性分离的研究。结果表明：在固体物料粒度74～105μm、反应温度90℃、浸出时间300 min、硫酸浓度1.4 mol/L的条件下,铟的浸出率达95%以上。采用收缩核模型对还原浸出动力学进行分析,不同条件下的浸出实验结果表明反应受穿过固体产物层的扩散控制,活化能为17.96 k J/mol,相对于硫酸浓度的反应级数为2.41。铁粉置换沉铜过程铜和砷的沉淀率均达99%以上。98%以上的铟从含高亚铁离子浓度的硫酸锌溶液中选择性分离,获得铟含量约为2.4%的富铟渣,经酸浸-萃取-电积工艺流程进一步处理后可得到纯铟。     

Recovery of gallium from phosphorus industry flue

The flue dust generated during electric furnace production of elemental phosphorus was investigated for the recovery of gallium. Then the flue dust was slurried with water and blended with concentrated sulfuric acid, followed by ageing. The gallium in the dust was thereby converted to soluble sulfate. The factors affecting the dust curing were investigated to understand the process chemistry of the pretreatment. The optimal curing conditions are determined as follows: the mass ratio of dust to water and acid is 1 : 1 : 1, ageing temperature and time are 200℃ and 2h, respectively. Almost all the gallium available to acid dissolution in the dust, about 90% gallium, can be extracted by leaching the cured dust at 80℃ for 1h.     

以废茶叶为还原剂在硫酸溶液中还原浸出氧化锰矿（英文）

采用废茶叶在硫酸溶液中还原浸出加蓬和湘西氧化锰矿石,探索废茶叶用量、硫酸浓度、固液比、浸出温度和反应时间对浸出过程的影响。对加蓬氧化锰矿,优化的浸出条件为:氧化锰矿与废茶叶的质量比10:4、硫酸浓度2.5 mol/L、固液比7.5:1、浸出温度368 K、浸出时间8 h;在此条件下,加蓬氧化锰矿的浸出率几乎达100%。对于湘西氧化锰矿,优化浸出条件为:氧化锰矿与废茶叶的质量比10:1、硫酸浓度1.7 mol/L、液固比7.5:1、温度368 K、浸出时间8 h;在此条件下,锰的浸出率达到99.8%。氧化锰矿的还原浸出过程符合内扩散控制模型,加蓬和湘西氧化锰矿石的还原浸出反应表观活化能分别为38.2 kJ/mol和20.4 kJ/mol。采用X-射线衍射(XRD)和扫描电子显微镜(SEM)对浸出前、后的锰渣进行表征。     

水钴矿中选择性提取铜和钴的新工艺

针对某水钴矿的特点,采取还原酸浸旋流电积新工艺选择性提取其中的铜和钴。系统考察初始硫酸浓度、温度、时间、Na2SO3用量、液固比等因素的影响,确定浸出最佳条件如下:初始硫酸浓度为75g/L,Na2SO3用量为7%,液固比L/S=4 mg/L,温度为70℃,时间为0.5 h。对浸出液进行了旋流电积提取铜和钴的探索实验研究,得到纯度分别为99.95%、99.97%的电积铜、钴产品,铜、钴的直收率分别达到98.23%和94.54%。     

富银低品位锰矿和闪锌矿的联合氧压酸浸

进行了硫化锌和含银低品位锰矿联合氧压酸浸的小型试验,以考察影响浸出的各种因素,诸如浸出温度、浸出时间、氧分压和搅拌速度.实验在2L的高压釜内进行.实验结果表明,硫化锌和含银低品位锰矿能相互促进浸出,但这种耦合作用须在一定条件下才会起作用.作者对一些浸出做了探讨.为下一步扩大试验的需要,本文给出了合理的浸出条件:浸出温度, 110℃; 浸出时间,2h; 氧分压, 0.6MPa; 硫酸用量, 1.2倍理论用量.     

Recovery of gallium from zinc concentrate by pressure oxygen leaching

Zinc concentrate with high gallium content is one of the main resources of gallium.The gallium presents in the form of isomorphism in tetrahedron coordination with sulfur in sphalerite.The research was to investigate the amenability of zinc concentrate with high gallium to pressure oxygen leaching.The particle size,sulfuric acid concentration,oxygen partial pressure,additive amount,and time of reaction were studied.The extraction yields of gallium and zinc are 86%and 98%,respectively.The optimal condition is 100 g of zinc concentrate with particle size smaller than 38 lm,sulfuric acid concentration150 g L-1,leaching temperature 150℃,leaching time120 min,oxygen partial pressure 0.7 MPa,additive amount of 0.2 wt%.     

一氧化碳还原热解硫酸镁制备高纯氧化镁

利用菱镁矿和工业纯硫酸制备七水硫酸镁,进一步脱水得到无水硫酸镁。以无水硫酸镁和CO气体为原料,经高温煅烧制备高纯氧化镁,考察硫酸镁粒径、CO气体流量、煅烧温度和煅烧时间对硫酸镁转化率的影响。通过单因素实验确定的最佳工艺条件为:粒径51.8μm,CO气体流量30 mL/min,煅烧温度800℃,煅烧时间2.0 h。样品氧化镁经X射线衍射和扫描电镜表征,纯度达到99%,其平均粒径为40nm,单分子表面为多孔蓬松状。经实验和动力学推导得到描述热解过程的经验方程,CO气体还原热解硫酸镁的动力学模型为收缩核动力学模型,热解表观活化能为234.055 kJ/mol,该模型表明热解过程中的控制步骤决定于氧化还原速率。     

氧化锌贫矿提锌渣中铅和银的氯盐一步浸出

运用X射线衍射、扫描电镜和X射线能谱等分析手段,对山东某地深度氧化锌贫矿提锌后渣进行工艺矿物学特征分析可知,矿物中金属赋存状态复杂,属难选矿物。开发出氯盐一步法浸出铅和银的新工艺,考察反应温度、NaCl浓度、添加剂用量、液固比、HCl加入量和浸出时间对浸出过程的影响。结果表明:加入添加剂对铅的浸出率没有影响,但可以显著提高银的浸出率。条件试验研究得出最佳工艺条件如下:浸出温度90℃、NaCl浓度390 g/L、添加剂用量15 mL、液固比(质量比)7?1、HCl加入量3 mL、浸出时间3 h。在此最佳工艺条件下,铅的浸出率达到95%左右,银的浸出率达到90%左右。     

Pressure leaching of zinc silicate ore in sulfuric acid medium

Zinc silicate ore was characterized mineralogically and the results showed that zinc exists mainly as hemimorphite and smithsonite in the sample. Sulfuric acid pressure leaching of zinc silicate ore was carried out to assess the effect of particle size, sulfuric acid concentration, pressure, reaction time and temperature on the extraction of zinc and the dissolution of silica. Under the optimum conditions employed, up to 99.25% of zinc extraction and 0.20% silica dissolution are obtained. The main minerals in leaching residue are quartz and small amounts of undissolved oxide minerals of iron, lead and aluminum are associated with quartz.     

镍红土矿高压酸浸过程的金属元素浸出行为

以镍、钴的提取为目的,研究褐铁矿型镍红土矿高压酸浸过程中各金属元素的浸出行为,探讨硫酸加入量、浸出温度、浸出时间及液固比对各金属元素浸出率的影响.实验结果表明,在优化条件下Ni、Co、Mn和Mg的浸出率分别达到97%、96%、93%和95%以上,则Fe的浸出率小于1%.对高压浸出渣的分析表明,渣中的铁和硫主要分别以赤铁...     

Recovery of valuable metals from anode material of hydrogen-nickel battery

Simultaneous recovery of rare earth, nickel and cobalt resources from the anode material of hydrogen-nickel battery was performed through a hydrometallurgical process. Most of rare earth elements are separated from nickel and cobalt in the form of sulfates when the anode material is firstly leached with sulfuric acid. Then, the precipitated rare earth sulfates are dissolved with sodium hydroxide to form rare earth hydroxides. The rare earth element, zinc and manganese ions in the lixivium are also separated from nickel and cobalt by using PC-88A extractant system, and the organic phase loaded rare earth is stripped with hydrochloric acid. By neutralizing the stripping solution with rare earth hydroxide, the rare earth chloride is obtained. Under the suitable leaching conditions of sulfuric acid 3 mol/L, leaching time 4 h and temperature 95 ℃, 94.5% of rare earth in the anode material is transformed into the sulfate precipitates, and the leaching ratios of nickel and cobalt can approach 99.5%. When the pH value of the extractive system is controlled in the range of 3.0-3.5, the rare earth elements in the lixivium can be extracted completely into the organic phase, and the stripping recovery of the rare earth can reach 98% in the extraction stage. The total recoveries of rare earth, nickel and cobalt are 98.9%, 98.4% and 98.5%, respectively.     

Recovery of Precious and Base Metals from Waste Printed Circuit Boards Using a Sequential Leaching Procedure

This paper considers the issue of recycling of waste printed circuit boards (WPCBs) containing precious and base metals in appreciable amounts. High-pressure oxidative leaching (HPOL) with dilute sulfuric acid resulted in removal of a significant amount of base metals from a WPCB ash sample obtained by incineration at 800°C. The parameters investigated in the precious metal leaching from WPCB residue after HPOL included the sulfuric acid concentration, thiourea concentration, oxidant concentration, leaching temperature, and leaching time. Recovery of gold, silver, and palladium of 100%, 81%, and 13% from the WPCB residue sample was achieved by thiourea leaching under optimized conditions. The results show that the efficiency of precious metal dissolution from the WPCB sample using thiourea solution depended strongly on the concentration of both thiourea and oxidant.     

低品位硫化铜矿细菌浸出

采用正交实验方法,考察了矿石粒度、矿浆含量、接种量对大冶露天矿石和井下矿石的摇瓶浸出过程的影响.结果表明:粒度对露天矿和井下矿的酸耗影响大,对井下矿的浸出率影响高度显著; 矿浆含量对酸耗影响大,对浸出率则不显著; 在接种的初期,接种量对酸耗与浸出率都有影响,2d后接种量的影响消失; 井下矿浸出的最优工艺条件为:粒度<0.154mm,矿浆含量10%,接种量10%; 露天矿浸出的最优工艺条件为:粒度<0.154mm,矿浆含量20%,接种量7.5%.