Extracting vanadium from stone-coal by oxygen pressure acid leaching and solvent extraction

Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction.The mineralogy of the stone-coal from Tongren City of Guizhou Province,China,was investigated by various determination methods. The effects of leaching time,leaching temperature,leaching agent concentration,leaching L/S ratio,granularity of material,additive consumption were investigated based on the mineralogy.The results show that under the conditions of leaching time of 3-4 h, temperature of 150℃,sulfuric acid consumption of 25%?30%,ratio of liquid to solid of 1.2:1,the granularity less than 0.074 mm, additive consumption of 3%-5%,and oxygen pressure of 1.2 MPa,and the vanadium leaching rate can be more than 92%by the method of two-step pressurized acid leaching.The powdery V2O5 product with 99.52%in V2O5 content is obtained by the flowsheet of acid recovery,removing iron by reduction process,solvent extraction,precipitating vanadium with ammonium water,and pyrolysis from the stone-coal oxygen pressure acid-leaching solution.The total recovery efficiency of vanadium is above 85%,which is more than 20%higher than that obtained in the conventional process.Furthermore,the new process does not cause air pollution since no HCl or Cl2 is released by calcination of the raw material.

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%.

Analysis of Oxygen Consumption in Lead and Zinc Metallurgy

This article analyzes the role of oxygen in lead zinc metallurgy,including shortening the metallurgical process,promoting energy conservation and environmental protection,improving metallurgical strength,enhancing raw material adaptability,and enhancing comprehensive recovery efficiency.This article introduces different lead zinc metallurgical processes and their oxygen consumption characteristics,including oxygen enriched side blowing lead smelting,oxygen bottom blowing lead smelting,oxygen enriched top blowing lead smelting,flash smelting lead,oxygen pressure leaching zinc smelting,and atmospheric pressure oxygen leaching zinc smelting.It is pointed out that oxygen enhanced metallurgy is the direction for the transformation and upgrading of lead zinc metallurgy.

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.

An investigation of oxygen pressure acid leaching of Gacun complex Cu-Pb bulk concentrate

The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The pri-mary copper and leadminerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in themineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.% and 94.9 wt.%,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.

Pressure acid leaching of zinc sulfide concentrate

Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indicate that the additive can improve the reaction kinetics and the conversion rate.And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur.The reasonable experiment parameters are obtained as follows:the leaching temperature 150℃,oxygen partial pressure 1 MPa,additive amount 1%,solid-to-liquid ratio 1:4,leaching time 2 h,initial sulfuric acid concentration 15%,and particle size less than 44μm.Under the optimum conditions,the leaching rate of the zinc can reach 95%and the reduction rate of the sulfur can reach 90%.

Oxygen pressure acid leaching of Gacun complex Cu concentrates

The treatment of the Gacun complex Cu concentrate with high contents of Pb,Zn,Ag,etc by oxygen pressure acid leaching was studied.It is unusual that tetrahedrite,whose treatment was rarely studied,is the primary copper mineral of the concentrates.Most of silver also occurs in the mineral.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Pilot scale test was carried out under the parameters,and the leaching rates of copper and zinc are as high as 97.10% and 89.83% while lead and silver are transformed into sulfate and sulfide respectively and stay in leaching residue.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by chloride leaching and thiourea leaching.The extraction rate of copper achieves 96%,and the leaching rates of lead and silver reach 90% and 95%,respectively.

熟化银精矿氧压强化酸浸行为研究

考虑到浮选银精矿极具回收价值及现有处理工艺存在不足,本文提出了浓硫酸熟化-氧压酸浸的工艺路线。首先,以湿法炼锌过程产出的浮选银精矿为对象,利用浓硫酸的强氧化性破坏浮选残留有机药剂矿物包裹体;其次,利用加压湿法冶金独特的技术优势,氧化溶解低价金属硫化物;最后,重点研究了工艺过程原理及氧压酸浸过程中Zn、Cu、In的浸出行为及浸出渣中物相演变规律。研究结果表明:银精矿颗粒表面残留浮选药剂的P—O、—OH极性基团被浓硫酸氧化破坏,避免产生水化膜及矿物包裹体,促进有价元素的高效浸出;在高温氧化性条件和Fe^(3+)/Fe^(2+)变价离子对的助浸作用下,银精矿物相组成由闪锌矿(Zn S)、锌铁尖晶石(Zn Fe_(2)O_(4))、雌黄铁矿(FeS)、氯化亚铜(Cu Cl)为主,转变为以单质硫(S8)、铁矾(MeFe_(3)(SO_(4))_(2)(OH)_(6),Me:1/2Pb、Na、K)为主,并残留少量的闪锌矿(Zn S);适当提高反应温度和初始质量浓度可促进Zn、Cu、In的浸出,但温度过高会加剧铟铁共沉淀问题,降低锌、铟的浸出率,而质量浓度过高会降低氧气对Zn S及Fe^(2+)的氧化能力,抑制银精矿的溶解;延长反应时间、提高液固比均有利于银精矿的解离浸出;增大氧分压可以加强Fe^(3+)的氧化能力,促进对Zn S及ZnxFe(1-x)S的氧化作用,有利于银精矿的溶解。在反应温度为155℃、初始质量浓度为160 g/L、反应时间为180 min、氧分压为0.8 MPa、液固比为7 m L/g的氧压酸浸条件下,银精矿中Zn、Cu、In的浸出率分别为94.84%、98.36%和91.13%,浸出渣中铅、银的富集率分别达2.88%和7754 g/t。

含锗银精矿浓硫酸熟化—氧压酸浸行为研究

针对银精矿极具回收价值及现有处理工艺不足,以湿法炼锌过程产出的含锗银精矿为研究对象,提出了浓硫酸熟化—氧压酸浸的工艺路线,重点研究了工艺过程原理、含锗银精矿熟化—氧压酸浸过程中Zn、Cu、Ge的浸出行为及浸出渣中物相演变规律。结果表明:浓硫酸熟化预处理可氧化分解含锗银精矿中低价金属硫化物及精矿表面残留的有机药剂;在高温氧化性条件和Fe^(3+)/Fe^(2+)变价离子对的助浸作用下,物相组成由含锗银精矿中的铁闪锌矿、闪锌矿、雌黄铁矿、氯化亚铜为主演变为浸出渣中单质硫、硫酸铅、铁矾(MeFe_(3)(SO_(4))_(2)(OH)_(6),Me:1/2Pb、Na、K)为主。最佳氧压酸浸条件为:反应温度150℃、初始酸度150 g/L、反应时间180 min、氧分压0.6 MPa、液固比7 mL/g,Zn、Cu、Ge的浸出率分别为95.34%、98.47%和84.35%,铅、银分别富集至3.19%和7074 g/t,实现了Zn、Cu、Ge的深度浸出和铅、银的高效富集。

硫化锌精矿氧压浸出过程中硫回收率的影响机制

锌氧压浸出特点是硫化锌精矿直接湿法浸出,硫不进入烟气,而是直接转化成单质硫回收。该工艺生产流程更简单,环境友好,综合能耗低,硫磺产品相对更容易储存、运输及销售。因此该技术的推广应用对锌冶炼行业节能减排乃至碳达峰碳中和目标的实现意义重大。从精矿矿物特征及金属的浸出率、浸出过程中硫元素的酸化率、硫磺结晶过程控制、硫浮选效率以及硫精矿热过滤温度等方面着手,研究其对硫磺回收率的影响机制。

铜冶炼烟尘加压浸出及物相演变行为

铜冶炼烟尘是火法炼铜过程产出的高附加值二次资源,为实现其中铜、锌等有价金属的浸出及砷的分离,本文研究铜烟尘加压浸出及同步沉砷新工艺技术参数优化与过程物相演变规律。结果表明:在反应温度由140℃升高至160℃过程中,锌、铜浸出率分别可达96%和97%,砷沉淀率由36%提高至92%,砷沉淀形态由非晶形砷酸铁(FeAsO_(4)·n H_(2)O,n≠2)向晶型砷酸铁(FeAsO_(4)·2H_(2)O)演变;整个反应过程显著分为以锌、铜高效浸出(0~60 min)和砷铁共沉淀(60~180 min)两个阶段;增大搅拌转速、提高液固比可进一步提升体系的氧化传质速率和砷酸铁矿化物相-臭葱石生成速率。在反应温度160℃、时间3 h、初始酸度40 g/L、液固比5∶1、搅拌转速600 r/min的优化技术参数条件下,锌、铜浸出率均可达96%以上,砷沉淀率可达92%,浸出渣主要物相为硫酸铅与臭葱石,实现了锌、铜的深度浸出和砷的高效分离。

富钴铜冶炼渣预浸—氧压浸出工艺研究

铜冶炼过程中产生的冶炼渣含有较多的铜、钴等有价金属,从冶炼渣中回收这些有价金属具有重要经济价值和环保意义。以Cu含量8.26%、Co含量1.52%的富钴铜冶炼渣为原料,采用预浸——氧压浸出工艺对其进行处理。系统考察了酸矿比、反应温度、反应时间、氧分压、液固比对Co、Cu、Fe浸出率浸出效果的影响,得出最佳工艺条件为:液固比3、反应温度230℃、反应时间1.5 h、酸矿比350 kg t、磨矿细度—0.074 mm占75%、氧分压0.2 MPa,在该条件下,Co、Cu、Fe浸出率分别达到98.38%、95.34%和2.07%。相较于常压浸出,该工艺能有效降低酸耗和浸液中铁离子浓度。

锌氧压酸浸硫精矿和硫尾矿的矿物学特征与环境活性

采用电感耦合(ICP-OES)、X射线衍射(XRD)、扫面电镜-能谱(SEM-EDS)、红外光谱(FTIR)及差热差重(TG-DTA)等多种检测方法对锌氧压酸浸硫精矿和硫尾矿的化学组成、粒度分布、物相结构等进行研究,并采用毒性浸出程序(TCLP)法考察其短期环境活性。研究结果表明:硫精矿1和硫精矿2主要以球状的单质硫存在,其中,少量硫精矿1以小颗粒的闪锌矿、黄铁矿等附着于块状的铁矾颗粒上,平均粒径为22.1μm;少量硫精矿2以块状的闪锌矿、黄铁矿和小颗粒存在,平均粒径为59.2μm。2种硫精矿在185~340℃时会剧烈燃烧,热稳定性较差。2种硫尾矿主要以块状的铅黄铁矾和黄钠铁矾存在,其中,硫尾矿1平均粒径为21.1μm;硫尾矿2含一定的赤铁矿(16.86%),平均粒径为5.3μm。硫精矿和硫尾矿中锌和镉元素在弱酸条件下不稳定,硫尾矿1中锌和镉的浸出量分别超标1.93倍和2.50倍;硫精矿2的分别超标2.03倍和6.56倍,硫尾矿1的分别超标3.46倍和5.94倍,硫尾矿2的分别超标11.10倍和8.28倍。

锌精矿氧压浸出技术研究进展

传统湿法炼锌工艺存在烟气处理成本高、综合回收率低等问题,近年来,氧压浸出技术因其原料适应性强、锌回收率高、环境污染小等优点被应用于锌冶炼工业。介绍了锌精矿氧压浸出的生产工艺,论述了浸出过程反应的基本原理,以及离子催化、温度、酸度、分散剂等条件对浸出过程离子行为的影响与机理。讨论了锌精矿中伴生元素铁、镓、锗、铟、银等的浸出行为及在不同浸出条件下浸出行为的变化。

加压湿法炼锌过程中镓的富集

针对冶炼厂在两段逆流氧压浸出湿法炼锌工艺流程中镓回收率不高的问题,采用XRD、SEM-EDS与ICP-AES对工艺流程中各阶段溶液及渣物相进行分析,对预中和、锌粉置换过程影响镓回收率低的机理进行探究。结果表明,一段浸出液预中和过程中由于酸度过低导致镓与黄钾铁矾或铅黄钾铁矾共沉淀的现象是导致镓损失的原因。在锌粉置换过程中,单纯依靠锌粉提高pH无法使镓完全沉淀。据此提出在预中和过程中通过将预中和终点酸度控制为6 g/L,反应时间1 h,反应温度85℃,以避免镓损失;在置换阶段采用先中和再置换的方法:先使用优质氧化锌7 g/L进行中和反应1 h,然后加入锌粉3 g/L反应1 h,反应温度85℃。经优化后,镓沉淀率可达100%。

含砷冶炼烟尘选择性脱砷与铜锌回收研究

含砷冶炼烟尘是有色金属冶炼中产生的一类重要二次资源,但砷具有剧毒性,限制其清洁化利用。根据含砷烟尘中砷物相特征,采用水浸-氧化酸浸梯次脱砷,利用溶剂萃取回收铜锌,实现烟尘资源的综合利用。对于易溶相As_(2)O_(3),在液固比4∶1、温度95℃和时间3 h的条件下进行水浸,砷浸出率为52.85%,水浸液在120℃温度下蒸发结晶回收砷。对于难溶相Pb5(AsO4)3Cl,在温度45℃、液固比3∶1、硫酸浓度1.5 mol/L、H_(2)O_(2)浓度3.0%和反应时间4 h的条件下进行氧化酸浸,砷、铜和锌的浸出率分别为99.16%、99.67%和98.64%,酸浸渣可作为炼铅原料。浸出液中,铜和锌分别采用萃取剂N902和P204进行回收,其萃取率分别为96.42%和96.38%。含砷冶炼烟尘选择性脱砷和铜锌萃取组合工艺具有脱砷效率高、成本低、安全环保等优点,适用于含砷物料资源化清洁利用。

草食性动物排泄物制备氧压酸浸炼锌中分散剂的研究与应用

对木质素在牧草中的分布进行了详细论述,对牧草的体外消化情况进行了分析,归纳了木质素的制备方法,创新地提出了草食性动物排泄物富集—提纯木质素—制备氧压浸出分散剂的技术新思路,利用草食性动物消化—分解—吸收植物中蛋白质、脂肪、淀粉和粗纤维等营养物质,使得木质素残留在排泄物中,实现了木质素的天然富集,并以草食性动物排泄物为原料制备了木质素磺酸盐,提高了木质素磺酸盐单位产出量,降低了处理成本。该技术结合了冶炼企业与草原生态的特点,充分利用当地草原牧草、草食性动物排泄物为锌氧压浸出提供制备木质素的原材料,形成了植物-动物-锌冶炼多领域相结合的特色工艺。

锌冶炼渣中有价金属回收工艺研究

回收利用锌冶炼渣中有价金属,对冶金行业可持续发展具有重要意义。某锌冶炼渣中锌、铅、金、银含量较高,试验采用酸浸—碱浸—氰化浸出湿法梯级浸出工艺回收锌、铅、金、银。试验结果表明:在硫酸质量分数20%,液固比2∶1,浸出温度80℃,浸出时间2 h的条件下,锌浸出率为90.31%;在氢氧化钠质量分数10%,液固比2∶1,浸出温度80℃,浸出时间2 h的条件下,铅浸出率为93.37%;在氰化钠质量分数0.20%,液固比2∶1,浸出时间16 h的条件下,金、银浸出率分别为82.61%、92.39%。该湿法梯级浸出工艺实现了锌冶炼渣的综合回收。

电解废液与锌冶炼烟尘协同处理试验研究

为综合回收锌冶炼烟尘及电解废液中的有价金属,本研究进行了利用锌电解废液浸出锌冶炼烟尘试验,通过一段浸出、碳酸钠中和沉锌及氢氧化钠中和沉镁等工序获得了沉锌产品和沉镁产品,通过两段浸出、萃取反萃、中和沉铟等工序获得了富铟渣及副产物铅银渣。试验最佳工艺条件:一段浸出为液固比4:1,浸出温度80℃,浸出时间75 min,在此条件下,锌浸出率可达78.69%,铟浸出率仅为8.4%;锌镁分离最佳终点pH值区间为6.86~7.80;二段浸出最佳工艺条件为终点pH值1.08,液固比3:1,浸出温度75℃,浸出时间10 h,在此条件下,铟浸出率可达86.84%。该研究可为炼锌厂开路除杂及综合回收有价金属提供新思路。

锌冶炼厂铅锌渣铅锌分离工艺研究

某锌冶炼厂铅锌渣的铅含量为43.89%,锌含量为14.29%,其中锌未得到有效回收,造成资源浪费。试验通过“焙烧+酸浸”的方式探索铅锌渣中锌回收的最佳条件。结果表明,铅锌渣在600℃温度下焙烧40min后,在浸出温度85℃、浸出时间2h、终酸浓度25g/L和钙硅比1.5的条件下,经浸出,铅锌渣的锌品位降低到小于1%的水平,锌回收率超过95%,浸出渣满足铅锌渣售卖要求,有效实现铅锌分离和铅锌资源的高效回收利用。