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.

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.

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.

Dissolution kinetics and thermodynamic analysis of vanadium trioxide during pressure oxidation

The dissolution kinetics of vanadium trioxide in sulphuric acid-oxygen medium was examined. It was determined that the concentration of sulphuric acid and stirring speed above 800 r min 1 did not significantly affect vanadium extraction. The dissolution rate increased with increasing temperature and oxygen partial pressure, but decreased with increasing particle size. The dissolution kinetics was controlled by the chemical reaction at the surface with the estimated activation energy of 43.46 kJ·mol-1. The leaching mechanism was confirmed by characterizing vanadium trioxide and the dissolution residue using SEM-EDS analysis. Combined with thermodynamic calculation, the pressure leaching of vanadium trioxide in the temperature range (100 to 140 ℃) studied occurs as follows: V2O3 + O2 + H2SO4 = (VO2)2SO4 + H2O.

Kinetics of indium dissolution from marmatite with high indium content in pressure acid leaching

The kinetics of oxygen pressure acid leaching marmatite with high indium content was studied. The effects of particle size, agitation rate, temperature, HSOconcentration, and oxygen partial pressure on leaching rate of indium were investigated. The results show that when the agitation rate is above 600 r-min, its influence on indium leaching rate is insignificant. It is determined that the leaching rates increase with the increase in sulfuric acid concentration, temperature, partial oxygen pressure, and the decrease in particle size. Moreover, the results demonstrate that the process of indium leaching is controlled by interface chemical reaction, with apparent activation energy of 65.7 k J-mol. The apparent reaction orders of sulfuric acid and oxygen partial pressure are determined to be 0.749 and 1.260, respectively. The leaching reaction process follows shrinking unreacted core model. And finally, the kinetics model equation is established for indium.

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。

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。

石煤氧压酸浸液萃钒除铁工艺研究

研究了从石煤氧压酸浸液中萃钒除铁的工艺过程,从萃取和反萃的相比、试剂组成、pH值、澄清时间等方面进行了详细试验。研究表明:浸出液经中和还原处理后,采用10%P204+5%TBP+85%煤油萃取钒时,经六级逆流萃取后萃取率为95%以上;负载有机相用15%H2SO4反萃时,经五级逆流反萃后反萃率可达99%以上。经萃取后,浸出液中的钒可富集到37.gL-1以上,铁可缩减至以0.6.gL-1以下,反萃水相中钒铁质量比大于60,钒铁分离效果较好。

贵州铜仁含钒石煤氧压酸浸萃取提钒研究

系统介绍了含钒石煤氧压酸浸萃取提钒新工艺的研究情况。利用多种检测方法，研究了贵州铜仁石煤矿中钒的矿物学特征，在此基础上，研究考察了影响钒浸取率的各种因素，试验结果表明，在浸出时间3—4h、浸出温度150℃、硫酸用量39％-42％、液固比1．2：1、矿石粒度-0．074mm、Po2 1．2MPa、添加剂用量3％-5％的条件下，经氧压酸浸后，钒的浸出率可达92％以上。浸出液采用酸回收．还原除铁．溶剂萃取．氨水沉钒．热解的工艺流程，得到V：0，含量为99．5％的粉状产品，钒的总回收率达80％以上。

铜阳极泥氧压酸浸预处理工艺研究

研究某铜阳极泥氧压酸浸工艺条件，考察硫酸初始浓度、温度、时间、液固比、压力等对金属浸出的影响。结果表明，当控制浸出硫酸初始浓度130～150g／L、温度155～165℃、时间6～8h、液固比5：1、压力1．0～1．5MPa时，酸浸渣率为55％～60％，碲浸出率为85％～90％，镍、铜浸出率均在99％以上。

转炉钒渣氧压酸浸过程V-Fe-H_2O系的电位—pH图

根据高温水溶液中计算各反应电位、pH及电解质活度因子的经验公式,通过热力学计算,得到氧分压为0.5 MPa、对应质量浓度为0.1mol/kg的离子活度以及温度分别为60、100、150和200℃条件下V-Fe-H2O系的电位—pH图。在pH<2的酸性条件下,V3+、VO2+和VO2+的稳定区一直位于Fe3+和Fe2+的稳定区范围内。随着温度由60℃提高到200℃,可溶性钒铁离子稳定共存区的pH值逐渐降低,氧化还原电位逐渐提高,钒铁尖晶石及氧化物的稳定区变化不大。热力学研究结果表明:采用氧压酸浸工艺能够将转炉渣中的钒充分浸出,但不能通过一步浸出在提钒的同时将钒、铁分离,浸出液还需要进一步净化来生产钒制品。转炉钒渣氧压酸浸的实验结果表明:在浸出温度140℃、氧分压0.5 MPa、粒度0.055～0.075 mm、液固比15-1(体积与质量比)、浸出时间120 min、搅拌速度500 r/min及初酸浓度200 g/L H2SO4的条件下,钒的浸出率为96.87%,铁的浸出率为89.25%,实验结果与热力学计算结果相符合。

石煤提钒传统工艺与氧压酸浸新工艺对比

综述了我国石煤提钒的研究情况和工艺现状,对目前我国石煤提钒传统工艺和改进工艺进行了分析评述,并介绍了氧压直接酸浸出提钒新工艺,提出了我国石煤提钒工艺应有的发展方向。

某云母型含钒石煤氧压与常压酸浸提钒比较

湖北某地云母型含钒石煤中85%以上的钒赋存于云母类矿物中,V3+、V4+分别占总钒的70.83%、29.17%,V3+以类质同象取代云母晶格中的Al3+离子,常压酸浸极难释放出晶格中的钒。为了确定该矿石的高效、低耗、环保浸钒工艺及参数,以常压酸浸效果为参照,对氧压酸浸工艺条件进行了研究。结果表明,浸出温度、硫酸浓度以及氧分压的升高可显著提高钒浸出率,压力场的引入可大幅度提高钒浸出率、缩短反应时间、降低酸耗;在硫酸体积浓度为20%、浸出时间为5 h、反应温度为160℃、氧分压为0.5 MPa情况下的氧压酸浸,钒浸出率可达75.98%,较硫酸体积浓度为20%、浸出时间为5 h、反应温度为98℃情况下的常压酸浸钒浸出率高45.12个百分点。

一种高铁低品位硫化铅锌混合精矿综合利用工艺探讨

目前在工业上唯一能对硫化铅锌混合精矿进行处理的方法只有ISP熔炼法,对于低品位混合精矿而言,采用ISP法是不经济的,如果分别作为铅精矿或锌精矿进行处理时,采用传统的冶炼方法都不能经济地高效利用。本文探讨了高铁低品位硫化铅锌混合精矿氧压酸浸工艺的可行性,提出了低品位硫化铅锌混合精矿一段氧压浸出和两段氧压浸出工艺。

复杂铜铅混合精矿氧压浸出综合回收工艺

呷村铜铅混合精矿中铜、铅矿物主要为黝铜矿和方铅矿,还含有较高的锌、银、砷和锑.本试验针对该矿采用一段氧压浸出综合回收工艺进行处理,通过条件优化实验确定了氧压浸出的操作条件.扩大验证实验表明Cu、Zn的浸出率分别高达98.89%、94.92%,Pb、Ag转化为矾类和硫化物形式留在浸出渣中,铜锌与铅银分离彻底.浸出液中的铜、锌分别通过萃取、电积进行回收.浸出渣中的铅、银通过碳酸盐转化-硅氟酸浸铅-硫脲浸银进行回收.铜萃取率,铅、银浸出率分别为96%、94%和93%.

石煤直接氧压酸浸提钒新工艺

研究石煤氧压直接酸浸提钒新工艺。结果表明,在浸出时间3～4h、浸出温度150℃、液固质量比1.2∶1、矿石粒度-74μm条件下,经两段氧压酸浸后,钒的浸出率可达90%以上。

复杂铜铅锌银混合精矿两段逆流氧压浸出工艺

通过对呷村铜铅锌银多金属混合精矿的矿物分析可知,铅、锌主要赋存于方铅矿、闪锌矿中,而大部分铜、银均为难浸出的黝铜矿所载。采用两段逆流氧压浸出工艺进行处理该矿,条件试验研究得出一、二段最佳浸出条件分别为硫酸浓度150 g/L、液固比3:1、反应温度135℃、氧分压0.75 MPa、浸出时间2.0 h;硫酸浓度80 g/L、液固比3:1、反应温度180℃、氧分压1.0 MPa、浸出时间2.5 h。9轮扩大验证循环浸出试验运行顺利,Cu和Zn的平均两段总浸出率分别为93.23%和99.47%,而杂质元素Fe和As的浸出率仅15.77%和6.9%,元素硫的硫磺转化率为54.26%。Pb和Ag大部分转化为铅矾、铅铁矾和硫化银而留在浸出渣中,铜锌与铅银彻底分离。

高硅低品位氧化锌矿氧压酸浸研究

通过对广西某地低品位高硅氧化锌矿矿物分析可知,矿物中的锌主要以硅锌矿和异极矿的形式存在。采用氧压酸浸技术对该矿进行了处理,条件试验研究得出最佳工艺条件为:矿物粒度0.104mm,硫酸浓度120 g/L,釜内压力1.0MPa,浸出时间90min,反应温度120℃,液固比3∶1。综合性试验研究得出:在最佳工艺条件下,锌浸出率可达97%以上,SiO2浸出率低于0.8%,浸出液产液速率大于880L/(m2.h),浸出矿浆具有良好的过滤性能。