Vanadium recovery from clay vanadium mineral using an acid leaching method

A technique including direct acid leaching, vanadium precipitation with alkaline, sodium hydroxide releaching, impurity removing by adjusting pH value, precipitation vanadium with ammonium chloride, and vanadium pentoxide by roasting steps was proposed according to the characteristic of Xichuan clay vanadium mineral. The factors influencing leaching vanadium such as temperature and the concentration of sulfuric acid were investigated and optimized. The experimental results indicate that the extract ratios of V205 can reach 94% and 92% at a sodium chlorate ratio of 3% and a manganese dioxide ratio of 3%, respectivdy. A completely chemical precipitation method was adopted to decontaminate and enrich the vanadium in the acid leaching solution. The X-ray diffraction （XRD） pattern and the purity analysis of vanadium pentoxide indicate that the purity of final vanadium pentoxide can reach 99% and meet the standard specifications. The total recovery can reach about 75%. The technique has the characteristics of simplicity, less investment, and more environment safety as compared with the traditional salt roasting method.

Leaching of vanadium from stone coal with sulfuric acid

The effects of roasting, mass ratio of H2SO4 to stone coal, leaching temperature, liquid-to-solid ratio, grinding fineness of stone coal, and two-stage counter-current leaching on the vanadium leaching ratio were studied. The results show that the vanadium leaching ratio of roasted stone coal through two-stage counter-current leaching can reach 65.1% at the mass ratio of H2SO4 to stone coal of 20%, leaching temperature of 95℃, leaching time of 30 h, liquid-to-solid ratio of 1.1 mL·g^-1, and grinding fineness of 0.1 mm, which can serve as an experimental basis for the production of vanadium from stone coal.

Calcified roasting-acid leaching process of vanadium from low-grade vanadium-containing stone coal

The low-grade vanadium-containing stone coal used in this experiment was collected from Wuxi Country, Chongqing City, China. The experiment focused on the vanadium recovery from roasted residue through optimizing the process conditions of an effective and environmentally-friendly technology, named calcified roasting-sulfuric acid leaching technology. By single-factor experiments and orthogonal experiments, the effects of roasting temperature, roasting time, sulfuric acid concentration and leaching time on the leaching ratio of vanadium were analyzed. The results showed that the leaching ratio of vanadium reached 85.5% under the proper technological conditions of roasting temperature=950℃, roasting time=4 h, 40% concentration of sulfuric acid and leaching time=6 h.

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.

Acid leaching of vanadium from roasted residue of stone coal

Through leaching from residue directly and leaching after a roasting treatment,respectively,the experimental research on sulfuric leaching of vanadium from residue of stone coal that came from power generation was conducted.Factors which influence the leaching of vanadium such as concentration of sulfuric acid,leaching temperature,leaching time and liquid-to-solid ratio were investigated in both processes.In the process of direct leaching,to achieve a leaching rate of 74.49%,H2SO4 concentration of up to 5.4%,leaching temperature of 90℃and leaching time of 8 h were necessary reaction factors.The results show that after a roasting treatment at the optimum condition of 950℃at 1 h,76.88%vanadium can be leached under the experimental condition of 0.45% of H2SO4,30℃for 1 h with a liquid-to-solid ratio of 2 mL/g.Leaching after an oxidation roasting treatment is an efficient way to leach vanadium from the residue of stone coal,which has some advantages,such as high recovery,low economic cost and less impurities in leaching solution.

Extraction of vanadium from vanadium slag by high pressure oxidative acid leaching

To extract vanadium in an environment friendly manner, this study focuses on the process of leaching vanadium from vanadium slag by high pressure oxidative acid leaching. Characterizations of the raw slag, mineralogy transformation, and the form of leach residues were made by X-ray diffraction, scanning electron microscopy, and energy dispersive X-ray spectroscopy. The result shows that the vana-dium slag is composed of major phases of fayalite, titanomagnetite, and spinel. During the high pressure oxidative acid leaching process, the fayalite and spinel phases are gradually decomposed by sulfuric acid, causing the release of vanadium and iron in the solution. Meanwhile, unreacted silicon and titanium are enriched in the leach residues. With the initial concentration of sulfuric acid at 250 g·L^-1, a leaching tem-perature of 140℃, a leaching time of 50 min, a liquid-solid ratio of 10：1 mL·g^-1, and oxygen pressure at 0.2 MPa, the leaching rate of vana-dium reaches 97.69%.

Solvent extraction of vanadium from sulfuric acid solution

The behaviour of vanadium（V） extracted from sulfuric acid solution was investigated using Cyanex 923 as an extractant. The effects of the concentration of Cyanex 923 and the pH of the solution were studied. The extraction of vanadium（V） increases with the increase of Cyanex 923 concentration and shaking time. Cyanex 923 can extract vanadium（V） from sulfuric acid solution at low pH conditions, and the best pH conditions for extraction of vanadium（V） are at pH 1.0-2.0. The species extracted into the organic phase is VO2HSO4 with one molecule of Cyanex 923. Equilibrium studies were used to assess the extraction efficiency of vanadium（V） recovery from the sulfuric acid solution.

A phosphotungstic acid coupled silica-Nafion composite membrane with significantly enhanced ion selectivity for vanadium redox flow battery

An ultra-high ion-selective Nafion composite membrane modified by phosphotungstic acid(PWA)coupled silica for vanadium redox flow battery(VRB)was constructed and prepared through solution casting.The composite membrane exhibits excellent proton conductivity and vanadium ions blocking property by incorporating the nanohybrid composed of silica and PWA into the Nafion ionomer.Simple tuning for the filling amount of the nanohybrid endows the obtained membranes preeminent vanadium barrier property including a minimum vanadium permeability of 3.13×10-7cm2min-1and a maximum proton conductivity of 0.081 S cm-1at 25°C.These indicate an ion selectivity of 2.59×105S min cm-3,which is 6.8times higher than that of recast Nafion(0.33×105S min cm-3).As a result,the VRB with the composite membrane shows superior battery performance containing a lower self-discharge rate,higher capacity retention and more robust cyclic stability compared with recast Nafion over a range of current densities from 40 to 100 m A cm-2.

Pressure acid leaching of black shale for extraction of vanadium

Extraction of vanadium from black shale was attempted in pressure acid leaching.The chemical components of the sample obtained from Guizhou Province of China show that it contains 3.258%V2O5,52.880%SiO2 and 16.140%Al2O3.Phase analyses of vanadium indicates vanadium mainly exists in the free oxide and mica.Vanadium contents in the two phases are 18%and 53%, respectively.The contents of V3 +,V 4+and V 5+are almost equal.Under the optimum parameters of one-step leaching(reaction time of 3 h,sulfuric addition of 25%,temperature of 150℃,liquid to solid ratio of 1.2 mL/g,catalyst(FeSO4)addition of 5%and size of 85%particle 0.074 mm),about 77%of vanadium is recovered.After two-step countercurrent leaching,the leach recovery of vanadium can reach above 90%.Air replacing oxygen is completely feasible.The impurity metals can dissolve into solution in different degrees.

Study on stabilities and electrochemical behavior of V(V) electrolyte with acid additives for vanadium redox flow battery

Several acid compounds have been employed as additives of the V(V) electrolyte for vanadium redox flow battery(VRB) to improve its stability and electrochemical activity. Stability of the V(V) electrolyte with and without additives was investigated with ex-situ heating/cooling treatment at a wide temperature range of-5 ?C to 60 ?C. It was observed that methanesulfonic acid, boric acid, hydrochloric acid, trifluoroacetic acid,polyacrylic acid, oxalic acid, methacrylic acid and phosphotungstic acid could improve the stability of the V(V) electrolyte at a certain range of temperature. Their electrochemical behaviors in the V(V) electrolyte were further studied by cyclic voltammetry(CV), steady state polarization and electrochemical impedance spectroscopy(EIS). The results showed that the electrochemical activity, including the reversibility of electrode reaction, the diffusivity of V(V) species, the polarization resistance and the flexibility of charge transfer for the V(V) electrolyte with these additives were all improved compared with the pristine solution.

Low temperature catalytic conversion of methane to formic acid by simple vanadium compound with use of H_2O_2

Selective oxidation of methane with hydrogen peroxide was catalyzed by several simple vanadium compounds in CH3CN. The reaction could afford formic acid as the major product. Vanadyl oxysulfate （VOSO4） was found to be an efficient catalyst. Specifically, the selectivity to formic acid of 70% at a methane conversion of 6.5% could be achieved over the VOSO4 catalyst under the reaction conditions of methane pressure 3.0 MPa and temperature 333 K for 4 h. The UV-Vis spectroscopic measurements revealed that the formation of V5＋ species during the reaction might be vital for the methane activation. The reaction probably proceeded via radical mechanism.

Catalytic-kinetic spectrophotometric determination of vanadium (V) based on the Celestine blue-bromate-vanadium (V)-citric acid reaction

A novel sensitive and relatively selective kinetic method is presented for the determination of V（V） based on its catalytic effect on the oxidation reaction of Celestine blue by potassium bromate in the presence of citric acid as an activator. The reaction was monitored spectropho- tometrically by measuring the decrease in absorbance of Celestine blue at a maximum absorption wavelength of 540 nm between 0.5 and 9 min （the fixed-time method） in an H3PO4 medium at 45℃. The effect of various parameters such as concentrations of H3PO4, citric acid, potassium bromate and Celestine blue, ionic strength, reaction temperature and time on the rate of V（V） catalyzed reaction was studied. The method is free from the most interferences, especially from large amounts of V（IV）. The decrease in absorbance is proportional to the concentration of V（V） over the entire concentration range tested （0.025-1.25 lag.mL^-1） with a detection limit of 6.80 tag.L^-1 （according to statistical 3Sblank/k criterion） and a coefficient of variation （CV） of 1.78% （for ten replicate measurements at 95% confidence level）. The proposed method suffers from a few interferences such as Cr（VI） and Hg（Ⅱ） ions. The method was successfully applied to the determination of V（V） in river water, lake water, tap water, natural drinking water samples and a certified standard reference material such as SRM-1640 with satis- factory results. The vanadium contents of natural water samples were detected by using both linear calibration curve and standard addition curve methods. The recoveries of spiked vanadium （V） into the certified standard water sample were found to be quantitative, and the reproducibility was satisfactory. It was observed that the results of the SRM 1640 were in good agreement with the certified value.

镁化焙烧钒渣中钒的酸浸动力学研究

为了促进钒渣镁化焙烧−酸浸提钒法的工业应用,研究钒渣镁化焙烧熟料中钒的酸浸出动力学。结合XRD、XPS和SEM,揭示在浸出初期的10 min内,MgSiO_(3)和大部分钒酸盐急剧地消耗酸,形成SiO2和V(V)的氧钒根离子。通过研究钒浸出效率随浸出时间的变化,建立钒浸出的动力学模型。结果表明,浸出过程分为两个不同的阶段,且两个阶段的动力学均遵循收缩核模型。0~10 min的浸出速率由固体产物层扩散和化学反应共同控制,而10~30 min的浸出速率仅由化学反应控制。因此,浸出初始10 min内加速传质很关键,而后续20 min内精确控温至关重要。

混合酸浸法回收废脱硝催化剂的研究

为提高废脱硝催化剂中钒的浸出率,将草酸和硫酸配制成混合酸,进行了废脱硝催化剂在混合酸下的钒浸出研究,考察了不同温度下混合酸的浓度配比对钒浸出率的影响。结果表明:以0.6 mol/L草酸+3 mol/L硫酸的混合酸作为浸出剂时,废脱硝催化剂中钒的浸出率达81.91%。在优选工艺下得到的浸出渣粉体的主要成分为锐钛矿型TiO_(2),在富集TiO_(2)的同时浸出渣的比表面积提高18.3%,钛钨含量占比为91.8%,可作为制备催化剂的原材料循环使用。

含钒页岩酸浸提钒废水处置的研究进展

含钒页岩酸浸提钒过程中会产生两类废水:酸性废水以及氨氮废水。基于提钒废水的性质特点,评述了其处置的研究进展。处理酸性废水时,中和沉淀法操作简单,但存在重金属离子时还需进行二阶段处理;膜分离法避免了中和沉淀法产生污泥多的情况,但膜易结垢使得寿命较短;酸性废水循环主要是将其返回至前端的浸出过程,利用废水中的H+等,可明显节约成本。处理氨氮废水时,吹脱法简单高效,但废水氨氮浓度较高时还需要进行二次脱氮;磷酸铵镁结晶法不仅操作简便,其产物还可作为肥料使用,但其消耗药剂较多,这增加了成本;离子交换法虽运行成本低、效率高,但尚有吸附容量小、金属负载型树脂离子易洗脱等关键问题待解决;催化氧化法效率高且产物清洁,但其使用成本高以及反应稳定性差。最后,展望了提钒废水处置的方向与趋势。

低品位石煤钒矿两段逆流酸浸提钒的研究

以陕南某低品位石煤钒矿为原料,在添加含氟助浸剂的条件下采用两段逆流硫酸浸出提钒。结果表明,在石煤钒矿磨矿细度为―0.074 mm占40%,助浸剂为1%NaClO_(3)+3%2^(#)含氟助浸剂,硫酸用量为22%,两段逆流浸出温度为90℃,一段浸出时间为1 h,二段浸出时间为2 h的条件下,钒浸出率可以达到94.95%。该流程可有效降低硫酸用量,提高钒浸出率,节省药剂,并能适当简化后续提钒工艺。

高离子选择性PI基两性离子交换膜的制备与应用

两性离子交换膜可以结合质子交换膜和阴离子交换膜的优点,解决质子传输能力与钒离子渗透能力之间的权衡关系。针对许多报道中阴阳离子位于同一侧链而限制了调控其比例的问题,本文以5-氨基-2-(4-氨基苯基)苯并咪,4,4′-二氨基-2,2′-二磺酸基-联苯和4,4′-二氨基二苯砜为二胺单体,以1,4,5,8-萘四甲酸酐为二酐单体,制备了含有苯并咪唑结构主链的磺化聚酰亚胺膜材料,利用苯并咪唑的易修饰性,制备了具有长功能基团侧链的两性聚酰亚胺膜材料,可灵活调控阴阳离子的比例。通过调控长侧链阳离子基团的含量制备了QBISPI-x(x=0,33,67,100)系列膜。QBISPI-0具有最高的质子传导率(27.4622 mS·cm^(−1))和离子选择性(6.598×10^(5)S·min·cm^(−3)),在VRFB的测试中表现出最佳性能,在高电流密度(200 mA·cm^(−2))下能量密度可达到73.6%。

钒铬渣钙化焙烧—酸浸过程多组元浸出行为

以钒铬渣钙化焙烧熟料为原料,探究了浸出pH值、浸出温度、浸出时间和浸出液固比对钒铬渣中钒以及铬、锰、硅、钛、铁组元浸出率(质量分数,下同)的影响.结果表明:在酸浸过程中浸出pH值为2.8、浸出温度为60℃、浸出时间为60 min、浸出液固比为5的条件下,钒浸出率达到87.81%,此时铬、锰、硅、钛、铁浸出率分别为0.44%,23.91%,2.28%,0.23%,0.03%;此外,延长浸出时间、提高浸出温度、适当增大浸出液固比均有利于钒的分离提取.

用空白焙烧—酸浸工艺从石煤钒矿中提取钒

研究了采用空白焙烧—酸浸工艺从某石煤钒矿中提取钒,考察了磨矿细度、焙烧时间、焙烧温度、硫酸用量、助浸剂CaF_(2)用量、浸出温度、浸出时间和液固体积质量比对钒浸出率的影响。结果表明:适宜的空白焙烧条件为磨矿细度-74μm占70%,焙烧温度775℃和焙烧时间5 h;适宜的酸浸条件为硫酸用量10%、助浸剂CaF_(2)用量3%,液固体积质量比1.5∶1,浸出温度25℃和浸出时间1.5 h;适宜条件下,钒浸出率可达81.5%。

P507萃取分离硫酸溶液中钒和铁的动力学研究

钒作为一种稀有金属常与其他金属伴生,含钒矿石的硫酸浸出液中含有大量铁等杂质离子,严重影响钒的后续分离富集。萃取动力学研究是判断萃取过程控制步骤和金属离子萃取分离难易程度的重要方法。本文采用自行设计的圆筒型恒界面池对P507萃取硫酸溶液中V(Ⅳ)和Fe(Ⅱ)的动力学进行了研究。结果表明:水相是萃取过程的扩散阻力区,该萃取过程为化学反应控制类型,随着搅拌速度和萃取温度的升高,V(Ⅳ)和Fe(Ⅱ)的萃取速率均呈上升趋势。萃取V(Ⅳ)的表观活化能为48.68 kJ/mol,萃取Fe(Ⅱ)的表观活化能为135.11 kJ/mol,P507更容易萃取硫酸溶液中V(Ⅳ),有望实现V(Ⅳ)与Fe(Ⅱ)的高效分离。