Recovery and separation of Fe and Mn from simulated chlorinated vanadium slag by molten salt electrolysis

Tailings from the vanadium extraction process are discarded each year as waste,which contain approximately 30 wt%of Fe.In our previous work,we extracted Fe and Mn from vanadium slag,and Fe and Mn existed in the form of FeCl_(2) and MnCl_(2) after chlorination by NH_(4) Cl to achieve effective and green usage of waste containing Fe and Mn.In this work,square wave voltammetry(SWV)and cyclic voltammetry(CV)were applied to investigate the electrochemical behaviors of Fe^(2+)and Mn^(2+)in Na Cl-KCl melt at 800℃.The reduction processes of Fe^(2+)and Mn^(2+)were found to involve one step.The diffusion coefficients of FeCl_(2) and Mn Cl_(2) in molten salt of eutectic mixtures Na Cl-KCl molten salt were measured.The electrodeposition of Fe and Mn were performed using two electrodes at a constant cell voltage.The Mn/Fe mass ratio of the electrodeposited product in Na Cl-KCl-2.13 wt%FeCl_(2)-1.07 wt%Mn Cl_(2) was 0.0625 at 2.3 V.After the electrolysis of NaCl-KCl-2.13 wt%Fe Cl_(2)-1.07 wt%MnCl_(2) melted at 2.3 V,the electrolysis was again started under 3.0 V and the Mn/Fe mass ratio of the electrodeposited product was 36.4.This process provides a novel method to effectively separate Fe and Mn from simulated chlorinated vanadium slag.

A novel method for extracting vanadium by low temperature sodium roasting from converter vanadium slag

Long-term high temperature in conventional vanadium extraction process would cause particles to be sintered and wrapped, thus reducing extraction efficiency of vanadium. Based on the purpose of directional conversion and process intensification, this work proposed a combination of low temperature sodium roasting and high efficiency selective oxidation leaching in vanadium extraction. The investigation of the reaction mechanism suggested that the structure of vanadium slag was changed by roasting, which also caused the fracture of spinel.The addition of MnO2 promoted the directional oxidation of low-valent vanadium into high valence. It also found that Na2S2O8 could oxidize low-valent vanadium effectively in leaching. The leaching efficiency of vanadium reached 87.74% under the optimum conditions, including a roasting temperature of 650 ℃, a roasting time of 2.0 h, a molar ratio of sodium-to-vanadium of 0.6, a MnO2(roasting additive) dosage of 5 wt% and a Na2S2O8(leaching oxidant) dosage of 5 wt%. This percentage is 7.18% higher than that of direct roasting-andleaching under the same conditions.

Microemulsion leaching of vanadium from sodium-roasted vanadium slag by fusion of leaching and extraction processes

The fusion of the leaching and purification processes was realized by directly using microemulsion as the leaching agent.The bis-(2-ethyhexyl)phosphoric acid(DEHPA)/n-heptane/NaOH microemulsion system was established to directly leach vanadates from sodium-roasted vanadium slag.The effect of the leaching agent on the leaching efficiency was investigated,in addition to the molar ratio of H_(2)O/NaDEHP(W),DEHPA concentration,solid/liquid ratio,stirring time,and leaching temperature.In optimal situations,the vanadium leaching efficiency reaches 79.57%.The X-ray diffraction characterization of the leaching residue and the Raman spectrum of the microemulsion before and after leaching demonstrate the successful entry of vanadates from the sodium-roasted vanadium slag into the microemulsion.The proposed method successfully realizes the leaching and purification of vanadates in one step,thereby greatly reducing production costs and environmental pollution.It also offers a new way to achieve the green recovery of valuable metals from solid resources.

Effect of chemical composition on the element distribution,phase composition and calcification roasting process of vanadium slag

The chemical composition of vanadium slag significantly affects its element distribution and phase composition,which affect the subsequent calcification roasting process and vanadium recovery.In this work,seven kinds of vanadium slags derived from different regions in China were used as the raw materials to study the effects of different components on the vanadium slag’s elements distribution,phase composition,calcification roasting,and leaching rate of major elements using scanning electron microscope,X-ray diffraction analysis,and inductively coupled plasma-optical emission spectroscopy.The results show that the spinel phase is wrapped with silicate phase in all vanadium slag samples.The main elements in the spinel phase are Cr,V,and Ti from the interior to the exterior.The size of spinel phase in low chromium vanadium slag is larger than the other vanadium slags with higher chromium contents.The spinel phase of high-calcium and high-phosphorus vanadium slag is more dispersed.The strongest diffraction peak of vanadium spinel phase in the vanadium slag migrates to a higher diffraction angle,and(Fe_(0.6)Cr_(0.4))_(2)O_(3)is formed after calcification roasting as the chromium content increased.A large amount of Ca_(2)SiO_(4)is produced because excess Ca reacts with Si in high-calcium and high-phosphorus vanadium slag.The vanadium leaching rate reaches 88%in some vanadium slags.The chromium leaching rate is less than 5%in all vanadium slags.The silicon leaching rate of high-calcium and high-phosphorus vanadium slag is much higher than that of the other slags.The leaching rate of manganese is higher than 10%,and the leaching rates of iron and titanium are negligible.

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 vanadium 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 temperature 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 vanadium reaches 97.69%.

Roasting and leaching behaviors of vanadium and chromium in calcification roasting–acid leaching of high-chromium vanadium slag

Calcification roasting–acid leaching of high-chromium vanadium slag(HCVS)was conducted to elucidate the roasting and leaching behaviors of vanadium and chromium.The effects of the purity of Ca O,molar ratio between Ca O and V_2O_5(n(Ca O)/n(V_2O_5)),roasting temperature,holding time,and the heating rate used in the oxidation–calcification processes were investigated.The roasting process and mechanism were analyzed by X-ray diffraction(XRD),scanning electron microscopy(SEM),and thermogravimetry–differential scanning calorimetry(TG–DSC).The results show that most of vanadium reacted with Ca O to generate calcium vanadates and transferred into the leaching liquid,whereas almost all of the chromium remained in the leaching residue in the form of(Fe_(0.6)Cr_(0.4))_2O_3.Variation trends of the vanadium and chromium leaching ratios were always opposite because of the competitive reactions of oxidation and calcification between vanadium and chromium with Ca O.Moreover,Ca O was more likely to combine with vanadium,as further confirmed by thermodynamic analysis.When the HCVS with Ca O added in an n(Ca O)/n(V_2O_5)ratio of 0.5 was roasted in an air atmosphere at a heating rate of 10°C/min from room temperature to 950°C and maintained at this temperature for 60 min,the leaching ratios of vanadium and chromium reached91.14%and 0.49%,respectively;thus,efficient extraction of vanadium from HCVS was achieved and the leaching residue could be used as a new raw material for the extraction of chromium.Furthermore,the oxidation and calcification reactions of the spinel phases occurred at 592and 630°C for n(Ca O)/n(V_2O_5)ratios of 0.5 and 5,respectively.

Extraction of vanadium from high calcium vanadium slag using direct roasting and soda leaching

The extraction of vanadium from high calcium vanadium slag was attempted by direct roasting and soda leaching. The oxidation process of the vanadium slag at different temperatures was investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM), and energy dispersive spectroscopy (EDS). The effects of roasting temperature, roasting time, Na 2 CO 3 concentration, leaching temperature, leaching time, and liquid to solid ratio on the extraction of vanadium were studied. The results showed that olivine phases and spinel phases in the vanadium slag were completely decomposed at 500 and 800°C, respectively. Vanadium-rich phases were formed at above 850°C. The leaching rate of vanadium reached above 90% under the optimum conditions: roasting temperature of 850°C, roasting time of 60 min, Na 2 CO 3 concentration of 160 g/L, leaching temperature of 95°C, leaching time of 150 min, and liquid to solid ratio of 10:1 mL/g. The main impurities were Si and P in the leach liquor.

A novel process for comprehensive utilization of vanadium slag

Traditional processes for treating vanadium slag generate a huge volume of solid residue and a large amount of harmful gas,which cause serious environmental problems.In this study,a new process for the comprehensive utilization of vanadium slag was proposed,wherein zeolite A and a V_2O_5/TiO_2 system were synthesized.The structural properties of the as-synthesized zeolite A and the V_2O_5/TiO_2system were characterized using various experimental techniques,including X-ray diffraction,X-ray fluorescence,scanning electron microscopy,and infrared spectroscopy.The results reveal that zeolite A and the V_2O_5/TiO_2 system are successfully obtained with high purity.The results of gas adsorption measurements indicate that the prepared zeolite A exhibits high selectivity for CO_2 over N_2 and is a candidate material for CO_2 capture from flue-gas streams.

Microwave-assisted leaching of vanadium from vanadium slag with sulfuric acid

Microwave irradiation was used to investigate vanadium slag in place of traditional heating.Factors associated with vanadium extraction ratio were studied following the concentration of leaching agent,oxidant dosage,microwave power,microwave irradiation time and mass ratio of liquid to solid.Results indicated that leaching ratio based on microwave leaching at atmospheric pressure can get to 68.48% under the conditions following sulfuric acid concentration 30%,mass ratio of liquid to solid ratio 2∶1,mass ratio of vanadium slag to manganese dioxide dosage 25∶1.4,and microwave power 800W for 3.5 h at 95 ℃.While,leaching ratio based on microwave leaching under pressure was up to 45.79% under the conditions following sulfuric acid concentration 20%,mass ratio of liquid to solid ratio 2∶1,mass ratio of vanadium slag to manganese dioxide dosage 25∶0.9,microwave power 800W for 10 min.Results showed leaching based on microwave irradiation under pressure can shorten time 94.44% in comparison with that at atmospheric pressure,and the leaching extraction can improve about 20%.

Recovery of vanadium from vanadium slag by composite additive roasting-acid leaching process

To minimize the vanadium content in the vanadium extraction tailings, composite additive roasting with (CaO + MgO) and subsequent acid leaching process was carried out dealing with vanadium-bearing converter slag. The effect of additive with different MgO/(CaO + MgO) molar ratios on the roasting and leaching behaviours of vanadium slag was investigated, and the optimum process conditions were obtained. The results show that in the roasting experiment, under the conditions of roasting temperature of 850 ℃ and roasting time of 2 h, the main kinds of vanadate transformed from Ca_(2)V_(2)O_(7) to Ca_(5)Mg_(4)V_(6)O_(2)4 and then to Mg_(2)V_(2)O_(7) with the increase in the MgO/(CaO + MgO) molar ratio. In the leaching experiment, under the conditions of particle size less than 75 μm, leaching temperature of 50 ℃, pH of 2.5, liquid–solid ratio of 20:1, and MgO/(CaO + MgO) molar ratio of 1:3, the leaching efficiency of vanadium is increased by about 5%, but the substitution of MgO for most or all of CaO will significantly reduce the leaching efficiency of vanadium. Furthermore, the leaching efficiency of impurities (P and Cr) can also be decreased by a composite addictive (CaO + MgO) roasting process. The X-ray diffraction, scanning electron microscopy, energy dispersive spectrometry, and X-ray photoelectron spectroscopy of the original vanadium slag and solid products of both roasting and leaching processes were also evaluated.

Leaching behavior and mineralogical evolution of vanadium released from sodium roasted-acid leaching tailing of vanadium slag

The sodium roasted-acid leaching tailing(SRALT)of vanadium slag with a certain amount of vanadium exhibits potential environmental risk.To investigate the leaching behavior of vanadium from the SRALT,neutral batch leaching tests were performed.The evolution of vanadium concentration,pH,redox potential(Eh),dissolved oxygen,and conductivity as a function of time was measured.Pourbaix diagrams of V–H2O system with different vanadium concentrations were obtained to identify the ionic speciation of vanadium in leachate.X-ray diffraction,X-ray photoelectron spectroscopy,field emission-scanning electron microscopy,and thermogravimetry–differential scanning calorimetry analysis were conducted to investigate the mineralogical evolution of the SRALT during the leaching process.It was found that the major minerals of the original SRALT are titanomagnetite,spinel,olivine,and augite.The valence states of V existing in the original SRALT are V^(3+)and V^(5+).The pH and Eh values of the obtained leachates are 10.00–10.58 and(−43)–(+67)mV,respectively.In this pH and Eh region,the released vanadium is mainly present as HVO_(4)^(2−).The FeOOH and CaCO_(3)would form during the leaching process.The HVO_(4)^(2−)would be mainly adsorbed by the FeOOH and slightly incorporated into the CaCO_(3),resulting in the decline in the vanadium concentration.The vanadium concentration above 27 mg L^(−1)and the dissolved oxygen value below 5.0 mg L^(−1)can be obtained after a short leaching period.As a V(V)-releasing and oxygen-depleting substance,the leaching toxicity of the SRALT should not be ignored.

Effects of Oxides Contents in Vanadium Slag on Corrosion Mechanism of MgO-C Bricks

The influence of V2O3, FeO, TiO2, MnO and MgO in vanadium slag on the corrosion mechanism of MgO-C bricks was studied by stationary immersion tests at vanadium-extracting temperature. Experimental results show that FeO, TiO2, and MnO could enhance the corrosion rate and V2O3 and MgO could decrease it. Microstructure and phase composition of worn samples were investigated by SEM-EDS, revealing the presence of Fe particles, produced by graphite reduction, and (Mg,Fe,Mn)O solid solution at the interface. The formation process of (Mg,Fe,Mn)O solid solution was discussed and the corrosion mechnism of MgO-C bricks was thus proposed.

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。

钒渣焙烧过程钒元素价态变化特征探讨

综合利用光电子能谱及矿物分析仪,研究了钒渣及无盐焙烧阶段不同温度节点下的钒渣熟料的矿物组成及钒元素价态组成。结果表明:钒尖晶石、铁橄榄石和钙铁辉石含量随焙烧温度升高而下降,氧化铁、铁板钛矿、钒酸锰等矿物含量随焙烧温度升高而增加;随焙烧温度升高,钒渣逐步由以V^(3+)氧化为V^(4+)为主的边缘氧化向V^(3+)氧化为V^(4+)与V^(4+)氧化为V 5+共存的内部氧化过渡;钒尖晶石中钒的价态为V^(3+),氧化态钒尖晶石中钒的价态为V^(4+);钒渣及熟料矿物分析结果与钒元素价态分析结果之间一致性较好。

熔融钒渣析晶调控高效提钒机理

基于熔融钒渣在不同冷却制度下导致钠化焙烧-水浸工艺提钒率波动大的问题,本文采用光学显微镜(OM)、X射线衍射(XRD)、扫描电镜(SEM+EDS)及电子探针(EPMA)等表征方法,研究熔融钒渣中钒的析出及钒尖晶石结晶特征对钒提取率的影响,并深入探讨了析晶调控高效提钒机理。结果表明:熔融钒渣的水淬温度对熔融钒渣中钒的析出率、钒尖晶石结晶粒度及钒渣钠化焙烧-水浸过程钒物相转化率及浸出率影响显著。在熔融钒渣降温过程中,1400~900℃为尖晶石快速成核生长期,900℃水淬后,钒基本完成从液相向尖晶石固相中的迁移,结晶相以钒铁尖晶石为主,其中钒呈均匀分散状态;900℃后,尖晶石生长变缓,并以钛铁尖晶石沿钒尖晶石边缘析出为主,尖晶石内部仍呈现钒的“中心聚集效应”。对1400℃、900℃和400℃水淬钒渣(Water-quenched vanadium slag,WQVS)分别进行钠化焙烧-水浸试验,水淬钒渣粒度小于0.15 mm,m(Na_(2)CO_(3))∶m(NaCl)∶m(WQVS)=10∶3∶100,焙烧温度为780℃,焙烧时间为200 min。将焙烧产物磨细至粒度小于0.18 mm后80℃水浸,浸出时间为45 min,钒浸出率分别为75.15%、95.09%和78.78%。这表明将熔融钒渣缓冷至900℃水淬可有效提高钠化焙烧-水浸工艺的钒提取率,并同时提高生产效率。

熔融钒渣钙化及酸浸过程动力学研究

针对现有转炉钒渣钙化工艺中存在流程长和能耗高的弊端,直接利用转炉中熔融钒渣的热量,在分离出半钢后,向熔融钒渣中加入氧化钙,使含钒物相发生钙化转型生成易溶于酸的钒酸钙。由Factsage8.1软件计算得到熔融钒渣可能发生的化学反应的热力学结果。研究过程中采用XRF、ICP、XRD和SEM/EDS等技术进行表征。XRD结果表明,随着CaO添加量的增加,熔融钒渣中逐渐形成CaV_(2)O_(5)、Ca_(7)V_(4)O_(17)、CaVO_(3)和Ca_(3)V_(2)O_(8)。当CaO添加过多后,也会形成Ca_(2)SiO_(4)和CaTiO_(3),这些物相会包裹部分钒,降低钒的浸出率。而后对钙化钒渣进行了酸浸提钒试验,考察了钒渣粒度、浸出温度、酸浓度和液固比对钒浸出率的影响,动力学计算结果可知,酸浸过程的表观活化能为11.56 kJ/mol,表明钙化钒渣的酸浸反应为内扩散控制。

钒钛磁铁矿冶炼渣场周边土壤-鱼塘-河流系统中钒铬分布特征及生态风险评价

为探究冶炼废渣中钒(V)和铬(Cr)对周边环境的影响,采集渣场周边土壤、沉积物、鱼塘底泥、鱼塘水和河水样品,分析V和Cr的浓度、赋存形态,并采用富集因子法、地质累积指数法和潜在生态风险指数法对V和Cr的污染状况和生态风险进行评价。结果表明,研究区土壤中V和Cr平均含量分别约为四川省背景值的2.6和4.0倍,富集程度分别为轻微和中度富集;沉积物中V、Cr平均含量分别约为对照沉积物的5.0和3.2倍,底泥中V、Cr平均含量分别约为对照沉积物的5.5和1.4倍,V和Cr富集程度均为中度和轻微富集;水样表现为V污染,鱼塘水和河水V平均浓度分别约为金沙江水样的17.6和1.8倍。研究区V和Cr主要以残渣态形式存在于土壤中,其占比分别为57.1%和89.4%,沉积物和底泥中V和Cr残渣态占比分别为20.1%和74.9%,各环境介质中V的生物可利用率及迁移率要远高于Cr;结合V和Cr地球化学性质分析,研究区环境介质中V和Cr可能主要以V 5+和Cr 3+形式存在。污染评价结果表明,土壤中V、Cr污染程度分别为轻微和中度污染,沉积物与底泥中V、Cr污染程度分别为中度和轻微污染,V和Cr均对生态系统构成轻微风险。因此,为保护研究区域生态系统和公共健康,应尽快修复受污染土壤和沉积物。

半钢炼钢条件下转炉护炉技术探讨

钒钛磁铁矿经过高炉炼铁得到含钒铁水,转炉炼钢时利用碳与钒的选择性氧化,将铁水中的钒富集后进入炉渣,然后从钒渣中提取钒,是目前国内主流的提钒方法。含钒铁水的冶炼,转炉的成渣条件及渣相与常规铁水冶炼差异较大,对护炉技术提出挑战,因此掌握半钢炼钢条件下转炉护炉技术,在生产过程中才能不影响生产组织,保证一座提钒转炉对应两座半钢冶炼转炉的顺行。本文基于半钢炼钢条件下钒渣与钢渣的成分和岩相结构分析研究的基础上,总结了影响转炉护炉的主要因素,针对性的提出改进措施,在某厂半钢炼钢条件下,转炉护炉取得了较好的冶金效果。

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.

高含碳煤渣硫酸熟化——水浸法回收钒、镍

印度某高含碳含钒煤渣含钒3.57%、镍1.19%,具有较高回收价值,但该渣碳含量较高(14.61%),采用传统直接浸出效果较差。首先采用氧化焙烧对该含钒煤渣进行脱碳预处理,然后采用硫酸熟化—水浸技术提取其中的钒和镍。重点研究了焙烧、硫酸熟化和水浸工艺参数条件对钒、镍的浸出的影响。结果表明,最优焙烧温度和时间分别为800℃、2h;焙砂的最佳浸出粒度为-0.075mm占比90%;最佳熟化工艺参数条件为:酸料比1.5∶1、熟化温度200℃、熟化时间1h;最佳水浸条件为:固液比4∶1、水浸温度95℃、水浸时间2.5h;在上述最佳工艺参数条件下,钒的浸出率可高达96%,镍的浸出率约53%。