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%。     

钒钛磁铁矿提钒尾渣浸取钒

采用硫酸氢氟酸次氯酸钠组合浸出体系浸取钒钛磁铁矿提钒尾渣中的钒,研究浸出过程中试剂浓度、浸出液固比、浸出温度、浸出时间、物料粒度对钒浸出率的影响。结果表明:钒的浸出率随试剂浓度、液固比、温度和时间的升高而增大;当矿物粒度小于0.20 mm时,钒浸出率有随矿物粒度变小而减小的趋势。在物料粒度0.15~0.25 mm、初始硫酸浓度150 g/L、初始氢氟酸浓度30 g/L、次氯酸钠加入量为矿量1.5%、矿浆液固比6:1、浸出温度90℃、浸出时间6 h、搅拌速度500 r/min的条件下,钒的浸出率可达85%以上。     

机械活化强化钒渣钙化提钒工艺

针对转炉钒渣钙化焙烧酸浸工艺中存在的钒转浸率低的问题,采用高能球磨对钒渣进行活化预处理,以期强化其提钒效果。采用激光粒度分析仪、BET比表面积测定仪和XRD对活化前后钒渣进行了粒度、比表面积及物相结构分析;采用浸出实验研究了机械活化对钙化焙烧和浸出的影响规律。结果表明:机械活化法增大了钒渣的比表面积,增加了晶格畸变与微观应力,使含钒物相充分解离,由此可改善钒渣钙化焙烧的动力学条件。在浸出20 min条件下,机械活化80 min可将钒浸出率提高10%,最佳焙烧温度降低100℃。     

石煤湿法强化提钒新工艺

分别考察石煤常压氧化酸浸和加压氧化酸浸过程中的钒浸取率.结果表明:采用石煤加压氧化酸浸强化提钒新工艺钒浸出率接近90%,比传统工艺提高15%以上,比现有常压浸出工艺的钒浸出率提高20%,特别是当加入催化剂R时,加压氧化酸浸工艺的优势更为突出.提出石煤加压氧化酸浸新工艺的较佳工艺条件:硫酸酸度约为250 g/L,时间约为4 h,加压釜内压强约为1.2 MPa, 浸出温度为100～120 ℃,催化剂R的用量约为石煤矿量的2.5%.     

石煤提钒低温硫酸化焙烧矿物分解工艺

针对石煤提钒常压硫酸浸出能耗高、作业周期长的缺陷,提出石煤低温硫酸化焙烧矿物分解新工艺.以贵州凯里石煤为原料,对石煤低温硫酸化焙烧的时间、焙烧温度、硫酸加入量以及焙砂水浸工艺参数进行研究.结果表明:先对石煤进行低温硫酸化焙烧处理,再将焙砂按液固比1.2 mL/g加水于100 ℃下搅拌浸出2 h,钒的浸出率可达78.2%;而在相同酸矿比和固液比的条件下,采用常压直接酸浸石煤时,在100 ℃下搅拌浸出48 h后,钒的浸出率只有67.8%.石煤通过低温硫酸化焙烧可有效强化矿物分解过程,缩短提钒作业周期,提高酸的利用率及钒的浸出率.     

转炉提钒工艺对钒渣质量的影响

分析了钒渣中影响后道工序提取V2O5的相关因素,并研究转炉提钒的特点对钒渣质量的影响,提出了相关的工艺改进思路。     

钒掺杂对铝合金微弧氧化层结构和性能影响

通过在电解液中添加NH4VO3制备了钒掺杂铝合金微弧氧化层,研究了不同添加浓度对铝合金微弧氧化层结构和性能的影响。利用扫描电镜(SEM)观察微弧氧化层表面形貌,能谱(EDS)仪分析了膜层V、O元素含量,XPS测定V、O元素的价态,X射线衍射(XRD)仪分析了相组成,极化曲线评定了耐蚀性。结果表明,微弧放电区温度高于1714.38K时?3VO开始转变形成V2O5,低熔点的V2O5在电弧作用下优先熔化而抑制了微弧氧化层表面多孔层的形成。钒掺杂对微弧氧化层相组成影响较小,有利于提高膜层的厚度和耐蚀性。     

低钒转炉钢渣提钒湿法工艺的动力学研究

为了提高湿法浸出低钒钢渣中钒的浸出效率,并对湿法浸出低钒钢渣中钒提供理论依据,从动力学角度分析整个浸出过程。考察温度、液固比、硫酸质量分数和搅拌速率对浸出过程的影响。研究结果表明:在90℃、液固比为10?1以及硫酸浓度6.0mol/L时,浸取9h,低钒钢渣中钒的浸出率可达到95.3%。通过正交实验和动力学推导,得到描述浸出过程的经验方程,低钒钢渣湿法浸出钒的动力学模型为收缩核动力学模型,浸出过程的表观活化能为12.794kJ/mol,该模型表明浸出过程中的控制步骤取决于固膜扩散速率。提高温度、液固比和硫酸质量分数,均可加速钒的浸出速度,提高钒的浸出率。     

KOH亚熔盐中钒渣的溶出行为

对钒渣在KOH亚熔盐体系中的分解动力学进行研究,考察反应温度、碱矿质量比、粒度、气流量等工艺参数对钒渣分解过程的影响,获得最优工艺参数,并对反应机理进行探讨。结果表明,反应温度是最重要的影响因素;钒渣最优浸出条件如下:在反应温度为180℃,碱矿比4:1,KOH碱浓度75%,搅拌速率700 r/min,反应时间300 min,常压通氧气流量为1 L/min的反应条件下,最终钒、铬的浸出率分别达到95%和90%以上。钒渣在KOH亚熔盐介质中氧化分解遵循缩核模型,并主要受内扩散控制,钒和铬分解的表观活化能分别为40.54和50.27 kJ/mol,钒铬尖晶石的氧化以铁橄榄石、石英相的氧化分解为前提。     

粗TiCl4铜丝塔除钒废水沉淀泥浆综合回收工艺

介绍了一种粗TiCl4铜丝塔除钒废水沉淀泥浆综合回收新工艺.该工艺由沉淀泥浆自氧化、碱洗脱氯、脱氯渣一次酸浸生产硫酸铜、一次酸浸渣苏打焙烧提钒和提钒渣二次酸浸5个主要工序组成.实验结果表明,粗TiCl4铜丝塔除钒废水沉淀泥浆在空气中能自氧化.沉淀泥浆在空气中堆放1个月,接近90%的金属铜变成CuCl2·2H2O,Cu2Cl(OH)3和Cu2(OH)3Cl;这些铜的氯氧化合物在碱性溶液中容易转化成Cu(OH)2;在控制液固比4-1,pH值为 11,温度为80 ℃的条件下搅拌1 h,转化率达96%.当酸浸液的pH值为2.0~2.5时,Fe、V、Ti等杂质留在渣中,浸出液蒸发浓缩至密度为1.38 g/cm3,冷却结晶得到的硫酸铜产品符合国标GB437-93的质量要求.酸浸渣按化学计量的2.5倍加苏打后在700 ℃焙烧3 h,焙烧后按液固比3-1加水在70 ℃搅拌1 h浸钒,水浸液按化学计量的3倍加氯化铵沉淀偏钒酸铵,偏钒酸铵在550 ℃热解2 h得到纯度为98.61%的V2O5.提钒渣再经二次酸浸.整个工艺过程铜和钒的总回收率分别达到98.63%和95.65%.     

Distribution behavior of vanadium and phosphorus between slag and molten steel

The correlation of the equilibrium behaviors of phosphorus and vanadium between slag and low carbon molten steel in inert atmosphere was investigated with respect to the experimental variables of slag basicity, the (P₂O₅) and (V₂O₅) content, and the reaction temperature. The distribution ratios of phosphorus and vanadium increased with an increase in the slag basicity. The logarithms of the vanadium distribution ratio were greater by a factor of about two than those of phosphorus in the range of low slag basicity, but the difference diminished with an increase in the slag basicity. The logarithms of the vanadium distribution ratio increased linearly with an increase in the (P₂O₅) and (V₂O₅) content in the slag, while those of phosphorus remained nearly constant. The logarithms of the phosphorus and vanadium distribution ratio decreased with an increase in temperature, and the dependence on temperature was greater for the phosphorus than for the vanadium. For both the maximization of the vanadium yield and the minimization of the rephosphorization of molten steel in the steelmaking process, the ratio of N_(V2O5)/N_(P2O5), the slag basicity, the ratio of f_[P]/f_[V], and the temperature should be maximized, and the (FeO) content in the slag should be minimized.     

Appropriate titanium slag composition during smelting of vanadium titanomagnetite metallized pellets

The suitable titanium slag composition with high titanium content for electric furnace smelting of vanadium titanomagnetite was investigated through thermodynamics and related phase diagram analysis. According to the thermodynamic results, low-melting-point regions and MgTi₂O₅ primary phase area in the phase diagrams, the suggested titanium slag composition for the present vanadium titanomagnetite metallized pellets should consist of 50% TiO₂, 8%–12% MgO and 13% Al₂O₃ (mass fraction) with a binary basicity of 0.8–1.2. Finally, the verified smelting experiments were conducted and successful separation of the molten iron from the titanium slag is obtained. The obtained vanadium-containing molten iron contains 0.681% V and 0.267% Ti, and the obtained titanium slag contains 52.21% TiO₂ (mass fraction), in which MgTi₂O₅ is the primary phase. The titanium resource in the final titanium slag production could be used to produce TiO₂ pigment by acid leaching methods.     

Recovery of vanadium from vanadium slag by composite roasting with CaO/MgO and leaching

A novel process of composite roasting with CaO/MgO and subsequent acid leaching was proposed to improve the recovery rate of vanadium from Linz–Donawiz (LD) converter vanadium slag. The effects of the MgO/(CaO+MgO) molar ratio and the roasting and leaching parameters on the recovery of vanadium were studied. The results showed that the leaching efficiency of vanadium decreased from 88% to 81% when CaO was replaced completely by MgO; however, it could be improved by roasting with the composite of CaO/MgO. The maximum vanadium leaching efficiency of 94% was achieved under the optimum MgO/(CaO+MgO) mole ratio of 0.5:1. The results from X-ray diffractometry (XRD) and scanning electron microscopy with energy-dispersive X-ray spectroscopy (SEM−EDS) confirm that the formation rate of acid-soluble vanadates can be enhanced during roasting with the composite of CaO/MgO and that the leaching kinetics can be accelerated owing to the suppression of calcium sulfate precipitation.     

One-step electrochemical preparation of metallic vanadium from sodium metavanadate in molten chlorides

Conventionally, metallic vanadium is produced from vanadium pentoxide (V₂O₅). Sodium metavanadate (NaVO₃) is an essential intermediate product for the V₂O₅ production. A novel environmentally friendly route for the metallic vanadium preparation from NaVO₃ by molten salt electrolysis is proposed. In the NaCl molten salt, NaVO₃ has a high solubility of 8.6 mol% (16.3 wt.%) at 1173 K and was easily electro-reduced to V₂O₃. However, V₂O₃ is almost insoluble in the NaCl molten salt, and thus can't be further electro-reduced to metallic vanadium. In the eutectic CaCl₂–NaCl molten salt, NaVO₃ was converted to CaV₂O₆. CaV₂O₄ as the intermediate product was produced during CaV₂O₆ electrolysis and can be dissolved into the eutectic CaCl₂–NaCl molten salt. The dissolved CaV₂O₄ was further electro-reduced to metallic vanadium.     

Study of the oxidation kinetics of vanadium carbide

The oxidation of an oxycarbide of vanadium, VO_0.6C_0.7, and of a vanadium carbide, VC_0.98, was studied athermally up to temperatures of 800° C and isothermally between 400 and 580° C at oxygen pressures ranging from 10^–2 to 1 atm. The oxycarbide followed the parabolic rate law below 450° C with V₂O₅ forming as the only reaction product. The activation energy was 49 kcal/mole. VC_0.98 did not form an oxide in this temperature range, but rather dissolved oxygen, the activation energy being 26.6 kcal/mole. No oxygen pressure dependence on the kinetics was found for either sample in this temperature range. Both samples followed the cubic rate law during oxidation in the range of 500–580° C during which V₂O₅ formed. There was a P^1/3 dependence and the activation energy was the same for both materials, 51 kcal/mole. The cubic rate law and the positive pressure dependency (rather than an anticipated negative dependency) were attributed to an electric field associated with oxygen ions chemisorbed on a thin layer of V₂O₅.     

Corrosion protection properties of vanadium films formed on zinc surfaces

Vanadium films were prepared on zinc surfaces by using a solution containing vanadate. Corrosion protection properties of vanadium-treated (V-treated), chromium-treated (Cr-treated), and untreated zinc surfaces in contact with a 3.5 wt.% NaCl solution were studied using potentiodynamic polarization, electrochemical impedance spectroscopy (EIS), and neutral salt spray (NSS) tests. According to these results, the V-treated layer significantly improved the corrosion resistance of zinc surfaces. In comparison with the Cr-treated layer, the V-treated layer exhibited a better corrosion resistance. The composition of the V-treated layer was studied using X-ray photoelectron spectroscopy (XPS). XPS measurements indicated that the vanadium layer formed on zinc surfaces and the vanadium-rich coating was a hydrated oxide with a composition of V₂O₅, VO₂, and its hydrates such as V₂O_5·nH₂O and VO(OH)₂.     

Cyclic metallurgical process for extracting V and Cr from vanadium slag: Part I. Separation and recovery of V from chromium-containing vanadate solution

The separation and recovery of V from chromium-containing vanadate solution were investigated by a cyclic metallurgical process including selective precipitation of vanadium, vanadium leaching and preparation of vanadium pentoxide. By adding Ca(OH)₂ and ball milling, not only the V in the solution can be selectively precipitated, but also the leaching kinetics of the precipitate is significantly improved. The precipitation efficiency of V is 99.59% by adding Ca(OH)₂ according to Ca/V molar ratio of 1.75:1 into chromium-containing vanadate solution and ball milling for 60 min at room temperature, while the content of Cr in the precipitate is 0.04%. The leaching rate of V reaches 99.35% by adding NaHCO₃ into water according to NaHCO₃/V molar ratio of 2.74:1 to leach V from the precipitate with L/S ratio of 4:1 mL/g and stirring for 60 min at room temperature. The crystals of NH₄VO₃ are obtained by adjusting the leaching solution pH to be 8.0 with CO₂ and then adding NH₄HCO₃ according to NH₄HCO₃/NaVO₃ molar ratio of 1:1 and stirring for 8 h at room temperature. After filtration, the crystallized solution containing ammonia is reused to leach the precipitate of calcium vanadates, and the leaching efficiency of V is >99% after stirring for 1 h at room temperature. Finally, the product of V₂O₅ with purity of 99.6% is obtained by calcining the crystals at 560 °C for 2 h.     

氟化物熔盐中快速电脱氧制备金属钒及其机理

以偏钒酸铵为原料,采用"煤气还原+原位烧结"工艺制备高活性V_2O_3阴极片,在氟化物体系熔盐中实现了快速电脱氧制备金属钒,并通过测定循环伏安曲线结合恒电位电解实验,研究了电解过程的反应机理。结果表明:V_2O_3在氟化物熔盐中可实现快速电脱氧,电解4 h后所得金属钒的氧含量降至0.218%(质量分数,下同);V_2O_3阴极电脱氧产生的O~(2-)在脱氧反应区可原位生成铝氧氟络合离子并进一步产生金属铝,从而引发阴极的铝热还原反应,导致V_2O_3熔盐电脱氧过程同时存在直接电还原反应和铝热还原反应,其中后者起着关键的加速作用;在熔盐中添加适量Al_2O_3可强化V_2O_3电脱氧过程,在其他条件不变的情况下电脱氧时间可缩短至3 h。     

Research and prospect on extraction of vanadium from vanadium slag by liquid oxidation technologies

A series of innovative green metallurgical processes using novel reaction media including the NaOH/KOH sub-molten salt media and the NaOH–NaNO₃ binary molten salt medium, for the extraction of vanadium and chromium from the vanadium slag have been developed. In comparison with the traditional sodium salt roasting technology, which operates at 850 °C, the operation temperatures of these new processes drop to 200–400 °C. Further, the extraction rates of vanadium and chromium utilizing the new approaches could reach 95% and 90%, respectively, significantly higher than those in the traditional roasting process, which are 75% and approximate zero, respectively. Besides, no hazardous gases and toxic tailings are discharged during the extraction process. Compared with the conventional roasting method, these new technologies show obvious advantages in terms of energy, environments, and the mineral resource utilization efficiency, providing an attractive alternative for the green technology upgrade of the vanadium production industries.     

Behavior of vanadium during reduction and smelting of vanadium titanomagnetite metallized pellets

The effects of CaO content, MgO content and smelting temperature on the vanadium behavior during the smelting of vanadium titanomagnetite metallized pellets were investigated. The thermodynamics of reduction and distribution of vanadium was analyzed and the high-temperature smelting experiments were carried out. The thermodynamic calculations show that the distribution ratio of vanadium between the slag and the hot metal decreases with the increments of CaO and MgO content in the slag as well as the increase of the smelting temperature. The smelting experiments demonstrate that the vanadium content in iron and the recovery rate of vanadium in pig iron increase as the CaO content, MgO content and smelting temperature increase, whereas the vanadium distribution ratio between the slag and iron tends to decrease. Moreover, the recovery rate of vanadium in pig iron has a rising trend with increasing the optical basicity of the slag. The addition of MgO in the slag to increase the slag optical basicity can not only improve the vanadium reduction but also promote the formation of magnesium-containing anosovite, which is beneficial to following titanium extraction.