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, Na₂CO₃ 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℃, respectively. Vanadium-rich phases were formed at above 850℃. The leaching rate of vanadium reached above 90% under the optimum conditions:roasting temperature of 850℃, roasting time of 60 min, Na₂CO₃ concentration of 160 g/L, leaching temperature of 95℃, 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.     

无焙烧钛白废酸氧压浸出转炉钒渣研究

采用无焙烧直接加压酸浸工艺,以钛白废酸为浸出剂,转炉钒渣为原料进行浸出提钒实验研究.热力学分析表明:可溶性含钒离子在酸性溶液中能够稳定存在.根据浸出实验得出:初始酸浓度是影响酸浸过程的重要因素,在初始酸质量浓度为250g·L-1,反应温度150℃,反应时间40min,液固比12∶1,氧分压02MPa的条件下,钒的浸出率为9851%.不同条件下的浸出渣XRD图谱表明:在钒浸出率增大的过程中,含钒尖晶石相逐渐消失,钛铁矿相发生转化形成锐钛矿相在浸出渣中富集.     

酸浸法提钒新工艺的研究

研究了用稀硫酸直接浸出—萃取—反萃—氨水沉钒—煅烧的提钒工艺。结果表明,采用稀硫酸直接浸出,原矿渣中总钒的一次浸取率可达95%以上;用萃取-反萃方式净化和浓缩浸出液,同时使用萃取促进剂处理酸浸液,使萃取效率比传统方法有明显提高,萃取级数大大减少;沉钒步骤摒弃了传统的铵盐沉钒工艺,使用氨水直接沉钒,提高了产品的纯度。钒的总回收率达86%以上,比传统提钒工艺效率提高了20%以上,同时由于避免了焙烧从而解决了传统提钒过程中因焙烧等产生的HCl、Cl₂等污染问题。     

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/Na OH 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_2O/Na DEHP(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.     

钒渣钙化焙烧的影响因素及焙烧氧化动力学

采用钙化焙烧方式处理转炉钒渣以提高钒的浸出率,考察了焙烧参数(渣样粒度,升温速率,焙烧保温温度及保温时间,配钙量)对钒浸出率的影响,根据钒渣氧化的TG-DSC曲线对钒渣氧化变温动力学进行了分析.结果表明:降低升温速率可提高钒氧化率,保温温度高于600℃时钒浸出率迅速增加.在钒渣粒径48~75μm,外配钙m(CaO)/m(V2O5)为042,升温速率2℃·min-1,保温温度850℃,保温时间150min的条件下,钒浸出率达9331%.钒尖晶石氧化过程受三级化学反应控制,升温速率为5和10℃·min-1的表观活化能分别为26765,25603kJ·mol-1.     

不同钙化剂对高钒渣酸浸提钒的影响

针对传统钒渣钠化焙烧-水浸提钒工艺的不足,确定对钒渣钙化焙烧-酸浸提钒进行研究。在理论分析的基础上,本研究以高钒渣为原料,研究了钙化焙烧-酸浸提钒过程中3种钙化剂(CaSO4、CaCO3、CaO)的焙烧机理以及对提钒效果的影响。研究结果表明:钒浸出率随焙烧温度的升高先增大后减小,且在1 450K时达到最大值;钙化剂配比为100%CaSO4时提钒率最大;在目前实验室研究条件下,钒的浸出率最大可达93.53%。     

熔融态高钙高磷钒渣氧化钠化—水浸提钒

采用XRD对钠化高钙高磷钒渣(11.48%V2O3、13.71%Ca O、0.78%P2O5)熟料的物相组成进行了分析,并研究了钒渣熟料提钒的最佳实验参数。结果表明:在Na2CO3加入量相对较少时(35%),V存在于Na4V2O7、Na3VO4、Na1.33V2O5和Na Ca VO4中,随着Na2CO3加入量的增加,Na4V2O7和Na Ca VO4会进一步与Na2CO3反应转化为Na3VO4;钒渣熟料中P存在于水溶性Na3PO4中;当实验条件如下:Na2CO3加入量为40%,液固比为5∶1 m L/g,浸出温度为90℃,浸出时间为4min,搅拌速度为150 r/min,高钙高磷钒渣熟料浸出率可超过90%。可见,熔融态高钙高磷钒渣氧化钠化水浸提钒的方法可行。     

采用氧化焙烧-酸浸法从高碳石煤中提钒试验研究

针对广西某难浸高碳石煤,比较相同焙烧和酸浸条件下静态焙烧矿和流态化焙烧矿钒的浸出率,优化流态化焙烧矿的酸浸条件。研究结果表明:流态化焙烧矿酸浸钒的浸出率比静态焙烧矿酸浸钒的浸出率平均高24%,所以,在相同焙烧温度、时间下流态化焙烧较静态焙烧更利于钒的浸出;在液固质量比为0.8:1.0,二氧化锰添加量为3%和氢氟酸添加量为2%的条件下,得最佳酸浸条件,即酸矿质量比为0.4:1.0,浸出温度为150℃,浸出时间为6 h,在此最佳酸浸条件下,钒浸出率可达88.26%。     

A new process of extracting vanadium from stone coal

A new process of extracting vanadium from the stone coal vanadium ore in Fangshankou, Dunhuang area of Gansu Province, China was introduced. Various leaching experiments were carried out, and the results show that the vanadium ore in Fangshankou is difficult to process due to its high consumption of acid and the high leaching rate of impurities. However, the leaching rate can be up to 80% and the content of V₂O₅ in the residue can be between 0.22%–0.25% in the process of ore fine grinding→oxidation roasting→mixing and ripening→aqueous leaching→P₂O₄ solvent extraction→sulfuric acid stripping→oxidation and precipitation→decomposition by heat. Also, the quality of flaky V₂O₅ produced by this process can meet the requirements of GB3283–87. The total leaching rate of vanadium is 70%. Also, three types of wastes are easy to treat. The vanadium extraction process is better in relation to the aspect of environmental protection than the sodium method.     

Influence of alumina on mineralogy and environmental properties of zinc-copper smelting slags

An iron-silicate slag, from a zinc-copper smelting process, and mixtures of this slag with 5wt%, 10wt%, and 15wt% alumina addition were re-melted, semi-rapidly solidified, and characterized using scanning electron microscopy equipped with energy dispersive spectroscopy, and X-ray diffraction. The FactSageTM6.2 thermodynamic package was applied to compare the stable phases at equilibrium conditions with experimental characterization. A standard European leaching test was also carried out for all samples to investigate the changes in leaching behaviour because of the addition of alumina. Results show that the commonly reported phases for slags from copper and zinc production processes (olivine, pyroxene, and spinel) are the major constituents of the current samples. A correlation can be seen between mineralogical characteristics and leaching behaviours. The sample with 10wt% alumina addition, which contains high amounts of spinels and lower amounts of the other soluble phases, shows the lowest leachabilities for most of the elements.     

Bioleaching of vanadium from barren stone coal and its effect on the transition of vanadium speciation and mineral phase

This study determined the optimal conditions required to obtain maximum vanadium extraction and examined the transition of mineral phases and vanadium speciation during the bioleaching process. Parameters including the initial pH value, initial Fe2+ concentration, solid load, and inoculum quantity were examined. The results revealed that 48.92wt% of the vanadium was extracted through bioleaching under optimal conditions. Comparatively, the chemical leaching yield (H₂SO₄, pH 2.0) showed a slower and milder increase in vanadium yield. The vanadium bioleaching yield was 35.11wt% greater than the chemical leaching yield. The Community Bureau of Reference (BCR) sequential extraction results revealed that 88.62wt% of vanadium existed in the residual fraction. The bacteria substantially changed the distribution of the vanadium speciation during the leaching process, and the residual fraction decreased to 48.44wt%. The X-ray diffraction (XRD) and Fourier transform infrared (FTIR) results provided evidence that the crystal lattice structure of muscovite was destroyed by the bacteria.     

Leaching of vanadium,sodium, and silicon from molten V-Ti-bearing slag obtained from low-grade vanadium-bearing titanomagnetite

The water leaching process of vanadium, sodium, and silicon from molten vanadium-titanium-bearing (V-Ti-bearing) slag obtained from low-grade vanadium-bearing titanomagnetite was investigated systematically. The results show that calcium titanate, sodium aluminosilicate, sodium oxide, silicon dioxide and sodium vanadate are the major components of the molten V-Ti-bearing slag. The experimental results indicate that the liquid-solid (L/S) mass ratio significantly affects the leaching process because of the respective solubilities and diffusion rates of the components. A total of 83.8% of vanadium, 72.8% of sodium, and 16.1% of silicon can be leached out via a triple counter-current leaching process under the optimal conditions of a particle size below 0.074 mm, a temperature of 90°C, a leaching time of 20 min, an L/S mass ratio of 4:1, and a stirring speed of 300 r/min. The kinetics of vanadium leaching is well described by an internal diffusion-controlled model and the apparent activation energy is 11.1 kJ/mol. The leaching mechanism of vanadium was also analyzed.     

Study on mechanisms of different sulfuric acid leaching technologies of chromite

The extraction of chromate from chromite via the sulfuric acid leaching process has strong potential for practical use because it is a simple and environmentally friendly process. This paper aims to study the sulfuric acid leaching process using chromite as a raw material via either microwave irradiation or in the presence of an oxidizing agent. The results show that the main phases in Pakistan chromite are ferrichromspinel, chrompicotite, hortonolite, and silicate embedded around the spinel phases. Compared with the process with an oxidizing agent, the process involving microwaves has a higher leaching efficiency. When the mass fraction of sulfuric acid was 80% and the leaching time was 20 min, the efficiency could exceed 85%. In addition, the mechanisms of these two technologies fundamentally differ. When the leaching was processed in the presence of an oxidizing agent, the silicate was leached first and then expanded. By contrast, in the case of leaching under microwave irradiation, the chromite was dissolved layer by layer and numerous cracks appeared at the particle surface because of thermal shock. In addition, the silicate phase shrunk instead of expanding.     

Innovative method for boron extraction from iron ore containing boron

A novel process for boron enrichment and extraction from ludwigite based on iron nugget technology was proposed. The key steps of this novel process, which include boron and iron separation, crystallization of boron-rich slag, and elucidation of the boron extraction behavior of boron-rich slag by acid leaching, were performed at the laboratory. The results indicated that 95.7% of the total boron could be enriched into the slag phase, thereby forming a boron-rich slag during the iron and slag melting separation process. Suanite and kotoite were observed to be the boron-containing crystalline phases, and the boron extraction properties of the boron-rich slag depended on the amounts and grain sizes of these minerals. When the boron-rich slag was slowly cooled to 1100℃, the slag crystallized well and the efficiency of extraction of boron (EEB) of the slag was the highest observed in the present study. The boron extraction property of the slow-cooled boron-rich slag obtained in this study was much better than that of szaibelyite ore under the conditions of 80% of theoretical sulfuric acid amount, leaching time of 30 min, leaching temperature of 40℃, and liquid-to-solid ratio of 8 mL/g.     

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 CaO, molar ratio between CaO and V₂O₅ (n(CaO)/n(V₂O₅)), 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 CaO 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_0.6Cr_0.4)₂O₃. 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 CaO. Moreover, CaO was more likely to combine with vanadium, as further confirmed by thermodynamic analysis. When the HCVS with CaO added in an n(CaO)/n(V₂O₅) ratio of 0.5 was roasted in an air atmosphere at a heating rate of 10℃/min from room temperature to 950℃ and maintained at this temperature for 60 min, the leaching ratios of vanadium and chromium reached 91.14% 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 592 and 630℃ for n(CaO)/n(V₂O₅) ratios of 0.5 and 5, respectively.     

硫酸中香蕉皮还原低品位锰矿的浸出行为及动力学模型

Manganese was leached from a low-grade manganese ore (LGMO) using banana peel as the reductant in a dilute sulfuric acid medium. The effects of banana peel amount, H₂SO₄ concentration, reaction temperature, and time on Mn leaching from the complex LGMO were studied. A leaching efficiency of ~98% was achieved at a leaching time of 2 h, banana peel amount of 4 g, leaching temperature of 120°C, manganese ore amount of 5 g, and sulfuric acid concentration of 15vol%. The phase, microstructural, and chemical analyses of LGMO samples before and after the leaching process confirmed the successful leaching of manganese. Furthermore, the leaching process followed the shrinking core model and the leaching rate was controlled by a surface chemical reaction (1 ? (1 ? x)1/3 = kt) mechanism with an apparent activation energy of 40.19 kJ·mol?1.     

A method for recovery of iron,titanium, and vanadium from vanadium-bearing titanomagnetite

An innovative method for recovering valuable elements from vanadium-bearing titanomagnetite is proposed. This method involves two procedures:low-temperature roasting of vanadium-bearing titanomagnetite and water leaching of roasting slag. During the roasting process, the reduction of iron oxides to metallic iron, the sodium oxidation of vanadium oxides to water-soluble sodium vanadate, and the smelting separation of metallic iron and slag were accomplished simultaneously. Optimal roasting conditions for iron/slag separation were achieved with a mixture thickness of 42.5 mm, a roasting temperature of 1200℃, a residence time of 2 h, a molar ratio of C/O of 1.7, and a sodium carbonate addition of 70wt%, as well as with the use of anthracite as a reductant. Under the optimal conditions, 93.67% iron from the raw ore was recovered in the form of iron nugget with 95.44% iron grade. After a water leaching process, 85.61% of the vanadium from the roasting slag was leached, confirming the sodium oxidation of most of the vanadium oxides to water-soluble sodium vanadate during the roasting process. The total recoveries of iron, vanadium, and titanium were 93.67%, 72.68%, and 99.72%, respectively.     

Effects of the mineral phase and valence of vanadium on vanadium extraction from stone coal

The influence of roasting on the leaching rate and valence of vanadium was evaluated during vanadium extraction from stone coal. Vanadium in stone coal is hard to be leached and the leaching rate is less than 10% when the raw ore is leached by 4 mol/L H₂SO₄ at 90℃ for 2 h. After the sample is roasted at 900℃ for 2 h, the leaching rate of vanadium reaches the maximum, and more than 70% of vanadium can be leached. The crystal of vanadium-bearing mica minerals decomposes and the content of V(V) increases with the rise of roasting temperature from 600 to 900℃, therefore the leaching rate of vanadium increases significantly with the decomposition of the mica minerals. Some new phases, anorthite for example, form when the roasting temperature reaches 1000℃. A part of vanadium may be enwrapped in the sintered materials and newly formed phases, which may impede the oxidation of low valent vanadium and make the leaching rate of vanadium drop dramatically. The leaching rate of vanadium is not only determined by the valence state of vanadium but also controlled by the decomposition of vanadium-bearing minerals and the existence state of vanadium to a large extent.     

钒渣无焙烧浸出液中钒铁萃取分离

利用本课题组提出的钛白废酸无焙烧加压浸出钒渣提钒的新技术,以P204为萃取剂从废酸浸出钒渣的浸出液中进行了提钒研究.实验结果表明:采用亚硫酸钠为浸出液预处理还原剂,将浸出液中三价铁还原成二价铁,从而防止三价铁的共萃;常温条件下,当浸出液初始p H=2.5、水相与有机相体积比为1∶3,震荡时间为4 min时,采用有机相组成为20%P204及10%TBP协同萃取体系,钒的萃取率可达98.61%以上,钒铁的分离系数可达135.3.     

模拟不同pH值的雨水对炼锌废渣浸出毒性的影响

通过采集贵州省毕节地区赫章县、威宁县以及六盘水市水城县一带土法炼锌废渣,参照黔西北地区特殊的大气降雨pH变化,以pH值为3、4、5、5.5和6的硫酸溶液作为浸提剂,采用水平振荡法对废渣进行浸出实验。结果表明,炼锌废渣浸出液中Cu、Cd的浸出浓度与浸提剂pH呈显著负相关,Zn的浸出浓度与浸提剂pH的相关性不明显。在不同pH值的硫酸溶液的浸泡下,浸出液中Zn、Cu浓度变化范围分别为0.53-24.39 mg/L、0.011-10.38 mg/L,其浸出毒性在最高允许浓度范围内;Cd浓度变化范围为0.012-3.34 mg/L,浸出液中Cd最高浓度超限值11倍。同时,浸提剂pH对浸出液pH无影响,浸出液pH均值在7.43-7.86之间变动,无腐蚀性;因此,在酸沉降环境下,炼锌废渣是一种有害固体废物,其浸出毒性主要表现为Cd污染。