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%。     

酸浸对钙化焙烧提钒工艺钒浸出率的影响

采用稀硫酸浸出法提取钙化焙烧后钒渣中的钒,考察了浸出参数:物料粒度、体系pH值、浸出温度和时间、液固比(L/S)、搅拌速度对钒及杂质元素浸出率的影响.结果表明:物料粒度小于75μm时对提高钒浸出率影响较小;液固比从2∶1增加到7∶1,搅拌速度由100增加到500r/min时,钒浸出率增长幅度均低于3%;钒浸出率在浸出前15min内迅速升高,之后增长变缓;浸出体系pH值对钒及杂质浸出率影响显著,pH值为2~3时钒浸出率达90%,杂质元素Ca,Mn,Mg,Al,Si,P浸出率为10%~30%;在较佳浸出条件下:粒度96～75μm,pH值为25,温度55℃,时间30min,L/S为3,搅拌速度500r/min,钒浸出率超过91%.     

强化氧化对石煤钙化焙烧提钒的影响

研究石煤钙化焙烧参数对提钒效果的影响,确定合理的焙烧参数:当焙烧温度为950℃,焙烧时间3 h,碳酸钙添加量为质量分数6%时,石煤焙烧料中钒的浸出率为63%。在此基础上研究增强氧化对提钒效果的影响,比较通空气和添加MnO2这2种情况下提钒的效果。用化学物相分析和钒价态分析等技术探讨加强氧化提高钒浸出率的原因。空气通入速度为0.48 L/h时,钒的浸出率为69%;MnO2添加量(质量分数)为3%时,钒浸出率为68%。研究结果表明,加强氧化后矿石的结构被破坏,V5+含量提高,生成更多易溶于酸的钒酸钙类物质。     

无盐焙烧法提取石煤钒矿中V2O5新工艺研究

本实验主要采用无盐焙烧法对暮石煤钒矿石进行焙烧和浸出研究。在700-900℃高温下焙烧．然后用硫酸做浸出分解试羽，在80-90℃条件下从舍钒石煤矿石中钢得硫酸钒酰溶液（浸出液），得到最佳焙烧和浸出工艺参数。     

碱法浸出石煤中的钒和硅

在石煤提钒工艺中,为了充分利用石煤中的有价元素硅,采用碱浸提钒工艺提取石煤中的钒和硅.经过预焙烧后,可以有效地破坏石煤结构,提高钒硅浸出率.在焙烧温度850℃、焙烧时间2h、浸出温度95℃、浸出时间4h、固液比(g∶mL)1∶1.4、矿碱质量比1.2∶1的条件下,钒的浸出率为86.6%,硅的浸出率为61.4%.     

钒离子指示电极在石煤焙烧料浸出动力学研究中的应用

以湘西吉首某地石煤矿为原料,在850℃焙烧6 h后,采用酸浸法从焙烧料提取钒,用钒离子指示电极在线跟踪石煤浸出过程中五价钒浓度的变化,研究浸出过程中的动力学;考察矿石粒径、pH值和浸出温度对浸出过程E--t曲线的影响.研究结果表明:钒离子指示电极电势测定结果受杂质Fe3+的影响很小,可以用于跟踪石煤焙烧料浸出过程中钒离子浓度的变化;硫酸浸出后,矿石粒度越小,溶液的酸度越大,浸出温度越高,五价钒离子越容易浸出.采用等高线法求算出石煤焙烧料浸出过程在高温段区的表观活化能Ea为1.56 kJ/mol,受扩散步骤控制;在低温段区的表观活化能Ea为3.99 kJ/mol,受化学反应步骤控制.     

石煤钠化焙烧料酸浸动力学

研究了石煤钠化焙烧料硫酸浸出过程中,浸出剂初始浓度、搅拌速度和浸出温度对浸出率的影响,并对浸出过程动力学进行了分析. 结果表明:浸出剂初始浓度和浸出温度对钒浸出率有显著影响,搅拌速度对钒浸出率影响不大;该浸出过程符合核收缩模型,与化学反应控制动力学方程式相吻合,浸出反应的表观活化能为50.88kJ·mol-1,浸出过程控制步骤为化学反应控制.     

从石煤酸浸液中萃取钒的工艺研究

研究了三正辛胺从石煤酸浸液中萃取钒的工艺过程 ,从萃取和反萃的 p H值、相比、有机相组成、澄清时间等方面进行了详细试验。研究表明 :用三正辛胺萃取钒时 ,其萃取率可达98%以上 ;而且易反萃 ,用 0 .5 M Na2 CO3反萃时 ,反萃率为 99.9%。经萃取后 ,浸出液中的钒可由每升几克富集到每升数十克以上 ,有利于后续的提钒工艺。     

氧化钙化焙烧法应用于从钒云母矿中提取钒的研究

本文介绍用氧化钙化焙烧法从钒云母矿中提取钒的试验研究,对焙烧、浸出、净化、沉钒过程中各影响因素进行了探讨。研究结果表明,用石灰和钒云母矿混合焙烧生成钒酸钙,然后用碳铵溶液浸出钒,提取率高达78%,工艺简单、可靠,并且对环境污染小,投资少,它不失为一种可取的提钒新方法。     

响应曲面法优化钒渣微波钙化焙烧提钒工艺

采用相应曲面法,建立钙化添加剂用量、焙烧温度、焙烧时间与钒浸出率关系的数学模型,对钒渣微波钙化焙烧提钒工艺进行优化,并对试验结果的可靠性进行分析与验证。研究结果表明,采用响应曲面法优化钒渣微波钙化焙烧提钒工艺参数是可行的;微波钙化焙烧工艺参数对钒浸出率的影响从大到小依次为钙化添加剂用量、焙烧温度、焙烧时间;最佳焙烧参数为焙烧温度861.69℃、钙化添加剂用量1.02、焙烧时间106.31 min,此时钒的浸出率可达93.82%。     

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.     

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.     

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.     

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.     

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%.     

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.