Mechanisms of aid-leaching reagent calcium fluoride in the extracting vanadium processes from stone coal

The mechanisms of aid-leaching reagent calcium fluoride in the extracting vanadium processes were researched by comparing the blank extraction with the extraction with calcium fluoride as aid-leaching reagent. The leaching experiments were carried out under the conditions of 95 ~C, 6 h, 1.5 ml.g-1, 15 vol % H2SO4, and 5 % calcium fluoride （mass ratio of CaF2 to stone coal）. The concentrations of hydrogen ion were adjusted from initial 1.82 and 4.79 mol.L-1 to the pH of 1.80 4- 0.05 by 200 g.L-1 lime milk, respectively. The solvent extraction experiments were carded out under the organic phase using 15 vol% D2EHPA, 5 vol% TBP, and 80 vol% sulfonated kerosene and O/A phase ratio of 1/4. The leaching rate of vanadium and six-stage countercurrent solvent extraction rate of vanadium can be remarkably improved from 66.71 to 92.97 % and 81.15 to 98.81%, respectively, when calcium fluoride was added as aid-leaching reagent. In the leaching process, fluorine can enhance the breakage of the structure of vanadium-bearing minerals. Meanwhile, fluoride ion can combine with iron ion to generate acid-insoluble iron fluoride hydrate （Fe2Fs.7H20） and consume all ferric ions and a portion of ferrous ions. In adjusting pH value process, the massive impurities are decreased to form KA1SO4 and FeA1Fs. In solvent extraction process, fluorine can prevent the formation of ferric hydroxide colloids and eliminate the formed unstable ferric hydroxide colloids and the silicon-containing colloids to promote the sufficient contact between the aqueous phase and the organic phase and the sufficient reaction.

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 moUL 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.

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.

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.

Extraction of vanadium from stone coal by modified salt-roasting process

In order to reduce the pollution of Cl2 and HCl released during extracting vanadium from stone coal by sodium chloride roasting, a modified salt-roasting process was proposed by adding calcined lime in roasting process followed by H2SO4 leaching. The effects of parameters including roasting temperature, roasting time, addition mass ratio of NaCl, calcined lime upon leaching rate of vanadium and curing rate of chlorine were investigated, and the effects of leaching time and leaching temperature on leaching rate of vanadium were also studied. The results show that the vanadium leaching rate and the curing rate of chlorine are 67.3% and 51.5% (mass fraction), respectively, at roasting temperature of 750 °C, roasting time of 4 h, 15% sodium chloride and 8% (mass fraction) calcined lime, leaching temperature of room temperature, and leaching time of 4 h.

Effect of vanadium occurrence state on the choice of extracting vanadium technology from stone coal

钒从的抽取技术借助于 no-salt-roasting 投掷煤并且冲淡酸答案沥滤处理有低污染，低投资，以及高钒沥滤效率的优点，它为扩大申请使它成为最好的技术。在现在的学习，钒出现的效果在石头煤声明钒的沥滤的效率上的温度和添加剂代理人被学习，烤。结果显示提取的钒技术上的钒出现状态的效果是明显的。如果在石头煤的钒部件在非结晶的阶段形式存在，这类石头煤能被 no-slat-roasting 对待并且冲淡酸答案沥滤技术；当忍受钒的水晶的阶段在石头煤存在时，烤的添加剂必须被增加破坏这种水晶的形式以便高获得钒沥滤率。

Comparison of the mechanisms of microwave roasting and conventional roasting and of their effects on vanadium extraction from stone coal

Experiments comparing microwave blank roasting and conventional blank roasting for typical vanadium-bearing stone coal from Hubei Province in central China, in which vanadium is present in muscovite, were conducted to investigate the effects of roasting tempera- ture, roasting time, H2SO4 concentration, and leaching time on vanadium extraction. The results show that the vanadium leaching efficiency is 84% when the sample is roasted at 800℃ for 30 min by microwave irradiation and the H2SO4 concentration, liquid/solid ratio, leaching temperature, and leaching time are set as 20vo1%, 1.5：1 mL.g-1, 95℃, and 8 h, respectively. However, the vanadium leaching efficiency achieved for the sample subjected to conventional roasting at 900℃ for 60 min is just 71% under the same leaching conditions. Scanning electron microscopy （SEM） analysis shows that the microwave roasted samples contain more cracks and that the particles are more porous compared to the conventionally roasted samples. According to the results of X-ray diffraction （XRD） and Fourier-transform infrared （FTIR） analyses, neither of these roasting methods could completely destroy the mica lattice structure under the experimental conditions; however, both methods deformed the muscovite structure and facilitated the leaching process. Comparing with conventional roasting, microwave roasting causes a greater deformation of the mineral structure at a lower temperature for a shorter roasting time.

A new process of extracting vanadium from stone coal

A new process of extracting vanadium from the stone coal vanadium ore in Fangshankou, Dtmhuang 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 fme 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 V205 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.

Leaching vanadium by high concentration sulfuric acid from stone coal

An effectively new technology of extracting vanadium from stone coals by high concentration sulfuric acid was researched. The effect of the concentration of sulfuric acid,leaching temperature,leaching time and helper leaching agent on the extraction of vanadium was explored.The results show that the optimal conditions of extraction are as follows:the concentration of sulfuric acid is 6 mol/L,the ratio of liquid to solid is 3-1;the temperature is 90℃;the leaching time is 3-5 h,the diameter of the ore particle is less than 180μm,and the concentration of helper leaching agent R is 6%.Under these conditions,the extraction of vanadium can reach 95.86%.

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.

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 p H value, initial Fe^(2+) concentration, solid load, and inoculum quantity were examined. The results revealed that 48.92 wt% of the vanadium was extracted through bioleaching under optimal conditions. Comparatively, the chemical leaching yield(H_2SO_4, pH 2.0) showed a slower and milder increase in vanadium yield. The vanadium bioleaching yield was 35.11 wt% greater than the chemical leaching yield. The Community Bureau of Reference(BCR) sequential extraction results revealed that 88.62 wt% 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 48.44 wt%. 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.

Desiliconisation of alkaline leaching solution of roasted stone coal with carbonation method

The effects of reaction time,reaction temperature,stirring speed and flowrate of CO2 gas on desilication rate and loss rate of vanadium were studied.The results indicated that desilication rate increases with the increase of flowrate of CO2 gas and reaction time.Reaction temperature and stirring speed have little effect on desilication rate,while influence the loss rate of vanadium significantly.Under the condition of reaction time of 2 h,reaction temperature of 95℃,stirring speed of 180 r/min,flowrate of CO2 gas of 60 mL/min and aging time of 2 h,desilication rate is more than 96%and the loss rate of vanadium is about 4.24%.The residue of desilication process can be processed for silicon materials,such as high-grade hydrated silica,which commonly known as white carbon black.In addition,with this carbonation method,leaching regents NaOH can be recycled by simple treatments.

电感耦合等离子体原子发射光谱(ICP-AES)法测定石煤钒矿石中钒、铁、钛的含量

研究一种快速准确测定石煤钒矿石中钒、铁、钛含量的方法对石煤钒矿检测行业的发展和石煤钒矿资源的开发利用、综合评价具有重要的意义。试样经过氧化钠在650℃完全熔融后保持1~3 min,取出冷却,热水浸提,盐酸酸化,定容稀释,选择V 292.402 nm、Fe 259.940 nm、Ti 334.941 nm为分析谱线,利用电感耦合等离子体原子发射光谱法(ICP-AES),建立了石煤钒矿石中同时测定钒、铁、钛含量的方法。优化了钒、铁、钛元素的分析谱线和背景校正模式,采用全程序样品空白稀释配制的校准曲线消除盐度钠基影响;研究了坩埚、熔剂种类、直接碱熔与灼烧-碱熔前处理方式、酸度效应的影响。结果表明:方法检出限为0.009%~0.019%,定量限为0.027%~0.057%;铁、钛、钒3种元素浓度在0.10~30.0μg/mL时,校准曲线线性回归方程的相关系数均大于0.9999;采用国家一级标准物质(GBW07876、GBW07877、GBW07878)进行验证,相对标准偏差(RSD,n=12)为0.49%~0.87%,且测定值与标准值吻合。方法具有快速简便、精密度好、准确度高、同时测定钒、铁、钛等优点,满足石煤钒矿石化学分析检测需求,且可为石煤钒矿化学分析标准方法的建立提供技术支撑。

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%。

悬浮焙烧法强化从难选石煤中提取钒

提出一种难选石煤在空气气氛下悬浮焙烧-酸浸强化的提钒新技术。在焙烧温度为800℃、焙烧时间为20 min、气体流量为400 m L/min的条件下钒的浸出率从过去的20%提高到目前技术的47.14%。在焙烧过程中,石煤表面逐渐变得粗糙和不规则,颗粒比表面积增加,硅酸盐矿物的片状结构被破坏,促进了钒的释放。同时,石煤中的钒被氧化为V(Ⅴ)或V(Ⅳ)。结果表明,由于悬浮焙烧过程中钒的释放和转化,因此,其浸出率得到了提高。

低钙镁石煤钒矿熟化浸出工艺研究

针对低钙镁钒矿碱度低、二氧化硅含量高的特点,采用硫酸熟化-水浸出工艺处理石煤钒矿,考察了矿物粒级、硫酸用量、熟化温度、熟化时间等参数对钒浸出的影响。结果表明,在矿石粒度-0.15 mm、硫酸用量20%、熟化温度120℃、熟化时间8.5 h的优化条件下,所得熟料在液固比2∶1、室温下用水搅拌浸出2 h,钒浸出率达到了88.6%。熟化过程矿物中包裹钒的云母结构得到高效解离,并释放可溶性钒,浸出液中硅含量低(0.08 g/L),比直接酸浸或焙烧酸浸工艺中硅含量大幅降低,对后续萃取工序非常有利。

石煤沸腾焙烧料提取五氧化二钒工艺研究

针对石煤沸腾焙烧料提钒原料差异大、工艺流程过长、钒回收率低(大多低于60%)、成本偏高等问题,采用拌酸熟化浸出-氧化中和-离子交换-铵盐沉钒-煅烧的工艺路线提钒。实验结果表明,在硫酸用量为原料质量的20%、熟化温度100℃、熟化时间6 h的条件下,石煤沸腾焙烧料中的钒浸出率达到90.2%,最终制备得到的产品V_(2)O_(5)粉末质量稳定可靠,达到了行业标准水平。该工艺流程短,工程建设投资低、钒回收率高、提钒冶炼成本低,取得了较好的经济技术指标。

添加石煤提钒尾渣对建筑用烧结砖性能的影响

本研究以石煤提钒尾渣为原料制备了建筑用烧结砖,研究了石煤提钒尾渣细度及添加量对制品抗压强度及吸水率的影响,采用XRD和SEM表征手段分析了石煤提钒尾渣添加量调控烧结砖相组成与显微结构的机理,检测了制品的抗冻融性与使用安全性。结果表明:提高尾渣细度改善了样品的致密度与抗压强度,引入尾渣使样品的抗压强度先增大后减小。B3样品(尾渣:黏土:粉煤灰=30:30:40)经1150 o C烧成后,其抗压强度≥35 MPa,吸水率小于≤13%,综合性能满足《烧结普通砖》(GB/T 5101-2003)中Mu30的要求。XRD与SEM分析表明,添加15%~30%尾渣有利于生成液相,促进钙长石与钙黄长石的生成,晶粒与液相相互胶结,使制品具有较高的抗压强度与致密度及较好的抗冻融性能。

浓硫酸熟化浸出提取湘西石煤中的钒

采用浓硫酸熟化浸出湘西石煤中的钒,通过单因素实验和正交实验研究了硫酸用量、脱碳温度、熟化时间、熟化温度以及浸出时间对钒浸出率的影响。实验结果表明,在脱碳温度600℃、硫酸用量24%、熟化温度180℃、熟化时间8 h、浸出时间8 h条件下,钒浸出率为83.2%。该工艺提钒过程无有害气体产生,减少了环境污染。