Bioleaching of vanadium from stone coal vanadium ore by Bacillus mucilaginosus:Influencing factors and mechanism

Vanadium and its derivatives are used in various industries,including steel,metallurgy,pharmaceuticals,and aerospace engineering.Although China has massive reserves of stone coal resources,these resources have low grades.Therefore,the effective extraction and recovery of metallic vanadium from stone coal is an important way to realize the efficient resource utilization of stone coal vanadium ore.Herein,Bacillus mucilaginosus was selected as the leaching strain.The vanadium leaching rate reached 35.5%after 20 d of bioleaching under optimal operating conditions.The cumulative vanadium leaching rate in the contact group reached 35.5%,which was higher than that in the noncontact group(9.3%).The metabolites of B.mucilaginosus,such as oxalic,tartaric,citric,and malic acids,dominated in bioleaching,accounting for 73.8%of the vanadium leaching rate.Interestingly,during leaching,the presence of stone coal stimulated the expression of carbonic anhydrase in bacterial cells,and enzyme activity increased by 1.335-1.905 U.Enzyme activity positively promoted the production of metabolite organic acids,and total organic acid content increased by 39.31 mg·L^(-1),resulting in a reduction of 2.51 in the pH of the leaching system with stone coal.This effect favored the leaching of vanadium from stone coal.Atomic force microscopy illustrated that bacterial leaching exacerbated corrosion on the surface of stone coal beyond 10 nm.Our study provides a clear and promising strategy for exploring the bioleaching mechanism from the perspective of microbial enzyme activity and metabolites.

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.

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.

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 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.

φ-pH diagram of V-Ti-H_2O system during pressure acid leaching of converter slag containing vanadium and titanium

To analyze the thermodynamic characteristics of leaching process of converter slag, φ-pH diagram of V-Ti-H2O system at oxygen partial pressure of 0.5 MPa, ionic mass concentration of 0.1 mol/kg and temperatures ranging from 60 to 200 ℃ was obtained by recently published critically assessed standard Gibbs energies and activity coefficients of various species. When pH2, stable regions of V3＋, VO2＋ and VO2＋ exist in the stable region of TiO2. The pH values of stable regions of vanadium and titanium decrease and redox potentials become more positive with the temperature increasing. Vanadium and titanium could be separated by one-step leaching based on thermodynamics. The experiment results of pressure acid leaching of converter slag show that leaching rates of vanadium and titanium are 96.87% and 8.76% respectively, at 140 ℃ of temperature, 0.5 MPa of oxygen partial pressure, 0.055-0.075mm of particle size, 15：1 of liquid to solid ratio, 120 min of leaching time, 500 r/min of stirring speed and 200 g/L of initial acid concentration. Vanadium and titanium could be selectively separated in the pressure acid leaching process, and the experiment result is in agreement with thermodynamic calculation result.

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 vana-dium 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 tem-perature 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 vana-dium reaches 97.69%.

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.

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

The extraction technology of vanadium from stone coal by means of no-salt-roasting and dilute acid solution leaching treatment has the advantages of low pollution, low investment, as well as high vanadium leaching efficiency, which makes it the best technology for extending application. In the present study, the effects of vanadium occurrence state in stone coal, roasting temperature and additive agent on the leaching efficiency of vanadium were studied. The results indicate that the effect of vanadium occurrence state on the extracting vanadium technology is obvious. If the vanadium component in stone coal existed in amorphous phase form, this type of stone coal can be treated by no-slat-roasting and dilute acid solution leaching technology; while the vanadium-bearing crystalline phase existed in stone coal, the roasting additive must be added to destroy this crystalline form so as to acquire high vanadium leaching rate.

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.

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.

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.

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.

Extraction of vanadium from molten vanadium bearing slag by oxidation with pure oxygen in the presence of CaO

A novel process of vanadium extraction from vanadium slag in its molten state was conducted at the laboratory scale by oxidation with pure oxygen in the presence of CaO. The effect of mass ratio of CaO to V2O5 on the recovery of vanadium was studied. The sintered samples were leached by H2SO4 solution and characterized by XRD, XPS, SEM and EDS techniques. Compared with the roasting process, the energy saving effect of the proposed process was also discussed. The results showed that vanadium-rich phases were formed and vanadium mainly existed in the forms of CaV2O5 and Ca2V2O7. The formation mechanism of calcium vanadates in the molten vanadium bearing slag was explained. The XRD and XPS results implied that there was a limit to the oxidation reaction of V(IV) to V(V) under the high temperatures even though oxygen-supply was sufficient. An increase in the CaO content led to an increase in the formation of Ca2V2O7. About 90%of the vanadium recovery was obtained under optimal experiment conditions (mass ratio of CaO to V2O5 of 0.6, particle size 120 to 150μm, leaching temperature 90 °C, leaching time 2 h, H2SO4 concentration 20%, liquid to solid ratio 5:1 mL/g, stirring speed 500 r/min). The energy of 1.85×106 kJ could be saved in every 1000 kg of vanadium bearing slag using the proposed process from the theoretical calculation results. Recovery of vanadium from the molten vanadium bearing slag and utilisation of its heat energy are important not only for saving metal resources, but also for energy saving and emission reduction.

钒渣酸浸提铁工艺优化及复合光催化剂的制备

钒渣来自转炉钒渣,含有大量铁元素,它的二次利用具有重大的经济价值与现实意义。本文以H2SO4为酸浸液,利用响应面法优化酸浸工艺参数,后通过煅烧-水热法成功制备出ɑ-Fe_(2)O_(3)/Bi_(2)WO_(6)复合光催化剂,在模拟可见光作用下光降解染料废水甲基橙(MO)。结果表明,响应面优化酸浸实验的最佳工艺参数为H2SO4浓度2.8mol/L、酸浸温度98℃、酸浸时间138min、固液比1∶4.2,钒渣中Fe的浸出率为71.0%;以钒渣为铁源,采用沉淀-煅烧法制备出α-Fe_(2)O_(3)后,通过水热法合成VS-Fe_(2)O_(3)/Bi_(2)WO_(6),其光降解性能明显优于商业Fe_(2)O_(3)/Bi_(2)WO_(6),VS-Fe_(2)O_(3)的最佳掺杂量为10%,当H2O2投加量为19.58mmol/L、p H=6.5、催化剂投加量为0.4g/L、MO初始浓度为10mg/L、光反应6h时,MO去除率达96%,COD去除率为88.4%,出水COD为13.34mg/L,符合《污水综合排放标准》一级标准,光反应速率常数为纯相Bi_(2)WO_(6)的4.8倍;MO的降解机理主要为H2O2协同VS-Fe_(2)O_(3)/Bi_(2)WO_(6)作用下的光芬顿反应。本文可为钒渣高效、清洁的资源化利用及其在光催化领域的资源化利用提供参考。

石煤钒矿的钡盐焙烧—酸浸提钒工艺研究

针对含钒物相以绢云母为主、钒以类质同象形式赋存的石煤钒矿,研究了用重晶石焙烧—酸浸工艺从中提取钒,考察了磨矿细度、重晶石用量、焙烧温度、焙烧时间、液固体积质量比、硫酸用量、浸出温度及浸出时间对钒浸出率的影响。结果表明:石煤钒矿磨矿细度为-200目占比70%、重晶石用量5%、焙烧温度850℃、焙烧时间16 h、液固体积质量比2/1、硫酸用量8%、浸出温度25℃和浸出时间2 h条件下,钒浸出率可达89%以上,浸出效果较好。

酸浸法富集含钒钢渣中铁/钒元素研究

本试验通过酸浸法对含钒钢渣中铁/钒元素进行富集,研究了醋酸浓度(体积分数,下同)、液固比、酸浸时间、酸浸温度对铁/钒氧化物富集的影响。结果表明,液固比是影响酸浸渣中铁/钒氧化物含量的主要因素,醋酸浓度次之,酸浸时间和酸浸温度影响较小。在醋酸浓度为5%,液固比为20 mL/g,酸浸温度为30℃,酸浸时间为1 h的条件下,酸浸渣中Fe_(2)O_(3)、V_(2)O_(5)富集效果最佳,分别达到52.06%和2.15%;此时,酸浸渣中CaO含量降至19.42%。在此条件下得到的酸浸渣碱度系数R为1.3,处于钢铁冶炼要求的推荐渣性区间,将其直接熔融还原,铁、钒氧化物还原率分别达到了91.84%和66.51%。

煤气化细渣高炭组分超声强化酸浸法制备多孔材料

煤气化细渣是煤炭清洁高效利用的副产物之一,其资源化应用迫在眉睫。通过简单筛分得到固定碳含量高于60%的高炭组分,并以此为原料,采用超声酸浸法制备多孔材料。以核废水中放射性碘的吸附处理为应用背景,用碘吸附值表征多孔材料的吸附性能。结合SEM、BET、XRD和FT-IR等性质和结构分析方法,系统研究了超声时间、超声功率、酸浓度和温度对多孔材料碘吸附性能和组成结构的影响规律;并探讨了超声强化酸浸对残炭的组成结构的影响机制和灰成分的迁移转化规律,总结出超声强化酸浸作用机理。结果表明,煤气化细渣高炭组分在酸浓度为4 mol/L、酸浸温度为50℃、超声功率为210 W,超声时间1.5 h的条件下,所制备多孔材料的碘吸附性能最佳,为468.53 mg/g,比表面积达到474.97 m^(2)/g,且具有以介孔为主的丰富孔隙结构。各因素对多孔材料碘吸附性能影响的顺序为:超声时间>酸浓度>超声功率>酸浸温度。超声强化酸浸作用机理是超声空化和机械波作用一方面强化炭灰黏附颗粒的解离,使堵塞在气化细渣孔道内的灰颗粒脱附,增加孔隙结构的连通性;其次,会导致炭灰颗粒表面裂纹的产生,增强碳颗粒内部无机组分的可及性;第三,能够提高酸浸过程的传质速率,强化气化细渣中的无机组分的浸出效果。

煤气化渣除杂富铝制备聚合氯化铝实验研究

煤气化渣是煤气化工艺产生的固体废弃物,随着煤气化工艺的广泛推广与应用,大量煤气化渣的产生对生态环境造成威胁的同时也制约着煤化工行业的发展。为实现煤气化渣的资源化、工业化利用,以煤气化粗渣为研究对象,对其矿物赋存形态和化学组成进行分析,提出煤气化渣净化除杂-酸浸提铝制备聚合氯化铝(polyaluminum chloride,PAC)的总体构思。利用煤气化渣中各元素与酸溶液的反应活性差异,在低酸体系下浸出Fe^(2+),Ca^(2+)及少部分Al^(3+),再利用离子沉淀pH差异,将钙、铁与煤气化渣选择性分离,实现酸渣富铝。以富铝酸渣为原料,酸浸制备PAC前驱体。考察铝酸钙粉添加量、聚合温度、聚合时间对产品品质及絮凝效果的影响。结果表明:在60%理论耗酸量下,煤气化渣浸出Al^(3+),Fe^(2+),Ca^(2+),将反应液pH调至3.5,Al^(3+)选择性沉淀,固液分离后制备富铝渣,钙、铁的去除率可达84.06%和79.79%,Al_(2)O_(3)质量分数由原来的10.84%提升至14.38%;富铝酸渣添加酸后再次溶出,酸溶液中氧化铝质量浓度可达到28.4 g/L;按固液比为10 g∶100 mL向酸溶液添加铝酸钙粉,在80℃下聚合2.5 h,可获得符合GB/T 22627-2022要求的PAC产品;通过FTIR和XRD对PAC产品进行表征,表明生成的产品为PAC,将该产品应用于模拟废水中,去浊率高达92.90%。