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.

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.

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.

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.

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.

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。

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。

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。

钙盐添加剂焙烧对石煤提钒的影响研究

文章以湖南某地含钒石煤矿样作为研究对象,采用钙化焙烧工艺,经过水浸酸浸联合浸出,通过优化添加剂的组成比例、焙烧温度和时间等参数,结果表明:钙质添加剂为CaO和CaCl_(2),添加配比分别为6%和4%,在焙烧温度850℃的条件下焙烧90 min,浸出时间2.5 h的条件下,钒的浸出效果达到最佳,为91.1%。实现了钙盐添加剂对钒浸出率的显著提升。

含钒石煤的氧化焙烧机理

采用X射线衍射（XRD）、扫描电镜（SEM）和能谱分析（EDX）等技术对石煤氧化焙烧过程进行研究，考察焙烧对钒浸出的影响。结果表明：石煤在氧化焙烧过程中，有机质和黄铁矿首先被氧化，含钒伊利石晶体结构在750-850℃被破坏，V（Ⅲ）和V（Ⅳ）氧化反应达到平衡时，钒浸出率达到最大值；当焙烧温度高达1050℃时，物料烧结使钒被包裹，这是在1050℃焙烧后钒浸出率急剧降低的主要原因。

石煤提钒过程中钒氧化和转化对钒浸出的影响

采用H2SO4、HF浸出石煤原矿和石煤焙烧料中钒,通过浸出对比实验,考察了钒氧化和转化对钒浸出的影响。结果表明:含钒矿物晶体结构未被破坏时,V(Ⅲ)无法浸出;石煤原矿在3.5 mol/L HF中于60℃浸出8 h后,含钒矿物的晶体结构完全被破坏,钒浸出率可达到97.91%;通过焙烧或在浸出过程中添加NaClO3,可使V(Ⅲ)氧化为V(Ⅳ)或V(Ⅴ)。

微波焙烧对石煤提钒的影响

通过微波焙烧与传统焙烧后的对比实验,考察焙烧温度、焙烧时间和添加剂用量对石煤提钒的影响。结果表明,微波焙烧温度为700℃、焙烧60 min、添加剂(Na2CO3)用量为矿样质量的6%时,V2O5浸出率达到64.1%以上;经微波700℃焙烧120 min,矿样中小于0.074 mm粒级的含量提高10.0%左右。根据微波焙烧前后矿样粒度的变化,提出了微波焙烧过程矿样裂解模型,用此模型可以解释微波焙烧能够提高石煤中V2O5浸出率的原因。

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。

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1.2mL/g加水于100℃下搅拌浸出2h,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48h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。

碱法从石煤中浸出钒试验研究

采用造球—氧化焙烧—碱浸的方法从石煤中浸出钒,考察了焙烧温度、焙烧时间、浸出温度、浸出剂浓度、浸出时间、浸出液固比对浸出率的影响,获得了88.38%的高浸出率。研究表明,焙烧温度、浸出温度、浸出剂浓度、浸出时间是浸出率的重要影响因素。适宜的工艺条件是:焙烧温度850℃、焙烧时间3 h、浸出温度90℃、浸出剂浓度2 mol/L、浸出时间2 h、浸出液固比3。

石煤微波空白焙烧-酸浸提钒工艺

通过微波空白焙烧-酸浸提钒与传统加热焙烧-酸浸提钒和直接酸浸提钒的对比实验,考察H2SO4用量和浸出时间对石煤中钒浸出的影响。结果表明:石煤在700℃下微波焙烧60 min,H2SO4用量为矿样质量的22%,浸出温度为90℃时,V2O5浸出率达到83.50%,比传统加热焙烧-酸浸提钒和直接酸浸的浸出率提高约30%。通过对焙烧熟料酸浸和直接酸浸时Al2O3浸出率的分析发现,V2O5浸出率与Al2O3浸出率呈正相关性。结合矿物晶体构造与微波加热原理,探讨了微波焙烧改善酸浸提钒的机理,认为微波焙烧可破坏含钒云母的晶体结构,是提高V2O5浸出率的主要原因。

湘西含钒石煤提钒工艺研究

根据湘西含钒石煤矿的特点,采用添加剂X焙烧-酸浸工艺,从石煤中提钒,考察了添加剂用量、焙烧温度、焙烧时间、浸出温度、酸用量和浸出时间对钒浸出率的影响。结果表明,在添加剂用量为9%、焙烧温度为750℃、焙烧时间为2 h、浸出温度为50℃、酸用量为2%、浸出时间为4 h的条件下,钒的浸出率可达95%以上。钒浸出液经过树脂吸附-沉钒-煅烧得V2O5产品,主要技术指标均达到了化工用粉钒标准(GB3283-87)。

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%。

石煤提钒焙烧过程钒的价态变化及氧化动力学

含钒石煤是我国重要的钒矿资源,钒的焙烧氧化是石煤提钒的基础。以湖北某地钒矿为研究对象,首先采用电位滴定对氧化焙烧过程中钒的价态变化进行了研究。在此基础上,探索了钒氧化过程中控制反应速率的步骤,测定了动力学参数——表观活化能。研究表明,石煤中钒主要赋存在伊利石中,焙烧过程中钒的氧化过程是逐级氧化过程(V(Ⅲ)→V(Ⅳ)→V(Ⅴ)),钒的氧化反应属扩散动力学控制,其动力学表观活化能为E1=81.19 kJ/mol。