A novel process for the recovery of iron,titanium, and vanadium from vanadium-bearing titanomagnetite:sodium modification-direct reduction coupled process

A sodium modification-direct reduction coupled process was proposed for the simultaneous extraction of V and Fe from vanadium-bearing titanomagnetite. The sodium oxidation of vanadium oxides to water-soluble sodium vanadate and the transformation of iron oxides to metallic iron were accomplished in a single-step high-temperature process. The increase in roasting temperature favors the reduction of iron oxides but disfavors the oxidation of vanadium oxides. The recoveries of vanadium, iron, and titanium reached 84.52%, 89.37%, and 95.59%, respectively. Moreover, the acid decomposition efficiency of titanium slag reached 96.45%. Compared with traditional processes, the novel process provides several advantages, including a shorter flow, a lower energy consumption, and a higher utilization efficiency of vanadium-bearing titanomagnetite resources.     

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.     

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.     

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.     

以煤泥为还原剂海滨钛磁铁矿直接还原焙烧反应历程

研究以煤泥为还原剂,印尼某海滨钛磁铁矿在直接还原焙烧过程中,不同焙烧温度下矿物组成变化规律. X射线衍射和扫描电镜分析结果表明,随着焙烧温度的升高,钛磁铁矿逐渐被还原. 其中铁矿物经过浮士体( FeO) ,最终被还原成金属铁;而钛则经过钛尖晶石最终生成钛铁矿和少部分的铁板钛矿. 在整个直接还原焙烧过程中,金属铁颗粒在1100℃左右生成,然后不断长大,在1250℃时金属铁颗粒明显增多,在之后的保温过程中,金属铁颗粒不断长大,并在此过程中将金属铁从中分离出来.     

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%.     

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%。     

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.     

碳酸钠在高磷鲕状赤铁矿直接还原中的作用

为了研究碳酸钠对尼日利亚某高磷鲕状赤铁矿直接还原焙烧-磁选脱磷效果的影响,采用X射线衍射(XRD)和扫描电镜(SEM)研究了添加碳酸钠后直接还原焙烧的产物.结果表明,还原焙烧过程中添加碳酸钠后可以实现脱磷:碳酸钠的加入抑制了铁橄榄石的生成,阻断了磷进入金属铁的过程;使得鲕粒结构破坏,促进金属铁颗粒的聚集长大,有利于金属铁颗粒与脉石的解离;原矿中含磷矿物在焙烧过程中与碳酸钠反应生成可溶性的Na3PO4,在磨矿磁选过程中溶于水,使直接还原铁中磷的含量降低.     

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。     

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.     

Effects of smelting parameters on the slag/metal separation behaviors of Hongge vanadium-bearing titanomagnetite metallized pellets obtained from the gas-based direct reduction process

Smelting separations of Hongge vanadium-bearing titanomagnetite metallized pellets (HVTMP) prepared by gas-based direct reduction were investigated, and the effects of smelting parameters on the slag/metal separation behaviors were analyzed. Relevant mechanisms were elucidated using X-ray diffraction analysis, FACTSAGE 7.0 calculations, and scanning electron microscopy observations. The results show that, when the smelting temperature, time, and C/O ratio are increased, the recoveries of V and Cr of HVTMP in pig iron are improved, the recovery of Fe initially increases and subsequently decreases, and the recovery of TiO₂ in slag decreases. When the smelting CaO/SiO₂ ratio is increased, the recoveries of Fe, V, and Cr in pig iron increase and the recovery of TiO₂ in slag initially increases and subsequently decreases. The appropriate smelting separation parameters for HVTMP are as follows: smelting temperature of 1873 K; smelting time of 30–50 min; C/O ratio of 1.25; and CaO/SiO₂ ratio of 0.50. With these optimized parameters (smelting time: 30 min), the recoveries of Fe, V, Cr, and TiO₂ are 99.5%, 91.24%, 92.41%, and 94.86%, respectively.     

煤种对钛磁铁矿直接还原-磁选钛铁分离的影响

研究了烟煤和无烟煤对海滨钛磁铁矿直接还原-磁选钛铁分离的影响机理.结果表明,在试验用量范围内,两种煤对还原铁指标的影响规律相近,煤用量低时钛磁铁矿还原不充分.随煤用量增加,被还原的金属铁越来越多,但粒度较小,与其他颗粒嵌布紧密,因此还原铁Fe品位低,Ti O2品位高,铁回收率则先提高后基本不变.所有煤用量下所得金属铁颗粒均纯净.和无烟煤相比,烟煤固定碳较低,还原气氛较弱,但灰分较高,有利于金属铁颗粒的聚集长大;因此相同用量的烟煤为还原剂时,焙烧矿中金属铁颗粒较少,但粒度较大,还原铁中Fe品位较高,铁回收率较低,Ti O2品位较低.     

利用二元钠盐体系处理钒渣的清洁焙烧工艺

试验研究了利用二元钠盐(NaOH-Na₂CO₃)对钒渣进行焙烧,并分析了相关动力学参数对焙烧效果的影响.结果表明:在二元钠盐焙烧过程中,焙烧温度、焙烧时间及NaOH与Na₂CO₃质量比对渣中钒、铬的浸出率影响重大;焙烧过程中,Fe₃O₄被氧化为Fe₂O₃,V₂O₅和Cr₂O₃分别被氧化为β钒酸钠型结构的Na₃VO₄与正交晶系结构的Na₂CrO₄;最佳焙烧条件下,NaOH与 Na₂CO₃质量比为1.5∶1,焙烧温度为600℃,焙烧时间为60min,此时钒与铬的浸出率分别为98.66%与83.57%;浸出尾渣的主要金属元素为Fe.     

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.     

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。     

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.     

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%.     

高铝硅氰化渣中铁回收工艺

研究一种处理磁选前高铝硅氰化渣的新工艺。采用复合添加剂焙烧-水浸-磁选工艺对一种铁品位为27.69%(质量分数),SiO2含量为23.9%,Al2O3含量为6.35%的高铝硅氰化渣进行杂质与铁分离的研究。研究结果表明:在最佳焙烧条件下,当水浸温度为60℃,液固比为15:1,水浸时间为5 min,转速为20 r/min,在激磁电流为2 A时,可获得铁品位57.11%,铁的回收率为72.58%的铁精矿。铁的品位和回收率都比单纯的复合添加剂还原焙烧-磁选法所获得的铁精矿的指标高,铁的品位提高了10%左右,回收率提高了30%左右。X线荧光(XRF),X线衍射(XRD)及能谱(EDS)分析研究结果表明:经水浸后,复合添加剂焙烧过程中所产生的可溶性复杂杂质化合物被洗除,不溶性物质经磁选后随之进入非磁性物,实现铁与杂质矿物之间的有效分离。     

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.