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%。     

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。     

石煤氧化焙烧-酸溶液浸出提钒工艺

针对某地石煤原料组成特点,为提高石煤中钒的浸出率本试验采用氧化焙烧-酸溶液浸出提钒的工艺,研究了焙烧温度,添加剂氯化钠的用量,焙烧时间,酸浸出液用量及浸出温度时间对钒浸出率的影响.结果表明:,焙烧水浸后渣再用稀酸浸出的方法,钒的浸出率可提高10 %左右,钒的浸出率可达到75 %~80 %.     

钙化焙烧提钒废渣浸出毒性试验初步研究

对贵州省某矿业有限公司刚出炉的钙化焙烧提钒废渣进行浸出实验,并利用斑马鱼胚胎染毒实验探讨了钙化焙烧提钒废渣浸出液中重金属的联合毒性效应.实验结果表明:根据《固体废物浸出毒性浸出方法-水平振荡法》(HJ557-2010)试验规范进行浸出,钙化焙烧提钒废渣浸出液中V浓度符合《钒工业污染物排放标准》(GB26452-2011),其它7种重金属浓度均符合《污水综合排放标准》(GB8978-1996)中的排放限值;染毒后的斑马鱼胚胎没有表现出强烈的感应,孵化出的幼鱼表现正常.由此说明钙化焙烧提钒废渣浸出液毒性较小,不会表现出对环境的强烈污染效应.     

钛白废酸无焙烧直接浸出提钒短流程新工艺及装备研究

钒钛磁铁矿是我国主要提钒资源,广泛地分布于我国的攀枝花、承德地区。以钠化焙烧—水浸为代表的焙烧浸出工艺存在着污染环境、金属回收率不高的问题,目前正被逐步改进。据统计每生产1 t钛白粉就会排出20%的废酸8~10 t,而中国钛白行业年产废硫酸达到600万吨,直接排放将造成严重的环境污染。该报告围绕无焙烧直接加压酸浸提钒技术中的直接加压酸浸、浸出液中有价元素分离、新型加压连续浸出反应器研发、系统内物流循环与利用、浸出渣的综合利用、工艺放大等研究内容进行,通过相关研究取得以下成果:(1)研究并对比了无焙烧常压酸浸、无盐氧化焙烧常压酸浸、无焙烧氧压酸浸3种提钒过程的现象,结果表明:相比无焙烧常压酸浸、无盐氧化焙烧常压酸浸等2个工艺,明显地具有反应快速、高效的特点。(2)采用硫酸体系加压浸出四川攀枝花地区的转炉钒渣,矿物学表明,转炉钒渣中的主要物相为尖晶石相、钛铁矿相以及铁橄榄石。加压浸出过程中,铁橄榄石和尖晶石相逐渐分解,钒、铁被浸出进入浸出液,部分未反应的钛、硅相在浸出渣中富集。(3)对该技术核心加压酸浸过程进行了放大实验研究,对实验室研究结果进行了验证,放大实验研究结果表明:钒的浸出率随着初始酸度的增加而增加,随着液固比的增大而增大。在加压温度150℃,硫酸浓度300 g/L,搅拌转速300 rpm,浸出时间90 min,液固比8∶1的条件下,钒的浸出率可达到99.10%。(4)提钒酸浸液萃取最优工艺条件为:常温,还原剂用量20 g/L、浸出液p H=2.0、有机相组成为20%P2O4,5%TBP,75%磺化煤油、相比(O/A)=1∶1、震荡时间5 min,钒的一级萃取率达到74.49%,Fe的萃取率仅为1.92%。在最优条件下,进行4级错流萃取,钒的总萃取率可达97.89%。以硫酸为反萃液进行反萃,其最优工艺条件为:反萃时间t=4 min、反萃液浓度200 g/L、反萃相比(O/A)=1∶1时,钒的反萃率达到98%以上。     

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.     

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.     

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%.     

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.     

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.     

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.     

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.     

A method for recovery of iron,titanium, and vanadium from vanadium-bearing titanomagnetite

An innovative method for recovering valuable elements from vanadium-bearing titanomagnetite is proposed. This method involves two procedures:low-temperature roasting of vanadium-bearing titanomagnetite and water leaching of roasting slag. During the roasting process, the reduction of iron oxides to metallic iron, the sodium oxidation of vanadium oxides to water-soluble sodium vanadate, and the smelting separation of metallic iron and slag were accomplished simultaneously. Optimal roasting conditions for iron/slag separation were achieved with a mixture thickness of 42.5 mm, a roasting temperature of 1200℃, a residence time of 2 h, a molar ratio of C/O of 1.7, and a sodium carbonate addition of 70wt%, as well as with the use of anthracite as a reductant. Under the optimal conditions, 93.67% iron from the raw ore was recovered in the form of iron nugget with 95.44% iron grade. After a water leaching process, 85.61% of the vanadium from the roasting slag was leached, confirming the sodium oxidation of most of the vanadium oxides to water-soluble sodium vanadate during the roasting process. The total recoveries of iron, vanadium, and titanium were 93.67%, 72.68%, and 99.72%, respectively.     

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。