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。     

从石煤酸浸液中萃取钒的工艺研究

研究了三正辛胺从石煤酸浸液中萃取钒的工艺过程 ,从萃取和反萃的 p H值、相比、有机相组成、澄清时间等方面进行了详细试验。研究表明 :用三正辛胺萃取钒时 ,其萃取率可达98%以上 ;而且易反萃 ,用 0 .5 M Na2 CO3反萃时 ,反萃率为 99.9%。经萃取后 ,浸出液中的钒可由每升几克富集到每升数十克以上 ,有利于后续的提钒工艺。     

含钒石煤矿高效清洁提取五氧化二钒

该研究采用XRD、SEM、BES、TG-DTG等检测手段,对多个地区(四川、贵州、甘肃、湖南)的含钒石煤进行了矿物学研究。研究表明,石煤中的主要矿物为石英、云母、粘土;各地石煤矿的化学成分、钒赋存状态、不同价态钒的分布差异较大,但钒主要赋存于铝硅酸盐矿物中。采用常压强化浸出工艺,对多个地区石煤进行了提取钒的工艺技术研究。结果表明,湖南、贵州、四川、甘肃等地石煤矿中钒浸出率可达80%、82%、62%、55%以上,含氟离子助浸剂的加入有利于含钒矿物的强化离解;钒浸出率与碳质成分的含量及二氧化硅含量有关。废酸回收率达到83%以上,钒离子和铁离子截留率分别达到93%~95%和92%~94%;钒萃取率和反萃率分别达60%和70%以上。酸浸渣用于制备建筑用陶粒和砖,产品达到优等品的等级要求。开展了日处理100kg含钒石煤矿连续浸出—连续萃取/反萃—精钒制备扩大实验,钒浸出率可达82%~83%,重现了小试结果;8级逆流萃取/反萃后,钒回收率达96.3%;产品五氧化二钒纯度为99.01%;全流程钒直收率达80%以上。研究了U、Th、Ra、K等4种放射性核素在全流程的走向及分布。结果表明,石煤中93.06%的U进入溶液。而经过萃取反萃后,98%左右的U在贫有机相中富集;而75%~77%的Th、Ra、K三种放射性核素在浸出过程中滞留于渣相中,在后续萃取/反萃过程中的分布与U是一致的。常压强化浸出—萃取工艺为实现石煤资源的高效清洁利用提供了一条可行的技术路线。     

高钙含钒煤灰渣提钒的工艺研究

本文阐述了高钙含钒煤灰渣提钒中,影响稀盐酸脱钙、焙烧、浸取、萃取,反萃取、沉钒的有关因素,并确定了最佳工艺条件。     

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.     

在压力场下从石煤中提取五氧化二钒的工艺

对贵州某地石煤进行加压酸浸提钒实验研究。在压力场条件下考察几种主要因素对钒浸出率的影响。得到的最佳技术条件为:反应时间3 h,硫酸质量浓度200 g/L,浸出温度180℃,搅拌转速580 r/min;在此条件下,钒浸出率(质量分数)为76%。两段逆流浸出实验结果表明:钒浸出率可达90%以上。浸出液经过废酸回收、还原、调整pH等预处理后,采用溶剂萃取的方法能够有效地分离和富集钒,钒萃取率可达98.1%,反萃率为99.14%;用氨水沉淀反萃液中的钒,沉淀物(多聚钒酸铵)在550℃煅烧3 h即可产出纯度为99%V2O5;全流程钒回收率为85%左右。     

碱法浸出石煤中的钒和硅

在石煤提钒工艺中,为了充分利用石煤中的有价元素硅,采用碱浸提钒工艺提取石煤中的钒和硅.经过预焙烧后,可以有效地破坏石煤结构,提高钒硅浸出率.在焙烧温度850℃、焙烧时间2h、浸出温度95℃、浸出时间4h、固液比(g∶mL)1∶1.4、矿碱质量比1.2∶1的条件下,钒的浸出率为86.6%,硅的浸出率为61.4%.     

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。     

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 %.     

直接酸浸含钒灰渣提钒的工艺研究

本文阐述了硫酸直接浸取含钒灰渣提钒工艺中,硫酸浸取,浸出液中提取铵明矾、再提取中间盐并溶解之,溶剂萃取、反萃及沉钒的有关因素。并确定了最佳工艺条件。     

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%。     

Synergistic effect of microwave irradiation and CaF<Subscript>2</Subscript> on vanadium leaching

The effect and mechanism of microwave irradiation on vanadium leaching were studied via a comparison between microwave heating and conventional heating. The results show a synergistic effect of microwave irradiation and calcium fluoride (CaF₂) on the vanadium leaching efficiency. It is confirmed that the vanadium leaching process can be improved by microwave irradiation when CaF₂ is present. The leaching rate of vanadium under microwave irradiation is increased by 8%-15% when 5wt% CaF₂ is added; by contrast, in the absence of CaF₂, the leaching rate is almost unaffected compared to that by conventional heating. Morphological analysis reveals that the particles are gradually eroded by acid under microwave irradiation, whereas some of the fine particles in samples subjected to conventional heating are tightly covered by a flocculent silicate product. Moreover, a large amount of Al and V and a small amount of Si are dissolved from samples under microwave heating, as revealed by the elemental analysis of leachates. Fourier transform infrared spectroscopic analysis also indicates a higher mass transfer coefficient in the diffusion layer of the raw material by microwave irradiation. When CaF₂ is present, the reaction energy barrier is lowered and the leaching process is controlled by the tightly covered product layer, resulting in a prominent effect of microwave irradiation.     

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%.     

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%。     

A new process of extracting vanadium from stone coal

A new process of extracting vanadium from the stone coal vanadium ore in Fangshankou, Dunhuang 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 fine 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 V₂O₅ 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.     

用废旧电路板酸浸-电沉积法回收金属铜

为了既保障环保又达到最佳铜回收效果的目的,以废旧电路板经过初步破碎分离,得到铜金属富集产品为研究对象,采用硫酸-双氧水浸出结合电沉积工艺回收物料中的铜。考察了铜浸出率与硫酸质量浓度、双氧水用量、浸出时间、液固比之间的关系,进行最佳金属浸出条件实验。结果表明：硫酸物质的量浓度为3.5 mol/L、浸出时间3 h、双氧水用量为20 mL、液固比10∶1条件下,铜金属浸出率最高可达到97.58%。电沉积尾液循环用以浸出铜,铜离子循环质量浓度不低于18 g/L,平均电流效率高于97.60%。在沉积温度40℃,阴极材料为T2紫铜,电流密度小于800 A/m2时,电流效率可以保持在97%以上。实验中铜回收率达到了97.58%。     

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.     

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.     

Microemulsion leaching of vanadium from sodium-roasted vanadium slag by fusion of leaching and extraction processes

The fusion of the leaching and purification processes was realized by directly using microemulsion as the leaching agent. The bis-(2-ethyhexyl) phosphoric acid(DEHPA)/n-heptane/Na OH microemulsion system was established to directly leach vanadates from sodium-roasted vanadium slag. The effect of the leaching agent on the leaching efficiency was investigated, in addition to the molar ratio of H_2O/Na DEHP(W), DEHPA concentration, solid/liquid ratio, stirring time, and leaching temperature. In optimal situations, the vanadium leaching efficiency reaches 79.57%. The X-ray diffraction characterization of the leaching residue and the Raman spectrum of the microemulsion before and after leaching demonstrate the successful entry of vanadates from the sodium-roasted vanadium slag into the microemulsion. The proposed method successfully realizes the leaching and purification of vanadates in one step, thereby greatly reducing production costs and environmental pollution. It also offers a new way to achieve the green recovery of valuable metals from solid resources.