砷滤饼制备砷酸铜的影响因素及工艺优化

采用氢氧化钠溶液浸出硫化砷滤饼,有效实现As与Cu和Bi等金属的分离,对浸出液经氧化脱硫后配入黑铜泥的酸性浸出液制备砷酸铜.研究结果表明:当NaOH的浓度为1.5mol/L、液固比10∶1、反应温度70℃、反应时间1.5h、搅拌速度为400r/min时,硫化砷滤饼中As的浸出率达到96.56%,Cu浸出率仅为0.12%;经氢氧化钠浸出,浸出渣中Cu、Bi的质量分数分别从原来15.93%和1.96%增加到56.31%和6.92%,使Cu和Bi得到高度富集;所制备的砷酸铜w(Cu)>29.8%,w(As)>26.19%,砷酸铜的结构式为Cu5As4O15.9H2O.     

高铁闪锌矿精矿加压浸出半工业试验研究

高铁闪锌矿经二段加压浸出，锌浸出率97.09%，铁浸出率15.2%，元素硫产率93%，浸出液残酸3.95g/L，经简单中和、净化，生产合格硫酸锌新液供应锌电积。为了研究比较，闪锌矿经一段加压浸出，锌的浸出率98.09%，铁浸出率37.88%，元素硫产率96%，浸出液残酸高（46.4g/L），可并入现有传统湿法炼锌工艺流程处理。     

用硫酸亚铁浸出同时沉淀铁矾法处理低品位锰矿

研究了用硫酸亚铁浸出同时沉淀黄钠铁矾的方法处理低品位软锰矿的过程．在该过程中,软锰矿中的ＭｎＯ２被还原成ＭｎＳＯ４同时ＦｅＳＯ４被氧化并以黄钠铁矾的形态沉淀．沉淀产生的酸可直接用于ＭｎＯ２浸出．考察了硫酸、硫酸亚铁和硫酸钠的加入量及温度等参数对锰浸出和沉铁效率的影响．讨论了过程动力学．实验结果表明,锰浸出率和沉铁效率（质量分数）在最佳条件下可分别达到９６％和９２％．     

硫酸铜结晶母液制备砷酸铜的工艺技术

采用二段中和法研究了硫酸铜结晶母液制备砷酸铜的工艺过程,用XRD图探讨了砷酸铜的分子结构式．结果表明只要控制一定的pH值、反应时间、温度和铜砷质量比,便可以产出合格的砷酸铜,砷酸铜的结构式为Cu5AS4O15·9H2O．利用含砷物料制备砷酸铜对于环境保护具有重要意义．     

氯化物中黄钠铁矾除铁的热力学与动力学

对氯化物体系中锰结核浸出溶液采用黄钠铁矾净化除铁的热力学与动力学进行了研究．热力学分析表明，氯化物溶液中氯离子浓度升高能增大黄钠铁矾的溶解度，但在一般情况下（如ｃＣｌ－＜１０ｍｏｌ／Ｌ），完全可以将溶液中的铁以铁矾形式降到理想的程度；黄钠铁矾生成的最佳ｐＨ值范围为１．０～３．０；溶液中钠离子浓度升高有利于铁矾沉淀．动力学研究表明，氯化物中黄钠铁矾形成的表观活化能为９４．６６ｋＪ／ｍｏｌ，沉矾过程受化学反应控制．这些研究结果对锰结核浸出液中铁的分离以及氯化物体系中铁矾的沉淀具有指导作用．     

苛性碱溶液氧压浸出高砷锑烟尘

对苛性碱溶液氧压浸出高砷锑烟尘进行研究,考察NaOH浓度、O2分压、温度、浸出时间和液固质量比等因素对砷、锑和铅浸出行为的影响。研究结果表明:在火法处理铅阳极泥产出的高砷锑烟尘中,砷、锑的主要物相分别为As2O3和Sb2O3,锑也有少量以Sb2O5存在;在苛性碱溶液氧压浸出高砷锑烟尘过程中,As(Ⅲ)氧化为溶解度更大的As(Ⅴ)进入溶液,同时Sb(Ⅲ)氧化为Sb(Ⅴ),并形成难溶化合物Sb2O3·2Sb2O5、Pb2Sb2O7和NaSb(OH)6,进入浸出渣中;实验确定的最佳工艺条件为:NaOH质量浓度40 g/L,O2分压2.0 MPa,浸出温度140℃,浸出时间2 h,液固质量比10。在此条件下,As的浸出率可达95%以上,而Sb和Pb的浸出率均小于1.0%。     

高砷烟尘酸性氧化浸出砷和锌的试验研究

采用酸性氧化浸出工艺对某冶炼厂高砷烟尘进行湿法浸出砷、锌的试验研究。通过单因素试验确定最佳浸出工艺条件。结果表明,采用pH值为2的稀硫酸溶液,在浸出温度80℃、浸出时间105min、液固体积质量比10∶1、H2O2添加量1.75mL/g(烟灰)、搅拌速度705r/min的条件下,砷、锌浸出率分别达到78.25%和85.42%。     

次氯酸钠一步法从磷酸富砷渣中浸砷的工艺研究

为解决传统浸砷工艺的浸出率低和操作时间长等问题,利用次氯酸钠溶液中OH-的溶砷作用和ClO-的氧化脱硫功能,用次氯酸钠溶液作浸出剂,研究磷酸富砷渣中砷浸出时次氯酸钠溶液用量、浸出温度以及浸出时间对砷浸出率和浸出液中AsS33-氧化脱硫程度的影响规律。结果表明,与传统的碱浸和空气氧化浸出工艺相比,次氯酸钠溶液一步法的浸出时间由原来的十多个小时缩短到10 min。在次氯酸钠的用量为理论用量的4倍、浸出时间为10 min、浸出温度为30℃时,砷浸出率达98.52%,浸出液中AsS33-的硫被完全氧化脱出,实现了砷高效浸出和硫完全脱出的一步法工艺。     

氯盐浸出湿法炼铜的物理化学(Ⅳ) Ⅳ高砷铜精矿的密闭浸出

绪言现有炼铜流程处理高砷铜矿时,砷广泛地分散于各产品、半产品、烟尘、废水及废渣中,既不能脱除,也不易回收,不但会使制酸的钒催化剂活性下降,而且使环境受到严重污染。因此,国内外有关单位对高砷铜矿的处理均在进行试验研究。如美国矿务局采用氧压二氯化铁浸出黝铜矿,加拿大国际镍公司采用氧压浸出高砷铜精矿,西德、奥地利采用氧压酸浸高砷高锑铜矿。其共同特点是氧压浸出,浸出结果铜的浸出率可达98％,As、Sb、Fe等不被浸出,几乎全部进入残渣。     

湿法炼锌浸出渣的处理

研究了常规搅拌浸出及机械活化浸出方式下，温度、酸度及浸出时间等对锌焙砂酸浸渣中锌、铁浸出率的影响，考查了铁酸锌的浸出行为．试验结果表明，提高温度及酸度有利于酸浸渣中锌的浸出；机械活化浸出可明显改善铁酸锌的浸出行为，提高锌的浸出率，并改善锌、铁选择性浸出分离的效果，相同条件下，锌的浸出率可比常规搅拌浸出提高１６％～２５％．     

低品位硫铜钴矿生物浸出液中铜的分离

生物氧化法处理低品位铜钴硫化矿时,浸出液常含有高浓度的铁、低浓度的钴及一定量的铜,因此在回收钴前对其中的铜进行选择性分离提取,并避免钴的损失.采用萃取剂LIX984N选择性分离低品位硫铜钴矿生物浸出液中的铜.结果表明,当采用LIX984N体积分数为25%的有机相,在环境温度为35℃,相比为1∶1时,混合时间为5min,平衡pH值为125的条件下,可达到994%的铜萃取率.该条件下铁夹带仅为403%,钴共萃率0849%.对负载有机相采用中性水在相比1∶1的条件下洗涤,使钴和铁夹带分别降至0008%和0766%.洗涤后,负载有机相采用200g/L硫酸水溶液反萃,当有机相与水相体积比为1∶1时,经过2级逆流反萃时,铜反萃率达到9813%.     

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 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%.     

铅碱性精炼废渣制取三氧化二锑

研究了脆硫锑铅矿精矿与铅碱性精炼废渣同时浸出制取三氧化二锑的工艺流程,得出了浸出、还原、水解、中和等过程的最优工艺条件.该工艺的技术特点是在浸出过程中,铅碱性精炼锑酸钠渣与脆硫锑铅矿精矿互作氧化剂和还原剂.实验结果表明浸出过程中Sb浸出率为94.56%,Pb入渣率为97.43%,很好地实现了Sb与Pb的分离;浸出液经还原后,冲稀水解率达99.55%;经碱液中和,得到的三氧化二锑颜色呈白色,且其化学成分平均含量中,Sb2 O3为97.69%,As为0.0055%,Pb为0.0034%,As和Pb含量低,在用等离子体法制取超细氧化锑时可作为原料.该工艺具有综合利用程度高、环境污染小、易于实现工业化生产等优点,对于铅碱性精炼废渣的资源化利用,消除因其堆存造成的环境污染,具有十分重要的意义.     

从磷酸废砷渣中回收砷的碱浸工艺研究

根据目前广西区大量砷渣得不到有效利用的现状,以磷酸净化过程中产生的含砷废渣为原料,通过物相分析确定了碱浸出法回收砷的工艺,考察了浸出温度、摩尔比、液固比和反应时间等工艺条件对砷浸出率的影响。结果表明,n(NaOH)∶n(As2S3)是影响砷浸出率的主要因素,较适宜的碱浸工艺条件为:浸出温度为70℃,n(NaOH)∶n(As2S3)=6.0∶1,液固比=6.0∶1,反应时间为30 min,在此条件下砷浸出率可达97.1%。在现有基础上,该工艺为磷酸废砷渣的综合利用提供了一条简单高效的技术路线。     

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.     

复杂高硫渣中有价金属的分离工艺

采用煤油脱硫-氯盐浸出-分步水解法对复杂高硫渣中有价金属的分离进行研究.研究反应时间、反应温度、液固比等因素对实验过程的影响.结果表明:在反应温度为95 ℃,反应时间为0.5 h,液固比分别为11-1时进行2次连续煤油脱硫实验,硫的脱除率为98%,脱硫渣中铋和锑富集,其含量约为复杂高硫渣的6倍.在硫酸质量浓度和氯化钠质量浓度均为150 g/L,液固比为10-1,反应温度为65 ℃时,锑的浸出率为96%,铋的浸出率为98%.采用分步水解,在氯盐浸出液中控制pH=0.8水解沉锑;在沉锑后液中控制pH=1.5水解沉铋,锑和铋的沉淀率分别为85.6%和98%.在整个优化工艺条件下,锑的回收率为82%,铋的回收率为96%.     

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.     

Influence of acid leaching and calcination on iron removal of coal kaolin

Calcination and acid leaching of coal kaolin were studied to determine an effective and economical preparation method of calcined kaolin. Thermogravimetric-differential thermal analysis (TG-DTA) and X-ray diffraction (XRD) demonstrated that 900℃ was the suitable temperature for the calcination. Leaching tests showed that hydrochloric acid was more effective for iron dissolution from raw coal kaolin (RCK), whereas oxalic acid was more effective on iron dissolution from calcined coal kaolin (CCK). The iron dissolution from CCK was 28.78wt%, which is far less effective than the 54.86wt% of RCK under their respective optimal conditions. Through analysis by using M?ssbauer spectroscopy, it is detected that nearly all of the structural ferrous ions in RCK were removed by hydrochloric acid. However, iron sites in CCK changed slightly by oxalic acid leaching because nearly all ferrous ions were transformed into ferric species after firing at 900℃. It can be concluded that it is difficult to remove the structural ferric ions and ferric oxides evolved from the structural ferrous ions. Thus, iron removal by acids should be conducted prior to calcination.