铜渣中铁铜分离新工艺研究

本文综述了铁铜分离的一般方法,并根据国内外氯化焙烧现状,结合铜渣特点,提出氯化焙烧法分离铜渣中铁铜的优势:可以在中温条件下进行选择性氯化,完成铁铜的分离过程;氯化焙烧过程是在氧化性气氛下进行的,铜渣中硫元素可以SO2的形式溢出;铜渣中的铜以Cu Cl2形式挥发,可以在反应器冷端凝结,铁留在渣中,实现了铜渣中铁铜的共同回收。     

铜渣氯化焙烧脱铜机理和动力学研究

采用氯化焙烧可有效实现铜渣中含铜、铁物相CuO和Fe3O4的选择性氯化挥发分离,氯化剂种类不同铜氯化挥发过程特异性明显.FeSO4·7H2O为添加剂,NaCl和CaCl2为氯化剂分别对CuO进行氯化焙烧时,物料失重可分为主要失水阶段、微量失水阶段和氯化失重阶段三部分,其中氯化失重部分反应属一级反应.CaCl2焙烧起始反应温度较NaCl较低为315.9℃,且其焙烧反应活化能4.826KJ/mol低于NaCl 6.001KJ/mol.影响CuO氯化挥发效果的主要因素为焙烧温度,焙烧温度1130℃、焙烧时间60min、氧气流量0.2L/min、CaCl2加入量1.6(CaCl2与铜渣质量比)和FeSO4·5H2O加入量4.42(FeSO4与铜渣质量比)条件下,焙烧物料失重率最大达62.3%,基本可实现物料中铜的氯化挥发,同时铁氯化损失率较小,可实现铜铁氧化物的选择性氯化挥发分离.     

诺兰达铜渣中有价元素的回收

采用电感耦合等离子体发射光谱仪、光学显微镜、X射线衍射仪等对诺兰达铜渣的化学成分、微观结构和物相组成进行了研究,并采用"焙烧-浸出"工艺探索铜渣资源化利用的途径.结果表明,诺兰达铜渣的主要矿相为铁橄榄石、磁铁矿、铜锍、玻璃相和少量金属铜.在680℃焙烧4 h后,采用2mol/L硫酸于70℃浸出2h后可以达到良好的浸出效果,使铜的浸出率达到92%以上.     

焙烧-浸出法分离回收白合金中的有价金属

白合金是铜钴矿冶炼过程的中间产物,存在大量钴、铜、铁金属元素,其中酸性溶液中的铁、钴分离是白合金湿法回收的难点之一。采用焙烧-浸出法分离回收白合金酸浸生成的硫酸铁-硫酸钴混合溶液。通过设计单因素实验探索最佳工艺条件,采用X射线衍射(XRD)以及扫描电子显微镜(SEM)研究白合金有价金属分离过程中的物相变化。研究结果表明：在硫酸浓度为1.9 mol/L、浸出温度为90℃、浸出时间为60 min和液固比为5:1 (L/g)的最佳条件下,白合金中钴和铁的浸出率达到91.78%和98.67%,而铜在硫酸溶液中保持稳定的斜方蓝辉铜矿结构残留在滤渣(含铜渣)中;随后将滤液浓缩结晶得到水合硫酸铁-硫酸钴混合晶体,并在600℃下焙烧,使硫酸铁热分解成难溶于水的氧化铁,而硫酸钴保持稳定,经水溶浸出后测得钴、铁浸出率为98.5%和0.1%,达到理想分离效果;最后将含铜渣在450℃下焙烧,形成硫酸铜-氧化铜化合物,经过酸浸可完全回收有价金属铜。     

含钼铜精矿氧化焙烧-浸出分离钼研究

四川攀西地区的含钼铜精矿中,由于钼、铜矿物组成复杂,共生关系紧密,提出了氧化焙烧-浸出工艺进一步分离钼。将试样置入焙烧炉中进行氧化焙烧,使硫化物转化为CuO、MoO3、Fe2O3等氧化物后;添加NaOH、H2O与MoO3反应生成可溶性Na2MoO4,浸出渣作为铜精矿产品。研究结果表明:铜、钼等以氧化物形式赋存于焙烧矿中,氧化焙烧矿中的硫含量较低为0.53%,硫以SO2气相形式挥发;在焙烧温度650℃、焙烧时间120 min、氢氧化钠用量为30%、浸出温度60℃、浸出时间120min、浸出液固比2∶1的综合条件下,钼的浸出率为94.24%,铜精矿(浸出渣)中铜的质量分数为24.27%,钼分离效果明显。     

氰化—萃取法从湿法处理铜阳极泥尾渣中回收有价金属

目前国内处理铜阳极泥的流程可归纳为四类。一类是传统的火法处理流程:将阳极泥经脱铜、脱硒、还原熔炼、氧化精炼,然后电解精炼得成品金银。第二类是选冶联合流程:先用氯化钠作氧化剂,浸出铜硒,后用选矿方法富集得银精矿,然后一次熔炼成金银合金阳极板,再电解精炼得成品金银。第三类是半湿法流程:阳极泥经硫酸化焙烧脱硒;稀硫酸浸铜;脱铜渣氨浸分银,水合肼还原沉淀银粉,铸阳极电解;分银渣硝酸分铅后再氯化分金,SO_2还原金粉,铸阳极板电解。第四类是全湿法流程:先将阳极泥水洗,过筛,过滤,然后用稀硫酸浸出脱铜,脱铜渣氯化浸出,氯化液有机萃取,草酸还原得海绵金,或氯化液用二氧化硫或草酸直接还原获粗     

某难处理多金属矿石综合浸出铀、铜、银试验研究

针对某难浸铀矿石,采用“氯化焙烧-硫酸浸出”工艺进行处理提取铀、铜、银。研究结果表明,最佳氯化焙烧实验条件为氯化钠用量6%,氯化焙烧温度 460 ℃,氯化焙烧时间2 h,焙烧液固比0.2∶1。对氯化焙烧后的矿样进行硫酸浸出,浸出条件为:硫酸浓度30 g/L、浸出时间30 min、浸出温度70 ℃、液固比2∶1,此时金属离子铀、铜、银的浸出率分别为铀85.08%、铜95.82%、银91.80%。     

难选锡中矿球团还原氯化挥发焙烧的研究

本文研究了锡中矿还原氯化挥发焙烧过程中,氯化温度、时间、气氛、还原剂和氯化剂用量,球团碱度和添加剂等因素对炉料中锡、铅、铜、锌和砷等挥发率的影响,并给出了焙烧的最佳条件。通过体系中有关化学反应的热力学计算、对金属氯化挥发的机理进行了分析。     

铜氨溶液电积铜的研究

工业铜渣和软锰矿在硫酸介质中经氧化还原,得到硫酸铜和硫酸锰混合溶液,净化过滤后,用氨与碳酸氢铵混合溶液分离锰得到的铜氨溶液,通过控制电解技术参数电解制得铜。研究表明在50℃、电流密度为150 A/m2条件下可以得到2#阴极铜。     

高炉灰为还原剂对海滨钛磁铁矿直接还原焙烧磁选--钛铁分离的影响

对高炉灰在直接还原焙烧-弱磁选工艺中用作印尼某海滨钛磁铁矿还原剂的可行性及其机理进行研究.结果表明,以萤石为添加剂的条件下,高炉灰可代替煤做还原剂,通过高炉灰与萤石的共同作用,可以在直接还原过程中提高还原铁粉中铁的回收率及品位并降低TiO2质量分数,同时回收高炉灰中铁.三种不同产地高炉灰还原效果的比较表明,高炉灰性质对还原效果有影响.在相同用量条件下,津鑫高炉灰( JX)还原效果最好;在JX高炉灰用量30%、萤石用量10%、焙烧温度1250益以及焙烧时间为60 min时,焙烧产物通过两段磨矿和两段磁选,最终得到最佳的还原铁粉中铁品位为91.28%,TiO2质量分数降至0.93%,包括海滨砂矿和高炉灰中铁的铁总回收率达到89.19%.     

金川镍闪速熔炼渣的物相与铜镍分布

采用X射线衍射、光学显微镜、扫描电镜和电子探针微区分析等方法,研究金川镍内速熔炼水淬渣的主要物相组成和有价金属铜、镍在炉渣中的存在形式。结果表明,水淬渣的主要结晶物相为铁镁橄榄石,结晶相之间由玻璃相填充,细小的铜镍铁硫化物以星散状分布于炉渣中,有价金属铜、镍主要以硫化物形态存在于炉渣中,还有部分存在于铁镁橄榄石相中。     

A novel process for Fe recovery and Zn,Pb removal from a low-grade pyrite cinder with high Zn and Pb contents

Comprehensive utilization of pyrite cinders is increasingly important because of their huge annual outputs and potential valuable metals recovery to cope with the gradual depletion of high-grade mineral resources. In this work, a new process, i.e., a high-temperature chlorination-magnetizing roasting-magnetic separation process, was proposed for recovering Fe and removing Zn, Pb from a low-grade pyrite cinder containing 49.90wt% Fe, 1.23wt% Zn, and 0.29wt% Pb. Various parameters, including the chlorinating conditions (dosage of CaCl₂, temperature, and time) and the magnetization roasting conditions (amount of coal, temperature, and time) were investigated. The results indicate that the proposed process is effective for Fe recovery and Zn, Pb removal from the pyrite cinder. Through this process, 97.06% Zn, 96.82% Pb, and approximately 90% S can be removed, and 89.74% Fe is recovered as magnetite into the final product under optimal conditions. A purified magnetite concentrate containing 63.07wt% Fe, 0.16wt% P, 0.26wt% S, and trace amounts of nonferrous metals (0.005wt% Cu, 0.013wt% Pb, and 0.051wt% Zn) was obtained. The concentrate can be potentially used as a high-quality feed material for producing oxidized pellets by blending with other high-grade iron ore concentrates.     

金川镍沉降渣的工艺矿物学

采用化学分析、X射线衍射(XRD)、光学显微镜、扫描电子显微镜(SEM)和能谱(EDS)等方法,研究了金川镍沉降渣的矿物组成、结构、嵌布特征、主要有价成分Fe、Ni、Cu、Co的分布等工艺矿物学性质.结果表明,金川镍沉降渣主要由铁镁橄榄石和玻璃质组成,并含少量的铜镍铁硫化物、辉铜矿、磁铁矿等;沉降渣的结构单一,微细粒的铜镍铁硫化物呈星散状无规律分散在硅酸盐基质中;铁主要存在于铁镁橄榄石内,镍和铜主要赋存在铜镍铁硫化物中,钴没有独立矿物存在,主要以类质同象形式赋存在其他矿物中.镍渣中有价成分的回收可考虑用深度还原法或湿法冶金工艺.     

Preparation of ferronickel from nickel laterite via coal-based reduction followed by magnetic separation

The sticking phenomenon between molten slag and refractory is one of the crucial problems when preparing ferronickel from laterite ore using rotary hearth furnace or rotary kiln processes. This study aims to ameliorate sticking problems by using silicon dioxide (SiO₂) to adjust the melting degree of the briquette during reduction roasting. Thermodynamic analysis indicates that the melting temperature of the slag gradually increases with an increase in the SiO₂ proportion (SiO₂/(SiO₂ + Al₂O₃ + MgO) mass ratio). Experimental validations also prove that the briquette retains its original shape when the SiO₂ proportion is greater than 75wt%, and sticking problems are avoided during reduction. A ferronickel product with 8.33wt% Ni and 84.71wt% Fe was prepared via reductive roasting at 1500℃ for 90 min with a SiO₂ proportion of 75wt% and a C/O molar ratio of 1.0 followed by dry magnetic separation; the corresponding recoveries of Ni and Fe reached 75.70% and 77.97%, respectively. The microstructure and phase transformation of reduced briquette reveals that the aggregation and growth of ferronickel particles were not significantly affected after adding SiO₂ to the reduction process.     

铜冶炼闪速炉烟尘氧化浸出与中和脱砷

介绍了废酸氧化浸出铜冶炼闪速炉烟尘和漫出液中和沉淀砷、铁过程。从化学热力学和实验2方面研究了浸出液中以砷酸铁形式中和沉淀脱砷过程，并对砷酸铁沉淀的稳定性进行了研究。研究结果表明：闪速炉烟尘中铜、砷和铁的浸出率分别可达到83％，92％和30％，浸出液中的铁和砷的量比n（Fe）／n（As）约为1．50；控制适当的pH值中和沉淀砷、铁，可使铜存留于溶液中，而砷以砷酸铁形式进入固相中，从而达到铜、砷分离的目的；不稳定的砷酸铁沉淀物进一步转型后，则可作为无毒稳定渣丢弃。     

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.     

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.     

Extraction of metals from complex sulfide nickel concentrates by low-temperature chlorination roasting and water leaching

The recovery of valuable metals from complex sulfide concentrates was investigated via chlorination roasting followed by water leaching. A reaction process is proposed on the basis of previous studies and the results of our preliminary experiments. During the process, various process parameters were studied, including the roasting temperature, the addition of NH₄Cl, the roasting time, the leaching time, and the liquid-to-solid ratio. The roasted products and leach residues were characterized by X-ray diffraction and vibrational spectroscopy. Under the optimum condition, 95% of Ni, 98% of Cu, and 88% of Co were recovered. In addition, the removal of iron was studied in the water leaching stage. The results demonstrate that this process provides an effective approach for extracting multiple metals from complex concentrates or ores.     

机械合金化AlCuFe 形成准晶体的XAFS 定量分析

为了准确辨别AlCuFe准晶体（QC）形成的相变过程,进而探索其形成机理,采用X射线吸收精细结构（XAFS）定量分析技术,计算不同球磨时间 Al70Cu20Fe10合金中的Cu原子,以及700 ℃退火后的Cu和Fe原子的局域结构参数,并与X射线衍射（XRD）分析进行比对。结果表明,经过不同球磨时间,样品中Al和Cu首先形成Al2Cu金属间化合物,进而转化为Cu9Al4,Fe原子仍然是原来的体心立方(bcc）α-Fe结构。α-Fe与Al和Al2Cu经过退火处理化合为Al7Cu2Fe化合物,进而转变为QC;而长时间球磨产生的Cu9Al4退火后化合为稳定的Al(Cu,Fe)固溶体,不会形成QC。