锌浸渣还原焙烧-磁选回收铁

在查明锌浸渣工艺矿物学的基础上,采用还原焙烧将铁酸锌分解为氧化锌和磁性氧化铁,再通过磁选的方法回收铁,达到锌、铁分离的目的。实验考查了焙烧温度、焙烧时间、还原剂用量对铁酸锌分解率、铁回收率和铁品位的影响。结果表明:在焙烧温度为950℃、焙烧时间为1 h及还原剂添加量为10%和5%的条件下,铁酸锌分解率达到72.05%,铁回收率可达到91.79%,精矿中铁的品位为50%左右。焙烧及磁选过程中颗粒的团聚包裹是铁精矿品位不高的主要原因。     

从柿竹园铅锌矿中提取铁和锌粉末

以柿竹园铅锌矿为原料，煅烧除硫后进行酸洗，实现了铅锌与其它物质的分离。通过加入1mol／LNaOH来控制pH=5．0和pH=7．5，分别得到铁和锌的氢氧化物，高温煅烧后，得到其氧化物。利用SEM和X-射线能谱分析及XRD相分析，相应地对制备的溶液和滤渣进行了表征，研究了焙烧物料的数量、焙烧温度和时间、用于酸洗的硝酸浓度等对产物的纯度和回收率的影响。本试验通过控制适宜的工艺条件，使铁和锌的酸性浸出率分别达到85．9％和98．8％，实现了资源的综合利用。     

Synchronous Volatilization of Sn,Zn, and As,and Preparation of Direct Reduction Iron (DRI) from a Complex Iron Concentrate via CO Reduction

Sn-, Zn-, and As-bearing iron ores are typical complex ores and are abundantly reserved in China. This kind of ore is difficult to use effectively due to the complicated relationships between iron and the other valuable metal minerals. Excessive Sn, Zn, and As contents would adversely affect ferrous metallurgy operation as well as the quality of the products. In this study, thermodynamic calculations revealed that it was feasible to synchronously volatilize Sn, Zn, and As via CO reduction. Experimental results showed that preoxidation was necessary for the subsequent reductive volatilization of Zn from the pellets, and the proper preoxidation temperature was 700–725°C under air atmosphere. Synchronous volatilization of Sn, Zn, and As was realized by roasting under weak reductive atmosphere after the pellets were preoxidized. The volatilization ratios of 75.88% Sn, 78.88% Zn, and 84.43% As were obtained, respectively, under the conditions by reduction at 1000°C for 100 min with mixed gas of 50% CO + 50% CO₂ (in vol.). A metallic pellet (direct reduction iron) with total iron grade of 87.36%, Fe metallization ratio of 89.27%, and residual Sn, Zn, and As contents of 0.071%, 0.009%, and 0.047%, respectively, was prepared. Sn and As were mainly volatilized during weak reductive atmosphere roasting, and those volatilized in the metallization reduction process were negligible. Most of Zn (78.88%) was volatilized during weak reductive atmosphere roasting, while the metallization reduction process only contributed to 16.10% of total Zn volatilization.     

Application of suspension magnetization roasting as technology for high-efficiency separation of valuable iron minerals from high-iron bauxite

A technology for suspension magnetization roasting?magnetic separation was proposed to separate iron minerals for recovery. The optimum parameters were as follows: a roasting temperature of 650 °C, a roasting time of 20 min, a CO concentration of 20%, and particles with a size less than 37 μm accounting for 67.14% of the roasted product. The total iron content and iron recovery of the magnetic concentrate were 56.71% and 90.50%, respectively. The phase transformation, magnetic transition, and microstructure evolution were systematically characterized through iron chemical phase analysis, X-ray diffraction, vibrating sample magnetometry, X-ray photoelectron spectroscopy, and transmission electron microscopy. The results demonstrated the transformation of hematite to magnetite, with the iron content in magnetite increasing from 0.41% in the raw ore to 91.47% in the roasted product.     

Simultaneous extraction of gold and zinc from refractory carbonaceous gold ore by chlorination roasting process

A novel process based on chlorination roasting was proposed to simultaneously recover gold and zinc from refractory carbonaceous gold ore by using NaCl as chlorination agent. The effects of roasting temperature, roasting time and NaCl content on the volatilization rates of gold and zinc were investigated. The reaction mechanism and the phase transition process were also analyzed by means of SEM, EDS and XRD. The results demonstrated that under the optimal conditions of NaCl content of 10%, roasting temperature of 800 °C, roasting time of 4 h and gas flow rate of 1 L/min, the rates of gold and zinc were 92% and 92.56%, respectively. During low-temperature chlorination roasting stage, a certain content of sulfur was beneficial to the chlorination reactions of gold and zinc; and during high-temperature chlorination roasting stage, the crystal structure of vanadium-bearing mica was destroyed, and the vanadium-containing oxides were beneficial to the chlorinating volatilization of gold and zinc. Eventually, the chlorinated volatiles of gold and zinc could be recovered by alkaline solution.     

Recovery of valuable metals from zinc leaching residue by sulfate roasting and water leaching

Zinc leaching residue (ZLR), produced from traditional zinc hydrometallurgy process, is not only a hazardous waste but also a potential valuable solid. The combination of sulfate roasting and water leaching was employed to recover the valuable metals from ZLR. The ZLR was initially roasted with ferric sulfate at 640 °C for 1 h with ferric sulfate/zinc ferrite mole ratio of 1.2. In this process, the valuable metals were efficiently transformed into water soluble sulfate, while iron remains as ferric oxide. Thereafter, water leaching was conducted to extract the valuable metals sulfate for recovery. The recovery rates of zinc, manganese, copper, cadmium and iron were 92.4%, 93.3%, 99.3%, 91.4% and 1.1%, respectively. A leaching toxicity test for ZLR was performed after water leaching. The results indicated that the final residue was effectively detoxified and all of the heavy metal leaching concentrations were under the allowable limit.     

Separation of arsenic and antimony from dust with high content of arsenic by a selective sulfidation roasting process using sulfur

The separation of arsenic and antimony from dust with high content of arsenic was conducted via a selective sulfidation roasting process. The factors such as roasting temperature, roasting time, sulfur content and nitrogen flow rate were investigated using XRD, EPMA and SEM–EDS. In a certain range, the sulfur addition has an active effect on the arsenic volatilization because the solid solution phase ((Sb,As)₂O₃) in the dust can be destroyed after the Sb component in it being vulcanized to Sb₂S₃ and this generated As₂O₃ continues to volatile. In addition, an amorphization reaction between As₂O₃ and Sb₂O₃ is hindered through the sulfidation of Sb₂O₃, which is also beneficial to increasing arsenic volatilization rate. The results show that volatilization rates of arsenic and antimony reach 95.36% and only 9.07%, respectively, under the optimum condition of roasting temperature of 350 °C, roasting time of 90 min, sulfur content of 22% and N₂ flow rate of 70 mL/min. In addition, the antimony in the residues can be reclaimed through a reverberatory process.     

Mineralogical changes occurring during the fluid-bed roasting of zinc sulfide concentrates

During the fluid-bed roasting of zinc sulfide concentrates, the sulfur in the sphalerite (Zn,Fe)S diffuses out of the particles, whereas the associated zinc and iron are converted to (Zn,Fe)O. The iron from the (Zn,Fe)O phase migrates outward to the peripheries of the particles, forming ZnFe₂O₄. The resulting porous ZnO+ZnFe₂O₄ particles agglomerate to form large, spherical masses. A compact shell subsequently develops on the surface of the agglomerates and continues to grow inward; eventually, the particles become compact. Rhythmic bands consisting of intergrowths of ZnO, ZnFe₂O₄, Zn₂SiO₄, and lead oxide/oxysulfate are often present in the agglomerated masses, and the Zn₂SiO₄ and lead oxide/oxysulfate phases seem to form by a vapor phase reaction. Occasionally, defluidization occurs during roasting. The defluidization agglomerates consist of calcine particles that are cemented by zinc oxysulfate and zinc sulfate. For more information, contact T.T. Chen, CANMETMMSL, 555 Booth Street, Ottawa, Canada K1A 0G1; (613) 995-9490; fax (613) 996-9673; e-mail tchen@nrcan.gc.ca     

流动注射-氢化物发生原子吸收分光光度法测定钢铁中微量铋

采用流动注射-氢化物发生原子吸收分光光度法测定钢铁中微量铋含量,测量范围是:0～0.01％,方法精密度考核结果为在RSD＜2％.用标准样品对照分析,结果令人满意.方法的选择性好,准确度高,是实验室当前痕量分析的重要手段之一,能够满足日常检验的需要.     

高砷铅阳极泥水蒸气焙烧脱砷实验研究

针对目前工业上难处理的高砷铜阳极泥 ,作者提出在水蒸气气氛中焙烧脱砷的新工艺。实验主要对反应气氛、焙烧温度、焙烧时间等影响因素进行了系统考查。结果表明 ,在水蒸气气氛下焙烧高砷铅阳极泥 ,脱砷率≥ 87% ,焙砂含砷 <3% ,脱砷效果明显好于空气气氛。同时 ,通过XRD分析 ,对焙烧脱砷过程中的物相变化及反应机理进行了探讨。     

Sulphatizing roasting of Cu-Zn and multimetal complx slphides

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次氧化锌中砷的物相及砷的浸出(英文)

研究烟化炉次氧化锌中砷的物相类型。结果表明:按砷的物相可将次氧化锌分为3种类型。在一型次氧化锌中砷以As2O3形态存在,而在二型和三型次氧化锌中砷分别以亚砷酸锌(Zn(AsO2)2)和砷酸铅(Pb(As2O6),Pb4As2O9)形态存在。在热力学分析基础上,对二型次氧化锌进行浸出脱砷。结果表明:采用30g/LNaOH溶液,在液固比3、温度20°C的条件下,砷的浸出率在1h内可达到65%～70%,而铅、锌的损失均小于1%。     

含铅锌难选铁矿石的矿物学及铅锌杂质产出特征

以新疆某含铁(FeT)47.04％、含Pb 0.39％、含Zn 0.30％的难选铁矿石为试样,采用化学分析、显微镜观察鉴定、EPMA和EDS等手段,考察其化学成分、铁铅锌的物相组成及铅锌杂质矿物的产出特征,探讨影响选矿工艺的矿物学因素与选矿前景.结果表明:含铁矿物中主要组合为赤褐铁矿、高达91.35％,少量磁性铁和硅酸...     

氯化焙烧-水浸法从锂云母矿提锂(英文)

采用氯化焙烧-水浸法处理锂云母矿,并对氯化处理温度、时间、氯化剂的类型及用量进行研究。条件优化实验表明,在锂云母、氯化钠、氯化钠的质量比为 1:0.6:0.4,氯化处理温度为 880 °C,氯化处理时间为 30 min时,锂的提取率可达 92.86%,钾、铷、铯的提取率分别为 88.49%、93.60%和 93.01%,。采用 XRD 对锂云母原矿及焙烧后物料的物相进行分析。XRD 结果分析表明,当将锂云母和混合氯化剂一起焙烧(氯化钙及氯化钠)时,所得物相为 SiO2、CaF2、KCl、CaSiO3、CaAl2Si2O8、NaCl 和 NaAlSi3O8。     

高磷鲕状赤铁矿添加脱磷剂还原焙烧脱磷机理(英文)

高磷鲕状赤铁矿是一种典型的难处理铁矿石,采用常规的选矿方法难以得到较好的提铁降磷指标。采用添加脱磷剂还原焙烧,然后对焙烧产物进行两段磨矿磁选来处理此类矿石,获得了较好的选别指标。实验结果表明,铁的品位从43.65%(原矿)提高到90.23%(磁选精矿),磷含量从0.82%(原矿)降低到0.06%(磁选精矿),铁的回收率达到87%。采用XRD、SEM、EPMA等分析方法对焙烧产物进行脱磷机理研究。结果表明,在还原焙烧过程中,原矿中有20%的磷灰石生成单质磷随气体挥发,80%的磷灰石没有参与生成单质磷的反应,仍以磷灰石的物相存在于焙烧产物中,而通过磨矿磁选被脱除到尾矿中。磁选精矿中少量的磷以磷灰石的形态存在。在焙烧过程中,加入的脱磷剂与原矿中的脉石矿物(SiO2、Al2O3)反应生成铝硅酸钠,此反应部分破坏原矿的鲕状结构,充分改善焙烧产物中矿物的单体解离程度,有利于后续的磨矿磁选。     

硫化锌银精矿富集银的工艺研究

基于国内外硫化锌矿处理的火、湿法研究进展,对含锌银精矿采用硫酸化焙烧、稀硫酸浸出工艺脱除锌、富集银,考察了焙烧和浸出过程中的主要影响因素。结果表明,硫酸配比为150%,在300℃焙烧90 min,以5%稀硫酸为浸出剂,液固比8:1,搅拌转速200~300 r/min,85℃浸出120 min,最终锌的浸出率可达到98%以上,浸出渣中银含量为7.24%,银被富集7倍。     

铟的还原浸出动力学及分步沉淀选择性分离（英文）

开展硫化锌精矿还原浸出高铁锌浸出渣高效浸铟及浸出液中铟选择性分离的研究。结果表明：在固体物料粒度74～105μm、反应温度90℃、浸出时间300 min、硫酸浓度1.4 mol/L的条件下,铟的浸出率达95%以上。采用收缩核模型对还原浸出动力学进行分析,不同条件下的浸出实验结果表明反应受穿过固体产物层的扩散控制,活化能为17.96 k J/mol,相对于硫酸浓度的反应级数为2.41。铁粉置换沉铜过程铜和砷的沉淀率均达99%以上。98%以上的铟从含高亚铁离子浓度的硫酸锌溶液中选择性分离,获得铟含量约为2.4%的富铟渣,经酸浸-萃取-电积工艺流程进一步处理后可得到纯铟。     

Thermodynamic analysis and process optimization of zinc and lead recovery from copper smelting slag with chlorination roasting

An efficient chlorination roasting process for recovering zinc (Zn) and lead (Pb) from copper smelting slag was proposed. Thermodynamic models were established, illustrating that Zn and Pb in copper smelting slag can be efficiently recycled during the chlorination roasting process. By decreasing the partial pressure of the gaseous products, chlorination was promoted. The Box−Behnken design was applied to assessing the interactive effects of the process variables and optimizing the chlorination roasting process. CaCl₂ dosage and roasting temperature and time were used as variables, and metal recovery efficiencies were used as responses. When the roasting temperature was 1172 °C with a CaCl₂ addition amount of 30 wt.% and a roasting time of 100 min, the predicted optimal recovery efficiencies of Zn and Pb were 87.85% and 99.26%, respectively, and the results were validated by experiments under the same conditions. The residual Zn- and Pb-containing phases in the roasting slags were ZnFe₂O₄, Zn₂SiO₄, and PbS.     

Separation of arsenic from arsenic—antimony-bearing dust through selective oxidation—sulfidation roasting with CuS

The feasibility of a new method for separating arsenic from arsenic—antimony-bearing dusts using CuS was put forward, in which Sb was transformed into Sb₂O₄ and Sb₂S₃ that stayed in the roasted calcine while As was volatilized in the form of As₄O₆. The factors such as roasting temperature and CuS addition amount were studied using XRD, EPMA and SEM—EDS. CuS has an active effect on the separation of arsenic due to the destruction of (Sb,As)₂O₃ structures in the original dust and the simultaneous release of As in the form of As₄O₆. At a roasting temperature of 400 °C and CuS addition amount of 130%, the volatilization rates of arsenic and antimony reach 97.80 wt.% and 8.29 wt.%, respectively. Further, the high As volatile matter can be used to prepare ferric arsenate after it is oxidized, with this treatment rendering the vapor harmlessness.     

PREPARATION OF IRON CARBIDE FROM MAGNETITE PELLETS BY CO-H2 MIXTURES REDUCTION

1. Introduction Iron carbide has some advantages over products of direct reduction iron and has been put into production in industrial scale[1,21, however, the present process only uses natural gas and hematite as material to prepare iron carbide, which has a great limit to the source of kind of reactant gas and iron ore. In recent years, people have done some investigations on preparation of iron carbide using other kind of reactant gas. For example, Hayashi et al[~1 made some investigation …