氧化镍和硫化铜矿联合熔炼炉渣的黏度（英文）

测定氧化镍和硫化铜矿联合冶炼熔渣的黏度。制备的模拟熔渣与实际熔渣的成分和结构相似,并限制研究的熔渣的成分范围,含铁熔渣的成分(质量分数)为:8.9%CaO、11.8%MgO、12.5%Al₂O₃、47.4%SiO₂、13.3%FeO和5.0%Fe₂O₃;无铁熔渣是将含铁熔渣中的氧化铁去除得到,其成分(质量分数)为:12.5%CaO、16.0%MgO、9.4%Al₂O₃和58.3%SiO₂。由振动法测试的含铁熔渣(1550～1300℃)和无铁熔渣(1550～1400℃)的黏度实验值分别为0.31～2.33 Pa·s和1.28～4.55 Pa·s。由Kalmanovitch-Frank模型预测的含铁熔渣的黏度值与实验值存在偏差。通过Weimann-Frenkel-Urbain方程对实验数据进行回归分析,提出可预测所考虑温度范围内的初级熔渣黏度的原始经验模型。当碱度从0.7(含铁熔渣)降低到0.6(无铁熔渣),黏性流动...     

鼓风炉水淬铅冶渣的矿物特征

通过X射线衍射分析、扫描电子显微镜、热重差热分析仪以及化学分析对铅冶炼渣进行系统矿物特征研究。结果表明:铅冶炼渣中主要金属矿物为硅锌矿、锌黄长石、锌铁尖晶石、方铁矿以及金属铅。锌元素高度分散在玻璃渣相中,其中锌在硅酸盐中占66.28%(质量分数),在铁酸盐中占31.63%(质量分数),选矿回收有价金属锌需要以含锌矿物矿相重构为基础。金属铅与锌铁尖晶石伴生,可以采用强磁选回收锌铁尖晶石而富集金属铅。铅冶炼渣的水溶液为弱碱性,水溶液pH值与渣料的粒度大小有关。在强酸或强碱条件下,金属的溶解行为不同,控制碱性浸出条件可作为冶炼渣中铅选择性提取的有效方法。     

不锈钢冶炼渣熔融还原及毒性浸出实验研究

通过半工业化试验证实了可以通过铁浴熔融还原工艺快速回收不锈钢冶炼渣中的铬。采用X射线衍射分析、荧光分析和化学分析方法,得到不锈钢渣的初始成分及物相,熔融还原处理后熔渣中Cr最优收得率达到98.6%,渣中残铬量最低可降至0.063%。并根据相关标准进行了毒性浸出试验,确认了不锈钢冶炼渣经过处理,浸出毒性从0.62 mg/L降低至0.017 86 mg/L,远远低于一般废弃物的标准。     

从氰化尾渣衍生物制备标准铅、锌精矿的新工艺

以氰化尾渣衍生物为原料,制备标准铅、锌精矿。采用X体系在常压下进行氧化浸出,研究浸出电位、氧化剂、总酸量、试剂A、试剂B、浸出时间和浸出温度对锌回收率、锌精矿品位以及铅浸出率的影响。结果表明:采用二段浸铅与气液固强化浸出相结合的方法在高效气液固反应器中进行实验,其最优条件如下:试剂A量15.6 g/L,试剂B量90 g/L,液固比L/S=10:1,鼓氧量1.5 L/min,浸出温度70℃,每段浸出时间均为3 h。所得铅精矿和锌精矿氧量品位分别高达75.49%和45%,副产品硫磺的品位达到99%,其中,铅、锌总回收率分别为90.68%和99%,单质硫的回收率高达99.1%。     

常压-氧压联合浸出锌置换渣中的镓和锗

针对锌置换渣中有价金属元素种类多、物相组成复杂的特点,本研究提出采用一段常压-二段氧压浸出的方法高效浸出锌置换渣中的有价金属。采用XRD和SEM-EDS对浸出渣物相以及形貌进行了分析。结果表明：在硫酸浓度1.5 mol/L、温度80℃、液固比7.5 mL/g、浸出时间3 h的条件下,常压浸出过程中Cu、Zn、Cd、Fe、Ni、Ga和Ge的浸出率分别为97.48%、99.43%、99.82%、97.21%、98.97%、97.74%、82.46%。对常压浸出渣进行二段氧压浸出,在氧气分压0.6 MPa以及硫酸浓度为0.25 mol/L条件下,Cu和Ge的浸出率可进一步分别提高至99.87%和91.66%。通过两段浸出,原来在置换渣中存在的Cu、Zn、Fe等物相消失,浸出渣主要由Pb和Si组成;铅的主要物相为PbSO₄,Si以粒径较小的聚合硅胶颗粒和块状SiO₂颗粒形式存在;聚合硅胶颗粒和块状SiO₂颗粒中Ge含量较高,对Ge的浸出造成不利影响。     

从锌置换渣中分段浸出锌、铅、镓和锗

采用分段浸出方法从锌置换渣中选择性提取锌、铅、镓和锗.首段浸出过程通过控制合适的硫酸浓度与液固比条件,锌和镓的浸出率达到90％和99％以上,超过92％的锗留在硫酸浸出渣中;二段采用盐酸作为浸出剂选择性提取铅,铅的浸出率为99％,锗损失率小于2％;三段采用1 mol/L氢氧化钠作为浸出剂,通过破坏硅锗固溶体的结构,实现锗...     

含钛夹杂物对连铸保护渣性能的影响

含钛钢连铸过程中析出氮化钛夹杂,并参与渣-金界面反应,严重恶化传统含钛钢连铸保护渣的理化性能。研究了不同含量的TiN和TiO₂对不同碱度保护渣的熔点、黏度、结晶比例以及凝固温度等性能的影响规律。测试结果表明,TiN对CaO-SiO₂基保护渣的性能影响最大,随着渣中TiN含量的升高,保护渣的黏度、结晶比例和凝固温度大幅度升高;当TiN质量分数大于5%时,熔点大幅升高。TiO₂对CaO-Al₂O₃基保护渣的性能影响最大,当TiO₂质量分数小于4%时,可小幅度降低保护渣的黏度和凝固温度;当TiO₂质量分数大于10%时,渣中钙钛矿成为主要析出物相,熔渣的性能严重恶化,影响对坯壳的润滑作用。通过对比3种不同碱度的基础渣发现,CaO-SiO₂-Al₂O₃基保护渣具有较好吸收含钛夹杂物的能力。     

生物浸出体系中中等嗜热菌对锌冶炼窑渣中金属的提取及其电化学特性的影响(英文)

研究中等嗜热菌对锌冶炼窑渣中金属提取的影响以及生物浸出过程中锌冶炼窑渣碳糊电极的电化学特性。结果表明,在矿浆浓度2%、pH1.0、温度65°C、转速为120r/min的浸出条件下,去除生物浸出体系中吸附菌后,废渣中Fe、Cu和Zn的浸出率分别为86.7%、90.3%和66.7%,而在没有去除吸附菌体系中3种金属的浸出率分别为91.9%、96.0%和84.5%。对生物浸出渣和酸浸渣表面细菌分泌物进行FT-IR测试分析可知,生物浸出渣颗粒表面出现了新的官能团振动峰,如1007cm1和1193cm1处的峰,间接说明残渣颗粒表面吸附细菌的存在。生物浸出体系和空白体系的循环伏安曲线和塔菲尔曲线特性进一步表明生物浸出体系中细菌促进了锌冶炼窑渣中有价金属的溶出。     

二氧化硫还原分解铁酸锌及锌浸渣工艺

锌冶炼浸出渣中锌主要以铁酸锌的形式存在,针对锌浸渣中铁酸锌难于分解的问题,以铁酸锌作为研究对象,研究二氧化硫作用下铁酸锌中锌的溶出和Fe(Ⅲ)的还原行为。考察初始硫酸浓度、液固比、二氧化硫通入量、反应时间、反应温度对二氧化硫还原分解铁酸锌行为的影响。结果表明:最佳反应条件如下,初始硫酸浓度120 g/L、液固比11:1、二氧化硫通入量0.41×10~(-2)mol/g、反应时间120 min、反应温度105℃。在最佳反应条件下,对锌浸渣开展还原浸出实验,锌的浸出率能达到99%以上,Fe(Ⅲ)的还原率能达到98%。通过ICP-MS和XRD分析表明,锌浸渣中的铁酸锌完全分解,还原浸出渣的主要成分为锌和铅,分别以ZnS和PbSO_4的形式存在。     

碱度与BaO含量对保护渣理化性能的影响

为获得适用于高拉速薄板坯生产的高结晶高润滑性保护渣组分范围，综合旋转黏度计、四探针法电导率测定仪、半球点熔点仪、X射线衍射分析仪，研究碱度和BaO质量分数对保护渣理化性能的影响规律。试验结果表明，保护渣碱度提高，熔点、转折温度、开始析晶温度、析晶比例均呈现先降低后升高的规律，而高温黏度变化较小，其中，转折温度和析晶比例分别在碱度为1.65和1.55时达到最低值；析晶矿相中枪晶石Ca₄Si₂O₇F₂矿相析出量减少，Ca₅MgSi₃O₁₂和CaF₂矿相析出量增加。当碱度相同时，保护渣中BaO质量分数增加，熔点略微降低；高温黏度呈现小幅波动；转折温度先降低后升高，在BaO质量分数为6%时最低；开始析晶温度降低，且开始析晶温度越来越接近转折温度。综合而言，碱度为1.65、BaO质量分数为6%的保护渣最有利于润滑和传热的协调控制。     

Leaching of iron concentrate separated from kiln slag in zinc hydrometallurgy with hydrochloric acid and its mechanism

It is taken as a novel prospective process to treat iron concentrate from hydrometallurgical zinc kiln slag for comprehensive utilization of valuable metals by a hydrochloric acid leaching-spray pyrolysis method. The leaching mechanism of different valuable metals was studied. The results revealed that the leaching rates of Ag, Pb, Cu, Fe, As and Zn were 99.91%, 99.25%, 95.12%, 90.15%, 87.58% and 58.15%, respectively with 6 mol/L HCl and L/S ratio of 10:1 at 60 °C for 120 min. The action of SiO₂ in leaching solution was also studied. The results showed that the precipitation and settlement of SiO₂(amorphous) adsorbed part of metal ions in solution, which greatly inhibited the leaching of Cu, Fe, As and Zn, so it is crucial to control the precipitation of amorphous SiO₂.     

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.     

湿法炼锌酸浸出渣浮选回收银试验

传统锌精矿伴生银回收,均是从挥发窑渣中进行,由于矿物性质变化,导致银无法有效回收。进行了从锌酸浸出渣中浮选回收银的研究,解决了锌精矿中伴生银回收的问题。某湿法炼锌浸出渣含Ag 350 g/t,Au0.01 g/t,Pb 3.87%,Zn 17.45%,Cu 1.38%,银主要以自然银存在,占60.13%。采用高效捕收剂HT-1#、起泡剂HT-2#进行浮选,经过一粗四精三扫工艺,获得银精矿产率4.39%,含银6616 g/t,银回收率82.98%,取得良好的经济技术指标。     

Ida~（2-）-H_2O体系浸出低品位氧化锌矿

采用Ida2--H2O体系（亚氨二乙酸盐水溶液）处理高碱性脉石型低品位氧化锌矿,考察浸出时间、液固比、配体总浓度、温度及pH值对矿物中主金属Zn及杂质元素Ca、Mg、Cu、Ni、Fe、Pb、Cd的溶出影响。结果表明：在弱碱性Ida2--H2O体系中,Ca、Mg、Fe不会被大量溶出,有价金属Cu、Ni、Pb、Cd可部分随主金属Zn溶出而进入浸出液;在浸出时间4h、液固比5：1、配体总浓度0.9mol/L、温度70℃、pH8的优化条件下,锌浸出率为76.6%。     

Extraction of manganese and zinc from their compound ore by reductive acid leaching

Comprehensive utilization of low grade manganese–zinc compound ore containing lead and silver with a method of reductive acid leaching was studied. According to the ?–pH diagram of Mn–Zn–H₂O system, Mn and Zn can be leached simultaneously in the pH range of –2 to 5.61. The results showed that both hydrogen peroxide and sucrose were effective reductants which could intensify the simultaneous leaching of Mn and Zn into leachate as well as enrich Pb and Ag in the residue. 95.88% of Mn and 99.23% of Zn were extracted when the compound ore was leached with hydrogen peroxide in sulfuric acid media, meanwhile the contents of Pb and Ag in the residue were enriched to 13.21% and 489.36 g/t, respectively. When sucrose was used as the reductant, the leaching efficiencies of Mn and Zn separately achieved 98.26% and 99.62%, and contents of Pb and Ag in the residue were as high as 13.92% and 517.87 g/t, respectively.     

Effect of B2O3 content on viscosity and structure of SiO2−MgO−FeO-based slag

The effect of B₂O₃ content on the viscosity of SiO₂?MgO?FeO-based molten slag system was investigated using the rotating cylinder method. The evolution process of the melt structure under different contents of B₂O₃ was comprehensively studied via FTIR spectroscopy and a model for calculating the degree of polymerization was developed. The results showed that the viscosity of the molten slag decreased with the addition of B₂O₃, which had a slight effect when its content exceeded 3 wt.%. As the addition of B₂O₃ increased from 0 to 4 wt.%, the break temperature of the slags decreased from 1152 to 1050 °C and the apparent activation energy decreased from 157.90 to 141.84 kJ/mol. The addition of B₂O₃ to the molten slag destroyed the chain silicate structure to form a more cyclic borosilicate structure. The Urbain model was improved to calculate the viscosity of the SiO₂?MgO?FeO-based slags, and the values were in good agreement with the experimentally measured values.     

Effect of Fe/SiO2 and CaO/SiO2 mass ratios on metal recovery rate and metal content in slag in oxygen-enriched direct smelting of jamesonite concentrate

The oxygen-enriched direct smelting of jamesonite concentrate was carried out at 1250 °C by changing the slag composition. The effects of Fe/SiO₂ and CaO/SiO₂ mass ratios on the metal recovery rate as well as metal content in slag were investigated. Experimental results indicated that the metal (Pb+Sb) recovery rate was up to 88.30%, and metal (Pb+Sb) content in slag was below 1 wt.% under the condition of slag composition of 21−22 wt.% Fe, 19−20 wt.% SiO₂ and 17−18 wt.% CaO with Fe/SiO₂ mass ratio of 1.1:1 and CaO/SiO₂ mass ratio of 0.9:1. The microanalysis of the alloy and slag demonstrated that the main phases in the alloy contained metallic Pb, metallic Sb and a small amount of Cu₂Sb and FeSb₂ intermetallic compounds. The slag was mainly composed of kirschsteinite and fayalite. Zinc in the raw material was mainly oxidized into the slag phase in the form of zinc oxide.     

Minimization of Copper Losses in Copper Smelting Slag During Electric Furnace Treatment

In the quest to achieve the highest metal recovery during the smelting of copper concentrates, this study has evaluated the minimum level of soluble copper in iron-silicate slags. The experimental work was performed under slag-cleaning conditions for different levels of Fe in the matte and for a range of Fe/SiO₂ ratios in the slag. All experiments were carried out under conditions where three phases were present (copper?Cmatte?Cslag), which is the condition typically prevailing in many slag-cleaning electric furnaces. The %Fe in the electric furnace matte was varied between 0.5?wt.% and 11?wt.%, and two different Fe/SiO₂ ratios in the slag were used (targeted values were 1.4 and 1.6). All experiments were performed at 1200°C. From thermodynamic considerations, from industrial experience, and from the results obtained in this study, the minimum soluble copper content in the electric furnace slag is expected to be near 0.55?wt.% Cu. This level does not account for a portion of the copper present as mechanically entrained matte/metal droplets. Taking this into account, the current authors believe an overall copper level in discard slag between 0.7?wt.% and 0.8?wt.% can be obtained with optimal operating conditions. For these conditions, the copper losses in the slag are roughly 75% as dissolved copper and 25% as entrained matte and copper. Such conditions include operating the electric furnace at metallic copper saturation, maintaining the %Fe in the electric furnace matte between 6?wt.% and 9?wt.%, not exceeding a slag temperature of 1250°C, and controlling the Fe/SiO₂ ratio in the smelting furnace slag at ??1.5. In addition, magnetite reduction needs to be performed efficiently during the slag-cleaning cycle so as to maintain a total magnetite content of ??7?wt.% in the discard slag. The authors further consider that under exceptionally well-controlled conditions, a copper content in electric furnace discard slag between 0.55?wt.% and 0.7?wt.% can be obtained, by minimizing entrained matte and copper solubility in the discard slag.     

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%。