Separation of alumina and silica from metakaolinite by reduction roasting−alkaline leaching process: Effect of CaSO4 and CaO

Limestone (CaCO₃), which could promote sulfur fixation, was added to coal gangue during roasting in a circulating fluidized bed (CFB) boiler. CaO and CaSO₄ were the main Ca-bearing minerals while metakaolinite was the major Al-bearing mineral in CFB slag. The effect of CaSO₄ and CaO on the separation of alumina and silica from metakaolinite by reduction roasting?alkaline leaching process was studied. Results showed that metakaolinite was completely converted into hercynite and silica solid solutions (i.e., quartz and cristobalite solid solutions) by reduction roasting with hematite. More than 95% of silica in the reduced specimen was removed by alkaline leaching. The addition of CaSO₄ and CaO remarkably decreased the separation efficiency of alumina and silica in metakaolinite, which could be attributed to the formation of Si-bearing minerals: (1) Fayalite and anorthite were formed during the reduction roasting process; (2) Fayalite was stable while anorthite was converted into sodalite and wollastonite during the alkaline leaching process. This study demonstrates that sulfur in coal gangue should be fixed by treating the exhaust gas instead of controlling the combustion process of CFB to achieve the comprehensive recovery of silica and alumina from the CFB slag.     

Preparation of pure SnO_2 powders from tin slag of printed circuit boards waste

The recycling method and principle of SnO2 from the tin slag of printed circuit boards(PCB) waste were investigated. In this study, pure SnO2 powders were obtained through a multi-step process including ball-milling, roasting, dissolving, precipitating, and pickling. The total recovery rate of tin can be up to 91 %. The SnO2 powders obtained is the single phase, and the content of SnO2 is up to 99.9 %. However, the SnO2 particles are easier to agglomerate during the precipitation process. The agglomerate SnO2 particles are about 7.778 lm in mean particle size(D50). This preparation method presents a viable alternative for the tin slag recycling. The tin is not only recycled, but also reused directly to prepare pure SnO2 powders.     

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.     

适用于带钢酸洗的盐酸缓蚀剂

提出了一种以乌洛托品、脂肪胺聚氧乙烯醚、十二烷基硫酸钠、0P-10乳化剂等为主要成份的盐酸缓蚀剂配方的一种高效复合型缓蚀剂,并研究了缓蚀剂添加量和温度对其缓蚀效果的影响,以及该缓蚀剂对废酸回收的影响.结果表明,该缓蚀剂的缓蚀率可以达到90%以上,控制合适的缓蚀剂添加量对废酸回收不会产生严重影响,适合大型钢板连续酸洗生产线使用.     

Reduction roasting–magnetic separation of vanadium tailings in presence of sodium sulfate and its mechanisms

Reduction roasting with sodium sulfate followed by magnetic separation was investigated to utilize vanadium tailings with total iron grade of 54.90 wt% and TiO_2 content of 17.40 wt%. The results show that after reduction roasting–magnetic separation with sodium sulfate dosage of 2 wt% at roasting temperature of 1150 °C for roasting time of 120 min, metallic iron concentrate with total iron grade of 90.20 wt%, iron recovery rate of 97.56 % and TiO_2 content of 4.85 wt% is obtained and high-titanium slag with TiO_2 content of 57.31 wt% and TiO_2 recovery rate of 80.27 % is also obtained. The results show that sodium sulfate has a catalytic effect on the reduction of tailings in the novel process by thermodynamics, scanning electron microscopy(SEM) and X-ray diffraction(XRD) and reacts with silica and alumina in the tailings to form sodium silicate and sodium aluminosilicate. Migration of elements and chemical reactions destroy the crystal structures of minerals and promote the reduction of vanadium tailings, resulting in that iron grains grow to large size so that metallic iron concentrate with high total iron grade and low TiO_2 content is obtained.     

Efficient separation of alumina and silica in reduction-roasted kaolin by alkali leaching

Alkali leaching was employed to investigate the separation of alumina and silica in roasted kaolin obtained by roasting kaolin alone in air at 1273 K for 60 min and in clinker prepared by roasting the mixed raw meal of kaolin, ferric oxide and coal powder with Fe₂O₃/Al₂O₃/C molar ratio of 1.2:2.0:1.2 in reducing atmosphere at 1373 K for 60 min. The thermodynamic analyses and alkali leaching results show that the composition of the Al?Si spinel in roasted kaolin is close to that of 3Al₂O₃·2SiO₂ and the spinel is dissolved with increasing leaching time, resulting in difficulty in deeply separating alumina and silica in kaolin by the traditional roasting?leaching process. On the contrary, the efficient separation of alumina and silica in kaolin can be reached by fully converting kaolinite into insoluble hercynite and soluble free silica, namely quartz solid solution and cristobalite solid solution, during reduction roasting, followed by alkali leaching of the obtained clinker. Furthermore, experimental results from treating high-silica diasporic bauxite indicate that the reduction roasting?alkali leaching process is potential to separate silica and alumina in aluminosilicates.     

废弃SCR脱硝催化剂与废NaCl盐焙烧回收催化剂中的钛、钒、钨（英文）

通过废弃选择性催化还原(SCR)脱硝催化剂与废NaCl盐焙烧,可以将催化剂中的钨和钒与钛分离。在最佳浸出条件下(焙烧温度900℃,焙烧时间3 h,废盐与废催化剂的质量比为0.5,浸出温度80℃,反应时间60 min),钨和钒的浸出率分别达到84.63%和66.42%,同时钛的损失率仅为1.3%。废NaCl盐和焙烧温度可以促进锐钛矿型TiO₂转化为金红石型TiO₂,反应后得到了金红石型TiO₂。金红石型TiO₂中的钛的价态为四价,晶格氧和化学吸附氧分别占57.26%和42.74%。该方法可以同时解决2种废弃物的处置问题。     

酸洗生产中硅泥问题分析与探讨

在硅钢酸洗和超高强钢酸洗工艺过程中,带钢表面的氧化层及少量基体会与盐酸反应,同时也会有少量不与盐酸反应的Si、Mn、Nb、Ti、Cr及其氧化物,这些物质会被带到盐酸循环系统中,容易在循环系统设备表面结垢,堵塞如酸加热器、喷梁、阀门等设备,影响带钢酸洗质量。通过钽管加热器在酸循环系统中的应用实例及高温除硅泥的试验分析,对比了几种硅泥问题解决方案,提出了采用钽管酸加热器和高温除硅泥工艺相结合的硅泥问题解决方案。     

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.     

报废银钨触头回收银和钨的工艺研究

银钨合金触点是重要的电接触材料,报废银钨触点中银钨的分离和资源化问题引起了越来越多的重视.以报废银钨触头为原料,提出了“一次分银-焙烧-二次分银”的回收工艺,考察了银钨分离的工艺条件,通过实践验证了其可靠性,银、钨的回收率达到99％以上.回收得到的银和钨可分别深加工成粒度为0.3～0.6 μm的超细银粉和纯净的WO3产...     

Digestion mechanism and crystal simulation of roasted low-grade high-sulfur bauxite

Low-grade high-sulfur bauxite was pretreated via suspension roasting and muffle furnace roasting to remove sulfur and enhance digestion properties. The results show that sulfur can be efficiently removed, and the alumina digestion properties are significantly improved after suspension roasting. Under optimal conditions (t=70 min, T=280 °C, w(CaO)=8% and N_k=245 g/L), the digestion ratios are 94.45% and 92.08% for the suspension-roasted and muffle-roasted ore, respectively, and the apparent activation energies are 63.26 and 64.24 kJ/mol, respectively. Two crystal models were established by Materials Studio based on the XRD patterns. The DFT simulation shows that the existing Al—O bands after suspension roasting can improve alumina digestion. The (104) and (113) planes of Al₂O₃ after suspension roasting are found to combine with NaOH more easily than those of Al₂O₃ treated in a muffle furnace.     

Extraction of Molybdenum from Molybdenite Concentrates with Hydrometallurgical Processing

Molybdenite concentrates are usually treated by roasting, but low-concentration SO₂ pollution is an associated problem. A hydrometallurgical process with pressure oxidation leaching (POX) and solvent extraction (SX) was developed in recent years. During POX, the oxidation of molybdenum (Mo) is above 98%. More than 95% of the rhenium (Re) and 15% to 20% of the Mo are leached into solution. The sulfur in the concentrate is converted to H₂SO₄, which results in high acidity of the solution. SX was used to recover the Re and Mo from the solution. The extraction of Re and Mo were above 98%. The loaded organic reagent is stripped with ammonia. More than 98% of the Mo can be stripped from the organic phase. Compared with the roasting process, the total recovery of Mo increased from 93% to 97% and that of Re from 60% to 90% when POX and SX are utilized.     

隧道式热风循环电热干燥炉及其应用

本文介绍了一种干燥效率高、节能降耗、温差小、干燥产品质量高的隧道式热风循环电热干燥炉.该炉采用多点分区控温,各均温区为一个热风循环组元,实现炉内双侧纵横双向热风循环,热风循环组元根据干燥传热学按干燥第一阶段、第二阶段在干燥炉长度方向间距不同布置.可根据产品规格和工艺需要,实现严格的温度曲线,同时也便于调整.热风循环装置应用于隧道式干燥炉,能强化对流换热,均匀炉温,提高干燥质量.推广应用结果表明:本炉型非常适合于规模化生产高质量树脂磨具,比箱式炉可节能约40%.     

High-temperature chlorination of gold with transformation of iron phase

Gold in cyanide tailings from Shandong Province is mainly encapsulated by hematite and magnetite at distribution rates of 76.49 % and 10.88 %, respectively.Chlorination–reduction one-step roasting of cyanide tailings was conducted under the following conditions: calcium chloride dosage of 6 %, bituminous coal dosage of30 %, calcium oxide dosage of 10 %(all dosages are vs.the mass of cyanide tailings) at 1000 °C of roasting temperature. X-ray diffraction(XRD), scanning electron microscopy(SEM), and chemical-phase analysis were performed to investigate the effects of iron phase transformation on the high-temperature chlorination of gold.Results indicate that the lattice structure of hematite undergoes expansion, pulverization, and reorganization when hematite is reduced to magnetite, which leads to42.03 % gold exposure, and the high-temperature chlorination rate of gold is 41.17 % at the same time. The structure of wustite formed by the reduction in magnetite is porous and loose, and thus 44.02 % of gold is exposed. The high-temperature chlorination rate of gold is increased by41.98 percentage points. When wustite is reduced to metallic iron, 4.42 % of gold is exposed, and the hightemperature chlorination rate of gold is increased by3.38 percentage points. Accordingly, the high-temperature chlorination of gold mainly occurs in two stages, in which Fe_2O_3 is reduced to Fe_3O_4, and Fe_3O_4 is reduced to Fe_xO finally.     

Controlling SO2 emissions in the roasting of gold concentrate

Acid pressure oxidation is generally believed to be the most effective pretreatment for refractory sulfide gold ores and concentrates, although it is ineffective for some ores. For such ores and concentrates (usually those containing both carbonaceous and sulfide materials), roasting remains the most effective pretreatment before cyanidation. However, the serious environmental damage that could result from the emission of SO₂ and As₂O₃ makes roasting an unattractive pretreatment step. In an effort to develop a technique having lower SO₂ emissions, a lime agglomeration roast (LAR) pretreatment process was investigated. The technique involves the agglomeration of calcium-based SO₂ sorbent with the refractory ore or concentrate before roasting. The results of the LAR pretreatment followed by cyanidation indicate that up to 95 percent of the sulfur dioxide can be captured and more than 90 percent gold extraction can be achieved.     

Oxidation Mechanism of Silver Selenide

The mechanism of selenium vaporisation from silver selenide, the major carrier of selenium in copper depleted anode slimes, was studied using isothermal oxidation of synthetic, massive Ag₂Se in flowing, dry oxygen and oxygen + 20 % sulfur dioxide mixtures at 450–550 °C. The reaction rates expressed as mass change rate indicate that in pure oxygen at 550 °C the reactions essentially cease within 1 h. Silver selenite scales were identified on the interface by SEM–EDS observations. The melting point of silver selenite is 540 °C, and above it selenite is formed in the molten state, where metallic silver as condensed deselenization product is embedded. Metallic silver as a roasting product of Ag₂Se will be partly sulfated in SO₂-bearing atmospheres. A small fraction of selenium is transformed into silver selenite, as the direct oxidation product of Ag₂Se. The reactions proceed much faster into the selenite matrix along the grain boundaries than the surface reactions. The reaction product layer is composed of very fine Ag₂SeO₃ and Ag₂SO₄ particles on the primary Ag₂Se, and significant porosity is generated next to the reaction zone.     

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.     

Thermal Recycling of Waelz Oxide Using Concentrated Solar Energy

The dominating Zn recycling process is the so-called Waelz process. Waelz oxide (WOX), containing 55–65% Zn in oxidic form, is mainly derived from electric arc furnace dust produced during recycling of galvanized steel. After its wash treatment to separate off chlorides, WOX is used as feedstock along with ZnS concentrates for the electrolytic production of high-grade zinc. Novel and environmentally cleaner routes for the purification of WOX and the production of Zn are investigated using concentrated solar energy as the source of high-temperature process heat. The solar-driven clinkering of WOX and its carbothermal reduction were experimentally demonstrated using a 10 kW_th packed-bed solar reactor. Solar clinkering at above 1265°C reduced the amount of impurities below 0.1 wt.%. Solar carbothermal reduction using biocharcoal as reducing agent in the 1170–1320°C range yielded 90 wt.% Zn.     

Recovery of iron from Baotou rare earth tailings by magnetizing roast

The difference of physicochemical properties among minerals in Baotou rare earth tailings is not significant,which leads to a great difficulty in separation of minerals.In this article,the process of magnetizing roast and low-intensity magnetic separation was used to recover iron.Effect of calcination temperature,holding time and carbon/oxygen ratio on roasting efficiency was investigated.The parameters evaluating magnetizing roast efficiency and theoretical value were determined.X-ray diffraction(XRD)analysis was used to investigate the conversion of Fe phase after roasting.The results show that the best magnetizing roast conditions are calcination temperature of 650℃,holding time of 2.5 h,and carbon/oxygen molar ratio of 3.85.The best magnetization rate is 2.36,which is close to the theoretical value of 2.33.Based on experiments of low-intensity magnetic separation under different intensities,the best current intensity is 2.0 A to obtain the best separation results.Under the best condition,the concentrate grade of iron is 45.45% and the recovery of iron is 68.36%.Most of rare earth,fluorine,and phosphorus are enriched in the magnetic separation tailings.The XRD analysis shows that Fe exists in Fe2O3 before roasting and exists in Fe3O4after roasting.     

Reaction mechanisms of low-grade molybdenum concentrate during calcification roasting process

The effects of Ca-based additives on roasting properties of low-grade molybdenum concentrate were studied. The results show that calcium-based additives can react with molybdenum concentrate to form CaSO₄ and CaMoO₄. The initial oxidation temperature of MoS₂ is 450 °C, while the formation of CaMoO₄ and CaSO₄ occurs above 500 °C. The whole calcification reactions are nearly completed between 600 and 650 °C. However, raising the temperature further helps for the formation of CaMoO₄ but is disadvantageous to sulfur fixing rate and molybdenum retention rate. Calcification efficiency of Ca-based additives follows the order: Ca(OH)₂>CaO>CaCO₃. With increasing the dosage of Ca(OH)₂, the molybdenum retention rate and sulfur-fixing rate rise, but excessive dosages would consume more acid during leaching process. The appropriate mass ratio of Ca(OH)₂ to molybdenum concentrate is 1:1. When roasted at 650 °C for 90 min, the molybdenum retention rate and the sulfur-fixing rate of low-grade molybdenum concentrate reach 100% and 92.92%, respectively, and the dissolution rate of molybdenum achieves 99.12% with calcines being leached by sulphuric acid.