加压浸出低品位锰矿的工艺

对低品位锰矿硫酸加压浸出工艺进行正交实验和单因素实验,通过正交实验得出:加压浸出低品位锰矿工艺中,影响锰浸出率的主要因素的较佳组合如下:初始酸浓度160 g/L、硫铁矿量50 g、液固比5:1(5 mL/g)、压力1 MPa、温度180℃、时间80 min。为分析低品位锰矿中锰、铁及铝的浸出行为,实现这3种金属元素的高效分离,参考正交实验结果适当地调整工艺参数,进行单因素实验研究,详细考察始酸浓度、反应温度、硫铁矿量、液固比、浸出时间和浸出压力对锰、铁及铝浸出率的影响,得到优化浸出工艺条件如下:低品位锰矿粉100 g,初始硫酸浓度120 g/L,浸出反应温度120℃,硫铁矿量50 g,液固比5:1(5 mL/g),浸出时间100 min,浸出压力0.7 MPa,搅拌转速500 r/min。本工艺具有良好的稳定性,在优化浸出条件下,锰的浸出率为96%,而铝和铁的浸出率分别为38.7%和7.12%,实现锰选择性高效溶出,锰和铝、铁等杂质的分离效果良好,为最终实现低品位锰矿中各种有价元素的清洁高效回收奠定了基础。     

废盐酸浸出菱锰矿制备四水氯化锰

以某低品位菱锰矿为原料,采用废盐酸浸出,研究浸出温度、浸出时间、液固比、反应时酸过量系数对矿石中锰的浸出效果的影响,讨论锰浸出的最佳工艺条件.结果表明:浸出温度80 ℃、浸出时间60 min、液固比2.5-1、酸过量系数1.3为最佳工艺条件.该工艺为低品位锰矿的开发利用及钛厂废盐酸的综合利用开辟了一条新途径.     

响应面优化蔗渣焙烧还原低品位软锰矿的工艺(英文)

采用基于统计的优化策略优化了无氧条件下蔗渣焙烧还原低品位软锰矿的工艺。用中心组合设计收集实验数据,用二次模型表示锰浸出率与渣矿比(蔗渣与锰矿质量比)、焙烧温度、焙烧时间的函数关系,用统计分析(ANOVA)研究变量及变量的相互作用对浸出过程的影响。结果表明,渣矿比和焙烧温度对浸出过程的影响比焙烧时间的大,渣矿比和焙烧温度的线性项、二次项及其交互作用影响显著,而焙烧时间的影响却较小。利用所得的二次模型可得最佳工艺参数:渣矿比0.9:10、焙烧温度450°C、焙烧时间30min。在优化条件下,锰浸出率的预测值为98.1%,实验值为98.2%.     

采用响应面方法和中心复合旋转设计对低品位锰矿浸出过程建模与优化(英文)

采用实验设计软件对从低品位锰矿中浸出提取锰的过程进行优化。在中心复合响应面实验设计中,考察了4个主要影响浸出过程的参数,即硫酸浓度、草酸浓度、浸出时间和温度。将锰和铁的浸出率作为考察指标。采用统计分析和方差分析确定了最优条件,即最高的锰和铁浸出率、最短的浸出时间和最低的温度。结果表明,硫酸浓度是影响浸出过程的最显著的参数,在最优条件下:硫酸浓度7%,草酸浓度42.5g/L,浸出时间60min,反应温度65℃,锰和铁的浸出率可分别达到93.44%和15.72%。     

低品位铂钯精矿的富氧压浸出

对高镁低品位复杂铂钯精矿进行工艺矿物学分析,提出采用硫酸氧压浸出工艺对该精矿中的贱金属铜、镍、铁选择性浸出分离并富集铂钯的处理工艺。考察磨矿粒度、反应温度、时间、初始硫酸浓度、氧压、搅拌速度、木质素磺酸钙用量、液固比对铜、镍、铁浸出率及渣率的影响,确定最佳工艺参数。实验结果表明:当精矿粒度小于43μm占有率为93%、时间3 h、浸出温度150℃、初始硫酸浓度2 mol/L、氧分压0.7 MPa、搅拌速度400 r/min、添加剂木质素磺酸钙用量0.6 g、液固比5:1的最佳工艺条件下,铜浸出率达99.27%、镍浸出率达98.04%、渣率为37%左右,铂钯几乎不被浸出,铂和钯在浸出渣中富集近3倍。     

四川某低品位微细粒氧化金矿非氰浸出试验研究

对金品位为2.02 g/t的某低品位氧化微细粒金矿开展了全泥浸出提取金的试验研究。优选出非氰浸出剂CC-1,确定了相应工艺参数,在此基础上开展了3个粒级柱浸试验,对柱浸含金溶液进行了活性炭吸附试验,研究表明该矿石适宜于利用非氰浸出剂CC-1堆浸回收金。矿石磨至-200目占80%、矿浆液固比2:1、石灰用量3000 g/t原矿、CC-1浓度0.10%、浸出时间30 h条件下金浸出率92.75%;在石灰用量3000 g/t、CC-1浓度0.10%、浸出时间10 d时-10 mm矿样Au浸出率92.46%,浸出时间15 d时-20 mm及-30 mm矿样Au浸出率分别为91.49%、89.24%。采用CC-1作为浸出剂的含Au溶液活性炭吸附率为95.72%~97.11%。     

以废茶叶为还原剂在硫酸溶液中还原浸出氧化锰矿（英文）

采用废茶叶在硫酸溶液中还原浸出加蓬和湘西氧化锰矿石,探索废茶叶用量、硫酸浓度、固液比、浸出温度和反应时间对浸出过程的影响。对加蓬氧化锰矿,优化的浸出条件为:氧化锰矿与废茶叶的质量比10:4、硫酸浓度2.5 mol/L、固液比7.5:1、浸出温度368 K、浸出时间8 h;在此条件下,加蓬氧化锰矿的浸出率几乎达100%。对于湘西氧化锰矿,优化浸出条件为:氧化锰矿与废茶叶的质量比10:1、硫酸浓度1.7 mol/L、液固比7.5:1、温度368 K、浸出时间8 h;在此条件下,锰的浸出率达到99.8%。氧化锰矿的还原浸出过程符合内扩散控制模型,加蓬和湘西氧化锰矿石的还原浸出反应表观活化能分别为38.2 kJ/mol和20.4 kJ/mol。采用X-射线衍射(XRD)和扫描电子显微镜(SEM)对浸出前、后的锰渣进行表征。     

低品位含铝矿物焙烧转型浸出铝试验研究

提出了从低品位含铝矿物中提取铝的新思路。采用微生物浸出技术,考察了低品位含铝矿物高温转型浸出铝的动力学,结果表明矿石在525℃下焙烧4h后铝的浸出效果最好,而铁的浸出率大幅度下降。系统研究了焙烧温度、浸出温度、浸出时间、更换浸出剂等因素对浸铝效果的影响。高温转型后的低品位含铝矿物浸出铝的条件为:配矿液固比100:10,浸出温度90℃,浸出时间10h,每2h更换一次浸出剂。结果铝的浸出率可达到90.99%。     

超声波辐射对低品位氧化锌矿氨浸行为的影响

研究超声波辐射对兰坪低品位氧化锌矿氨浸过程的影响规律.研究表明:无超声波辐射时,兰坪低品位氧化锌矿在NH3-NH4Cl-H2O体系中浸出3 h后的最高浸出率为69.4%;引入超声波辐射后,显著缩短了浸出时间,无超声波辐射时Zn浸出率达到61.8%需要80 min,而采用超声波辐射浸出时仅需20 min;反应温度、浸出剂浓度和NH4Cl与NH4OH的摩尔浓度比等参数对超声波辐射的强化作用效果显著.当反应温度和浸出剂浓度较低,NH4Cl与NH4OH的摩尔浓度比较大时,超声波辐射的强化作用显著;超声波辐射可望降低氨浸低品位氧化锌矿的反应温度和浸出所需NH4OH浓度,大幅缩短浸出时间;同时,超声强化氨浸对锌的浸出具有较高选择性.     

含银废电子浆料中银的浸出和除杂工艺条件研究

用硝酸作为浸出剂,通过正交试验考察了各工艺条件对含银电子废料中的银的浸出效率的影响。结果表明,硝酸浓度和浸出温度是影响浸出效率的主要因素;在硝酸浓度为8 mol/L、搅拌速度为100 r/min、浸出温度和时间分别为80℃和80 min的最佳条件下,浸出效率可达88%以上。控制温度为70℃,以 NaOH 溶液调节浸出液至近中性(pH=6.0),共存杂质 Cu、Fe、Pb、Ni 和 Bi的一次去除率达到80%以上,可得到较纯净的硝酸银溶液。     

Kinetic study on pressure leaching of high iron sphalerite concentrate

The kinetics of pressure leaching high iron sphalerite concentrate was studied.The effects of agitation rate,temperature, oxygen partial pressure,initial acid concentration,particle size,iron content in the concentrate and concentration of Fe2 added into the solution on the leaching rate of zinc were examined.The experiment results indicate that if the agitation rate is greater than 600 r/min,its influence on Zn leaching rate is not substantial.A suitable rise in temperature can facilitate the leaching reaction,and the temperature should be controlled at 140-150℃.The increase trend of Zn leaching rate becomes slow when pressure is greater than 1.2 MPa,so the pressure is controlled at 1.2-1.4 MPa.Under the conditions of this study,Zn leaching rate decreases with a rise in the initial sulfuric acid concentration;and Zn leaching rate increases with a rise of iron content in the concentrate and Fe 2 concentration in the solution.Moreover,the experiment demonstrates that the leaching process follows the surface chemical reaction control kinetic law of“shrinking of unreacted core”.The activation energy for pressure leaching high iron sphalerite concentrate is calculated,and a mathematical model for this pressure leaching is obtained.The model is promising to guide the practical operation of pressure leaching high iron sphalerite concentrate.     

闪锌矿氧压酸浸过程中银的催化作用

进行了银对闪锌矿氧压酸浸催化作用的研究.通过间歇性小型试验考察了各种因素对浸出率的影响,诸如浸出温度、浸出时间.试验在2升的高压釜内进行,结果表明银离子对闪锌矿氧压酸浸有显著的催化作用.     

Reduction treatment of low-grade manganese ore by biomass roasting

The reduction of manganese dioxide in low-grade manganese ore by biomass roasting process was investigated for extracting manganese from poor manganese ore more effectively. In this study,the cinder of ore fines and sawdust was further leached by sulphuric acid to obtain MnSO4. Over 97% manganese in ores can be converted into MnSO4. Effects of the mass ratio of manganese ore to sawdust, roasting temperature and time, leaching temperature and time, leaching agent concentration and liquid-solid ratio were studied. The manganese recovery achieved 97.71% under the conditions: the mass ratio of manganese ore to sawdust of 5:1, roasting temperature 500℃ for 40min, leaching temperature 60℃ for 40min, sulphuric acid concentration of 1mol/L and liquid-solid ratio of 10:1. This technology can be suitable for extraction of Mn in low-grade manganese ore.     

氧压酸浸低品位富银硫化矿富集提取银和锌

由于含有大量的黄铁矿和白铁矿(它们约占原矿的70%,质量分数),以及闪铅矿一定程度上的氧化,云南澜沧铅矿股份有限公司所产的富银硫化矿难以富集.本文通过对该矿在90-170℃下氧压酸浸,以期连同后面的氰化能提取精矿中的银.通过进行2L高压釜的小型试验,考察了温度、酸度、碘化钠用量、氧分压、氧气流速对银和锌回收率的影响.结果表明,银的回收率取决于银是否进入黄钾铁矾渣,或者与碘化钠反应生成碘化银沉淀.在优化的条件下,银和锌的回收率分别达到71.5%和41.29%.     

Recovery of gallium from zinc concentrate by pressure oxygen leaching

Zinc concentrate with high gallium content is one of the main resources of gallium.The gallium presents in the form of isomorphism in tetrahedron coordination with sulfur in sphalerite.The research was to investigate the amenability of zinc concentrate with high gallium to pressure oxygen leaching.The particle size,sulfuric acid concentration,oxygen partial pressure,additive amount,and time of reaction were studied.The extraction yields of gallium and zinc are 86%and 98%,respectively.The optimal condition is 100 g of zinc concentrate with particle size smaller than 38 lm,sulfuric acid concentration150 g L-1,leaching temperature 150℃,leaching time120 min,oxygen partial pressure 0.7 MPa,additive amount of 0.2 wt%.     

高压酸浸镍钼矿提取钼和镍

研究一项针对镍钼矿用高压酸浸的方法回收镍和钼的全湿法工艺。采用该工艺避免了传统上艺焙烧镍钼矿（15％～25％s）带来的人量S02和As2O3排放,减小了对环境的污染;与现有的湿法碱浸回收钼工艺相比,本工艺存酸浸过程中回收了儿乎全部的镍和人部分的钼。在氧压环境下,几乎全部的镍和大部分的钼都进入溶液,少部分的钼留在酸浸渣中,睃浸渣进一步用碱（NaOH）浸出。在最佳的实验条件下,97％的镍和96％的钼分别被浸出。     

Copper and elemental sulphur from chalcopyrite by pressure leaching

Processes employing direct oxidation under an over-pressure of air or oxygen in an aqueous sulphuric acid medium have been developed in the Sherritt Gordon Laboratories for iron, nickel, cobalt, zinc and lead sulphide concentrates. This study has recently also been extended to chalcocite, Cu₂S, concentrates. The rising interest in processes employing direct aqueous oxidation is stimulated by the fact that elemental sulphur can be produced as a by-product rather than sulphur dioxide or sulphuric acid.The present paper outlines a process which features the direct pressure oxidation of the most abundant copper sulphide mineral, chalcopyrite, CuFeS₂. The optimum conditions for a practical pressure leaching step have now been developed in the laboratory which results in the production of copper sulphate solution suitable for copper winninq by electrolysis, hydrogen reduction, solvent extraction combined with electrolysis, or other means. The leach residue yields pure elemental sulphur by-product. Copper and elemental sulphur recoveries of 98 and 85% respectively have been recorded. The fastest oxidation rate, corresponding to a leach retention time of 2.5 hr, was obtained when the copper concentrate was ground to 99.5% — 325 mesh, when a 50% stoichiometric excess of concentrate over the amount of available sulphuric acid for copper was used and when the oxygen partial pressure and temperature were maintained at 500 psi and 240°F, respectively. In an idealized form, the pressure leaching reaction can be expressed as follows:—CuFeS₂ + H₂SO₄ + 1 1/40₂ + 1/2H₂O → CuSO₄ + Fe(OH)₃ + 2S°After separation of the copper sulphate solution by filtration, elemental sulphur and excess concentrate ore recovered from the iron oxide tailing by flotation. The tailing, containing iron oxide and insolubles, is rejected. The elemental sulphur is separated from the concentrate by hot filtration, solvent extraction, distillation, or other means, and the unleached chalcopyrite is recycled to the leaching step.     

Coupling leaching of sphalerite concentrate

Coupling process of sphalerite concentrate leaching in H2SO4-HNO3 and tetrachloroethylene extracting of sulfur was investigated. Effects of leaching temperature, leaching time, mass ratio of liquid to solid and tetrachloroethylene addition on zinc leaching processes were examined separately. SEM images of sphalerite concentrate and residues were performed by using JEM-6700F field emission scanning electron microscope. The relationship between the number of recycling and extraction ratio of zinc was studied. The results indicate that 99.6% zinc is obtained after leaching for 3 h at 85℃ and pressure of 0.1MPaO2, with 20g sphalerite concentrate in 200 mL leaching solution containing 2.0mol/L H2SO4 and 0.2mol/L HNO3, in the presence of 10 mL C2Cl4. The leaching time of zinc is 50% shorter than that in the common leaching. The coupling effect is distinct. The recycled C2Cl4 exerts little influence on extraction ratio of zinc.