硝酸溶液中氧化铜矿浸出工艺参数优化及动力学模拟

研究Tunceli孔雀石矿物在硝酸溶液中的溶出行为,以评估各种实验参数的影响.研究为分两个阶段.在第一步中,确定浸出过程的最佳条件,而在第二步中,对该过程进行动力学评估.在优化实验中,以硝酸浓度、温度、搅拌速度和固液比为自变量,采用中心组合设计法(CCD)获得实验数据.确定硝酸浓度、温度、固液比和搅拌速度的最佳值分别为...     

焙烧氟碳铈矿硫酸浸出稀土的动力学(英文)

研究了硫酸浸出德昌稀土与天青石共伴生矿的焙烧矿过程。考查粒度、搅拌速度、硫酸浓度和温度对稀土浸出率的影响,并对稀土的浸出动力学进行分析。在选定的浸出条件下:粒径0.074～0.100mm、硫酸浓度1.5mol/L、液固比8:1、搅拌速度500r/min,稀土浸出反应受内扩散控制,表观活化能为9.977kJ/mol。     

Preparation of sodium manganate from low-grade pyrolusite by alkaline predesilication–fluidized roasting technique

Low concentration alkaline leaching was used for predesilication treatment of low-grade pyrolusite. The effects of initial NaOH concentration, liquid-to-solid ratio, leaching temperature, leaching time and stirring speed on silica leaching rate were investigated and the kinetics of alkaline leaching process was studied. The results show that silica leaching rate reached 91.2% under the conditions of initial NaOH concentration of 20%, liquid-to-solid ratio of 4:1, leaching temperature of 180 °C, leaching time of 4 h and stirring speed of 300 r/min. Shrinking-core model showed that the leaching process was controlled by the chemical surface reaction with activation energy E_a of 53.31 kJ/mol. The fluidized roasting conditions for preparation of sodium manganate were optimized by the orthogonal experiments using the desiliconized residue. The conversion rate of sodium manganate was obtained to be 89.7% under the conditions of silica leaching rate of 91.2%, NaOH/MnO₂ mass ratio of 3:1, roasting temperature of 500 °C and roasting time of 4 h, and it increased with the increase of silicon leaching rate.     

用碱性Na_2EDTA溶液从次氧化锌烟灰中选择性回收铅(英文)

采用碱性Na2EDTA溶液从次氧化锌烟灰中回收铅。探讨温度、浸出时间、Na2EDTA浓度和起始NaOH浓度对铅、锌浸出率的影响。得到最优实验条件如下:液固比5:1 mL/g、搅拌速度650 r/min、Na2EDTA浓度0.12mol/L、NaOH初始浓度0.5 mol/L、温度70°C、浸出时间120 min。在最优实验条件下,铅、锌、氟和氯的平均浸出率分别为89.92%、0.94%、62.84%和90.02%。浸出液用于电沉积铅粉。在温度为60°C、电流密度为200A/m2、H3PO4浓度为1.5 g/L、铅离子浓度不低于5 g/L时,电沉积铅粉平均电流效率大约为93%,阴极铅纯度高于98%。电沉积1 kg铅粉大约消耗0.218 kg Na2EDTA和0.958 kW·h电能。     

A NEW PROCESS FOR PRODUCING HIGH GRADE GRAPHITE FOR USE IN MANUFACTURING Mg-C BRICKS IN STEEL MAKING

Based on the characteristics of graphite ore from Panzhihua district, the present research tried to find a new technology to increase the purity of the graphite carbon from 93% to 97%. Study was conducted on leaching agents, time, temperature, stirring speed and concentration of leaching solution affecting the process. The technological parameters chosen were： L / S=3：1, reaction time 30min, 5% concentration of sulfuric acid, reaction temperature 70℃, stirring speed 200rev/min. The results of the continuing tests with one kilogram showed that the graphite content can be increased from 93% to 97%, which reaches the requirement for high grade graphite. When waste acid from coke making is used to leach graphite, the result is also satisfactory, but the treatment of waste solution containing organic substances has to be further investigated.     

Na2SO3浸出法提纯粗硒工艺研究

采用Na2SO3浸出法提纯粗硒,研究浸硒过程Na2SO3溶液浓度、浸出温度、浸出时间、搅拌速度和净化过程Na2S溶液浓度对提纯粗硒的影响,采用XRD、SEM对所得样品进行表征.结果表明,Na2SO3溶液浓度为300 g/L、浸出温度为98℃、浸出时间为2 h,搅拌速度为150 r/min、Na2S溶液浓度为2 g/L时...     

低钒转炉钢渣提钒湿法工艺的动力学研究

为了提高湿法浸出低钒钢渣中钒的浸出效率,并对湿法浸出低钒钢渣中钒提供理论依据,从动力学角度分析整个浸出过程。考察温度、液固比、硫酸质量分数和搅拌速率对浸出过程的影响。研究结果表明:在90℃、液固比为10?1以及硫酸浓度6.0mol/L时,浸取9h,低钒钢渣中钒的浸出率可达到95.3%。通过正交实验和动力学推导,得到描述浸出过程的经验方程,低钒钢渣湿法浸出钒的动力学模型为收缩核动力学模型,浸出过程的表观活化能为12.794kJ/mol,该模型表明浸出过程中的控制步骤取决于固膜扩散速率。提高温度、液固比和硫酸质量分数,均可加速钒的浸出速度,提高钒的浸出率。     

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。     

Leaching performance of Al-bearing spent LiFePO4 cathode powder in H2SO4 aqueous solution

The leaching performance and leaching kinetics of LiFePO₄ (LFP) and Al in Al-bearing spent LFP cathode powder were systematically studied. The effects of temperature (273?368 K), stirring speed (200?950 r/min), reaction time (0?240 min), acid-to-material ratio (0.1:1?1:1 mL/g) and liquid-to-solid ratio (3:1?9:1 mL/g) on the leaching process were investigated. The results show that the concentration of reactants and the temperature have a greater impact on the leaching of Al. Under the optimal conditions, leaching efficiencies of LFP and Al are 91.53% and 15.98%, respectively. The kinetic study shows that the leaching of LFP is kinetically controlled by mixed surface reaction and diffusion, with an activation energy of 22.990 kJ/mol; whereas the leaching of Al is only controlled by surface chemical reaction, with an activation energy of 46.581 kJ/mol. A low leaching temperature can effectively suppress the dissolving of Al during the acid leaching of the spent LFP cathode material.     

采用Na2SO3溶液从硒渣中选择性浸出Se及其动力学

采用SO2还原沉金后液制得硒渣,再用Na2SO3选择性浸出硒渣,使Se得到有效分离;通过研究浸出过程中Se浸出率随时间的变化,建立该反应的动力学方程,确定Na2SO3溶液浓度、液固比、搅拌速度及反应温度对Se浸出率的影响,并计算相应的表观活化能。结果表明:增加Na2SO3溶液浓度和升高反应温度可以较大幅度提高Se的浸出率,液固比和搅拌速度对浸出Se的影响较小;Na2SO3浸出Se过程为Avrami模型混合控制,其特征参数n和表观活化能E分别为0.235和20.847 kJ/mol,Se的浸出率受反应温度的影响较大。     

Leaching kinetics of antimony-bearing complex sulfides ore in hydrochloric acid solution with ozone

The leaching kinetics of Sb and Fe from antimony-bearing complex sulfides ore was investigated in HCl solution by oxidation–leaching with ozone. The effects of temperature, HCl concentration, stirring speed and particle size on the process were explored. It is found that the recoveries of Sb and Fe reach 86.1% and 28.8%, respectively, when the reaction conditions are 4.0 mol/L HCl, 900 r/min stirring speed at 85 °C with <0.074 mm particle size after 50 min leaching. XRD analysis indicates that no new solid product forms in the leaching residue and the leaching process can be described by shrinking core model. The leaching of Sb corresponds to diffusion-controlled model at low temperature (15–45 °C) and mixed-controlled model at high temperature (45–85 °C), and the apparent activation energies are 6.91 and 17.93 kJ/mol, respectively. The leaching of Fe corresponds to diffusion-controlled model, and the apparent activation energy is 1.99 kJ/mol. Three semi-empirical rate equations are obtained to describe the leaching process.     

含钒钛转炉渣氧压酸浸过程V-Ti-H2O系的电位-pH图

为了分析含钒钛转炉渣氧压酸浸过程的热力学特点,根据高温水溶液中计算标准摩尔吉布斯自由能和活度系数的经验公式,通过计算得到了氧分压0.5MPa、对应离子的质量浓度0.1mol/kg、温度60～200℃条件下V-Ti-H2O系的电位-pH图。在pH〈2的酸性条件下,可溶性V3＋,VO2＋,VO2＋的稳定区几乎全部包含在TiO2的稳定区范围内,随着温度由60℃升高到200℃,钒、钛稳定共存区对应的氧化还原电位逐渐增大,pH逐渐降低。钒、钛稳定区的共存特点从热力学角度为转炉渣采用氧压酸浸工艺通过一步酸浸将钒浸出的同时使钛富集在渣中提供了理论依据。含钒钛转炉渣的氧压酸浸实验结果表明,在浸出温度140℃、氧分压0.5MPa、粒度0.055～0.075mm、液固比15：1、浸出时间120min、搅拌速度500r/min、初酸浓度200g/L的条件下,钒的浸出率为96.87%,钛的浸出率为8.76%。钒与钛通过氧压酸浸工艺能够得到有效分离,实验结果与热力学计算结果一致。     

Optimization of operating parameters and kinetics for chloride leaching of lead from melting furnace slag

The feasibility and kinetics of lead recovery from the slag of traditional lead melting furnace using chloride leaching were investigated. The effects of operating parameters such as leaching time, NaCl concentration, FeCl₃ concentration, liquid/solid ratio, stirring rate, temperature, and particle size on recovery of lead were studied and the optimization was done through the response surface methodology (RSM) based on central composite design (CCD) model. The optimum conditions were achieved as follows: leaching time 60 min, 80 °C, stirring rate 800 r/min, NaCl concentration 200 g/L, FeCl₃ concentration 80 g/L, liquid/solid ratio 16, and particle size less than 106 μm. More than 96% of lead was effectively recovered in optimum condition. Based on analysis of variance, the reaction temperature, liquid/solid ratio, and NaCl concentration were determined as the most effective parameters on leaching process, respectively. Kinetics study revealed that chloride leaching of galena is a first-order reaction and the diffusion through solid reaction product and chemical reaction control the mechanism. The activation energy of chloride leaching of galena was determined using Arrhenius model as 27.9 kJ/mol.     

Extraction of Vanadium from LD Converter Slag by Pressure Leaching Process with Titanium White Waste Acid

研究了以钛白废酸直接加压浸出转炉钒渣提钒的工艺。矿物学研究表明：钒、钛、铁、锰、铬等金属元素形成的尖晶石是转炉钒渣的主要物相。绘制了V-Fe-H2O、V-Ti-H2O、V-Mn-H2O、V-Cr-H2O等三元系150 ℃高温电位-pH图,明确了酸浸提钒过程的热力学：在低酸度浸出提钒条件下,可溶性离子Fe2+、Fe3+、Mn2+、Cr2+、Cr3+等的热力学稳定与可溶性含钒离子的热力学稳定区重合,酸浸过程中与钒共同进入浸出液中。钛白废酸酸浸正交试验结果表明：温度和初始酸浓度是影响酸浸过程的主要因素。基于正交试验结果,进一步考察温度对浸出过程的影响,结果表明,随着酸浸温度由100 ℃升高到160 ℃的过程中,浸出渣中的钛有效富集含量在4.56%至12.0%之间变化,其他离子主要赋存于浸出液中。在较优条件下：温度 140 ℃,液固比10:1,初始酸浓度200 g·L-1,搅拌转速500 r/min,酸浸时间90 min,钒的浸出率为 96.85%     

废锂离子电池中集流体与活性物质的分离

针对废旧锂离子电池回收工艺中电极集流体的分离问题,根据集流体、活性物质、粘结剂的物理化学性质差异,对高温焙烧法、物理擦洗法和稀酸浸出-搅拌擦洗法分离集流体与活性物质进行研究。结果表明：高温焙烧与物理擦洗法都不能完全使集流体分离出来,而通过稀酸溶解-搅拌擦洗联合作用分离效果良好,在硫酸浓度为0.5 mol/L、固液比1：10、搅拌速度200 r/min、反应时间为40 min的条件下,可以实现正负极活性物质与集流体的分离,铝箔和铜箔可直接作为产品回收,只有极少部分进入浸出液,浸出渣用硫酸再浸,可以使钴、锂全部溶出,净化除杂后可回收钴和锂。     

Leaching kinetics of selenium from copper anode slimes by nitric acid-sulfuric acid mixture

The leaching kinetics of selenium from copper anode slimes was studied in a nitric acid-sulfuric acid mixture. The effects of main parameters on selenium leaching showed that the leaching rate of selenium was practically independent of stirring speed, while dependent on temperature and the concentrations of HNO₃ and H₂SO₄. The leaching of selenium includes two stages. The activation energy in the first stage is 103.5 kJ/mol, and the chemical reaction is the rate controlling step. It was almost independent of H₂SO₄ concentration and dependent on HNO₃ concentration since the empirical reaction order with respect to HNO₃ concentration is 0.5613. In the second stage, the activation energy is 30.6 kJ/mol, and the process is controlled by a mixture of diffusion and chemical reaction. The leaching of selenium was almost independent of HNO₃ concentration.     

盐酸蒸馏-磷酸三丁酯萃取法从锗煤烟尘中综合回收锗和镓

先采用二氧化锰氧化-盐酸蒸馏提取煤烟尘中的锗,经提纯后制备得到高纯二氧化锗,同时煤烟尘中的镓在蒸馏过程中充分溶解到了锗蒸馏残渣液中,过滤后镓和蒸馏后残留的锗进入到了蒸馏残液中,调节蒸馏残液的酸度至5.6mol/L,采用磷酸三丁酯的260#溶剂油溶液对镓进行萃取,蒸馏残液中的少量锗亦被同时萃取进入到有机相中,用氢氧化钠溶液进行反萃取,反萃取完后调节pH至6.5~6.7,升温至85~95℃水解,得到氢氧化镓沉淀和二氧化锗沉淀,经过滤烘干后得到含锗镓精矿。锗的蒸馏回收率可达到92.37%以上,残液中锗的萃取率可达到86.18%以上;镓的盐酸浸出率可达81.23%以上,镓的磷酸三丁酯萃取率可达98.81%以上,反萃取率可达99.11%以上,中和沉淀直收率可达93.20%以上;工艺方法具有工艺流程简短、设备简单、过程易于控制,所用生产辅料较少,是一种从煤烟尘中综合回收锗和镓的高效且经济的方法。     

低品位氧化铜矿氨-硫酸铵体系过硫酸铵氧化浸出

以过硫酸铵为氧化剂,研究低品位氧化铜矿在氨-硫酸铵体系氧化浸出工艺。讨论氨/铵离子摩尔比、总氨浓度,氨、硫酸铵和过硫酸铵的浓度,反应温度,液固比,反应时间和搅拌速度等操作条件对铜浸出的影响。结果表明:在92.8%的矿样粒径小于0.045 mm,氨、硫酸铵和过硫酸铵浓度分别为2.4、1.8和0.100 mol/L,浸出时间为90 min,温度为30℃,液固比(mL/g)为5:1,搅拌速度为500 r/min时的优化条件下,低品位铜矿的铜浸出率达87.7%。     

烧结电除尘灰中钾盐的回收及其浸出动力学

为了研究烧结电除尘灰中回收钾盐的强化浸出措施,使用ICP-AES、SEM-EDS和XRD分析技术对除尘灰的表面和内部形态,特别是钾盐的赋存形式进行分析。结果表明,该电除尘灰的主要成分是铁氧化合物,在其表面裸露吸附着一定含量的KCl晶体。水浸实验表明,该粉尘中的KCl可以通过水浸出、蒸发结晶的方式回收,其收率为18.56%。结晶产物的分析结果表明,KCl占61.21%,NaCl占13.40%,CaSO4占14.62%,K2SO4占10.86%。其水浸出动力学符合外扩散控制模型控制。强化浸出实验表明,提高浸出温度、加强搅拌、增加液固比等措施可以提高钾盐的浸出率和浸出速率。     

Leaching of cuprite through NH4OH in basic systems

Cuprite is a difficult oxide to leach under acidic conditions (for the maximum extraction of 50%). In this research, the feasibility of leaching cuprite in an ammoniacal medium was studied. The working conditions addressed here were the liquid/solid ratio (120:1–400:1 mL/g), stirring speed (0–950 r/min), temperature (10–45 °C) and NH₄OH concentration (0.05–0.15 mol/L). In addition, different ammoniacal reagents (NH₄F and (NH₄)₂SO₄) were analyzed. The experiments were performed in a 2 L reactor with a heating mantle and a condenser. The most important results were that the maximum leaching rate was obtained at pH 10.5, 0.10 mol/L NH₄OH, 45 °C, 4 h, 850 r/min and a liquid/solid ratio of 400:1, reaching a copper extraction rate of 82%. This result was related to the non-precipitation of copper in solution by the formation of copper tetra-amine. The liquid/solid ratio and stirring speed were essential for increasing the cuprite leaching. The maximum leaching rate was achieved at higher temperatures; however, significant copper leaching rate occurred at temperatures near the freezing point of water (17.9% over 4 h). Increasing NH₄OH concentration and decreasing particle size increased the cuprite leaching rate. The two ammoniacal reagents (NH₄F and (NH₄)₂SO₄) had low extraction rate of copper compared with NH₄OH. The kinetic model representing cuprite leaching was a chemical reaction on the surface. The order of the reaction with respect to the NH₄OH concentration was 1.8, and it was inversely proportional to the radius of the ore particles. The calculated activation energy was 44.36 kJ/mol in the temperature range of 10–45 °C.