Effects of mechanical activation on the kinetics of terbium leaching from waste phosphors using hydrochloric acid

The effect of mechanical activation(MA) on the kinetics of terbium(Tb) leaching from waste phosphors using hydrochloric acid was investigated. Leaching kinetics, such as apparent reaction rate, activation energy and reaction order, were determined using the shrinking-core model and the Arrhenius equation. Results obtained from experiments with different concentrations of HCl and under different leaching temperatures were used for the determinations. The impacts of factors such as rotational speed, HCl concentration and leaching temperature on the leaching rate of Tb were also discussed. The results showed that MA could dramatically increase the leaching rate of Tb from waste phosphors, and the apparent reaction rate(k_(ap)) of leaching was accelerated as well. For inactivated waste phosphors, the apparent activation energy(E_(ap)) was 52.82±3.95 kJ/mol, indicating that the rate-controlling step of the leaching process was the chemical reaction. The E_(ap) dropped to 25.96 ±3.90 kJ/mol and 10.96±2.79 kJ/mol when the waste phosphors were mechanically activated at rotational speeds of 400 and 600 r/min, respectively; the leaching process transformed to a hybrid(chemical-reaction and diffusion) control process, and even a reagent-diffusion(through the product layer) control process. The apparent reaction order for Tb leaching from 400 r/min-activated waste phosphors was 2.49±0.11, and it decreased to 1.16±0.17 when the rotational speed of 600 r/min was used. Kinetics results indicated that MA could make Tb leaching occur spontaneously, and the activation intensity of waste phosphors was strengthened with higher rotational speed.     

Decomposition of mixed rare earth concentrate by NaOH roasting and kinetics of hydrochloric acid leaching process

A new clean extraction technology for the decomposition of Bayan Obo mixed rare earth concentrate by NaOH roasting is proposed.The process mainly includes NaOH roasting to decompose rare earth concentrate and HCl leaching roasted ore.The effects of roasting temperature,roasting time,NaOH addition amount on the extraction of rare earth and factors such as HCl concentration,liquid-solid ratio,leaching temperature and leaching time on the dissolution kinetics of roasted ore were studied.The experimental results show that when the roasting temperature is 550℃ and the roasting time is 60 min,the mass ratio of NaOH:rare earth concentrate is 0.60:1,the concentration of HCl is 6.0 mol/L,the ratio of liquid to solid(L/S) 6.0:1.0,and the leaching temperature 90℃,leaching time 45 min,stirring speed 200 r/min,and the extraction of rare earth can reach 92.5%.The relevant experimental data show that the process of HCl leaching roasted ore conforms to the shrinking core model,but the control mechanism of the che mical reaction process is different when the leaching temperature is different.When the leaching temperature is between 40 and 70℃,the chemical reaction process is controlled by the diffusion of the product through the residual layer of the inert material.The average surface activation energy of the rare earth element is E_a=9.96 kJ/mol.When the leaching temperature is 75-90℃,the chemical reaction process is controlled by the interface transfer across the product layer(product layer interface mass transfer) and diffusion.The average surface activation energy of rare earth elements is E_a=41.65 kJ/mol.The results of this study have certain significance for the green extraction of mixed rare earth ore.     

Preparation of enriched cerium oxide from bastnasite with hydrochloric acid by two-step leaching

Roasted with sodium carbonate, bastnasite (Ln(Ce)CO₃F) was converted to calcine containing rare earth oxides (REO), among them cerium, which existed mainly as CeO₂. The calcine was first leached with diluted hydrochloric acid, which resulted in a sludge with the enriched cerium (IV) dioxide. The sludge was further leached with a concentrated hydrochloric acid, adding hydrogen peroxide as a reducing agent; in this manner, the enriched cerium tri-chloride (CeCl₃) was prepared. The optimal technological parameters are suggested as follows: first, the hydrochloric acid concentration, the leaching temperature, the ratio of solid to liquid, and the leaching time are 1 mol/L, 60°C, 1:20, and 90 minutes, respectively; second, the hydrochloric acid concentration, the dosage of hydrogen peroxide in every 5 g of the sludge, the ratio of solid to liquid, the leaching temperature, and the leaching time are 6 mol/L, 6 mL, 1:20, 50°C, and 90 minutes, respectively. As a result, the cerium-enriched rare earth (RE) solution, containing over 95 pct cerium oxide, is obtained, which is in turn available for use in preparing a kind of polishing powder containing high cerium. The total recovery of cerium was 91 pct (85.3 pct, in the second step).     

Preparation of enriched cerium oxide from bastnasite with hydrochloric acid by two-step leaching

Roasted with sodium carbonate, bastnasite (Ln(Ce)CO₃F) was converted to calcine containing rare earth oxides (REO), among them cerium, which existed mainly as CeO₂. The calcine was first leached with diluted hydrochloric acid, which resulted in a sludge with the enriched cerium (IV) dioxide. The sludge was further leached with a concentrated hydrochloric acid, adding hydrogen peroxide as a reducing agent; in this manner, the enriched cerium tri-chloride (CeCl₃) was prepared. The optimal technological parameters are suggested as follows: first, the hydrochloric acid concentration, the leaching temperature, the ratio of solid to liquid, and the leaching time are 1 mol/L, 60 °C, 1:20, and 90 minutes, respectively; second, the hydrochloric acid concentration, the dosage of hydrogen peroxide in every 5 g of the sludge, the ratio of solid to liquid, the leaching temperature, and the leaching time are 6 mol/L, 6 mL, 1:20, 50 °C, and 90 minutes, respectively. As a result, the cerium-enriched rare earth (RE) solution, containing over 95 pct cerium oxide, is obtained, which is in turn available for use in preparing a kind of polishing powder containing high cerium. The total recovery of cerium was 91 pct (85.3 pct, in the second step).     

高铁铝土矿分步硫酸高温浸出的性能

采用分步酸浸处理高铁铝土矿的新工艺,研究了高温酸浸过程主要工艺条件对一段酸浸渣中铝浸出性能的影响.结果表明:高温酸浸可以实现一段酸浸渣中铝的高效浸出,在浸出温度180℃、硫酸质量分数60％、液固质量比10:1的酸浸条件下,铝浸出率可达到95％以上.     

高磷鲕状铁矿酸浸脱磷

研究了酸浸处理高磷铁矿脱磷及其影响因素.实验用鄂西鲕状高磷矿Fe的质量分数为51.7%,P的质量分数约0.5%,S的质量分数为0.34%.通过硫酸浸出,浸出矿中磷的质量分数降低至0.07%左右,而铁损只有0.18%,S的质量分数为0.35%,满足钢铁生产的要求.通过扫描电镜观察和能谱分析表征了高磷矿中磷的脱除,在实验酸度下能明显看出磷灰石溶解,而铁相基本不反应,并得到了热力学计算证明.实验确定了最佳的酸浸条件:浸出时间1 h,液固比100mL:8 g,酸度0.2 mol·L^-1,振荡频率150 Hz.通过微波加热预处理,高磷铁矿中产生微裂纹,增加了矿石的比表面积,但是这并没有明显促进酸浸脱磷的进行.通过补酸的方式循环利用酸浸液处理高磷铁矿能得到较好的脱磷效果,脱磷率稳定在80%,能有效减少酸耗、保护环境.     

Reaction kinetics for the leaching of MnO2 by sulfur dioxide

The reaction kinetics for the leaching of MnO₂ by sulfur dioxide have been studied using studied using monosize particles at dilute solid phase concentrations in a stirred reactor to determine the important chemical factors which govern the kinetic response of the system. The conclusion that the reaction rate is limited by a chemical reaction at the MnO₂ surface is supported by: (1) an apparent activation energy of 35.9 kJ/mol (8.6 kcal/mol), (2) the inverse first-order relationship between the rate constant k and the initial particle diameter, (3) the independence of the reaction rate on stirring speed and, more importantly, the magnitude of the calculated reaction velocity constant (～10^?3 cm/s compared to predicted mass-transfer coefficients of 10^?2 cm/s), and (4) the one-half order dependence of the reaction rate with respect to the SO₂ concentration.The rate-limiting step is considered to be an electrochemical surface reaction, a conclusion which is substantiated by electrode half cell potential measurements. Using the Butler-Volmer equation, a theoretical analysis of the electrochemical reaction resulted in the following rate equation:

which was consistent with the experimental results.     

Preparation of sodium aluminate from the leach liquor of diasporic bauxite in concentrated NaOH solution

A new process to produce monosodium aluminate hydrates (MAH) by fast crystallization from the leach liquor of a diasporic bauxite in concentrated NaOH solution is presented. The crystallization of MAH was carried out easily compared to the precipitation of gibbsite and the effect of agitation, initial concentration of sodium aluminate, seed amount and the presence of red mud were systematically studied in a batch crystallizer. The apparent kinetics of crystallization followed a second order rate law with an apparent activation energy for MAH crystallization of 38.0 kJ/mol which implies a surface-diffusion controlled mechanism. X-Ray diffraction and scanning electron microscopy identified the structure of MAH as Na₂O·Al₂O₃·2.5H₂O with a flake crystal morphology. The molar ratio α of Na₂O to Al₂O₃ in the MAH products was < 1.2 after a simple wash by dilute sodium aluminate.     

Extraction kinetics of alunite in sulfuric acid and hydrochloric acid

Extraction kinetics of alunite in sulfuric acid and hydrochloric acid were studied in a batch reactor. The effects of reaction temperature, acid concentration, particle size, calcination temperature, calcination time and KF/Al₂O₃ molar ratio on the extraction process were investigated. Experimental studies were carried out in the ranges of 35–95 °C for reaction temperature, 0.25–3.0 M for sulfuric acid, 0.5–0.6 M for hydrochloric acid, 76–182 μm for average particle size, 100–900 °C for calcination temperature, 15–60 min for calcination time and 0.15–0.90 for KF/Al₂O₃ molar ratio. The calcination temperature was the most important parameter affecting the extraction process followed by reaction temperature, particle size and acid concentration. Others had less effect. It was determined that the extraction process is controlled by diffusion through a product layer. The activation energies of the processes were found to be 19.1 and 18.5 kJ/mol for sulfuric acid and hydrochloric acid, respectively. The apparent rate constants were similar for both acids and found to be a function of acid concentration as C^0.33 and particle size r^−0.8.     

Thermodynamics and kinetics of lutetium extraction with HEH(EHP) in hydrochloric acid medium

Solvent extraction has been the most widely used technique for rare earths separation. In this study, thermodynamics and kinetics of lutetium extraction with HEH(EHP) in hydrochloric acid medium were investigated. The extraction mechanism and the relevant parameters were determined by experiment research which can guide the practical extraction process. The data indicated that chloride ion had no effect on lutetium extraction, the rate constant increased when stirring speed was enhanced. Effects of temperature, HEH(EHP) concentration, acidity, and chloride concentration were also studied. Thickness of the diffusion film was also calculated to be 4.66×10~(–3) cm at 150 r/min.     

盐酸法富集低品位锰矿及酸介质高值再生工艺

针对印度尼西亚某低品位锰矿,提出盐酸法富集与酸介质高值再生的工艺。借助X射线衍射分析,光学显微镜和电子显微镜等表征方法进行工艺矿物学分析。结果表明:该锰矿矿物组成简单,主要由方解石、软锰矿和少量菱锰矿、褐铁矿、高岭石等组成。筛析结果显示该锰矿粒度越小,锰含量相对越高。粗碎后以2 mm筛孔的筛子过筛可得到锰质量分数为33.32%的锰中矿。锰中矿盐酸直接浸出的最佳条件为:浸出pH 3.0、浸出时间1.5 h、搅拌转速200 r·min?1、液固比4∶1 mL·g?1,此条件下产出的锰精矿品位为54.50%,钙质量分数为0.57%。常温下盐酸再生可产出二水硫酸钙晶须,其长径比可达50以上。再生盐酸返回浸出锰中矿,产出的锰精矿品位为52.16%,钙质量分数为1.39%,验证了该工艺流程的可行性。X射线衍射分析、扫描电镜及能谱分析结果显示产出的锰精矿主要组成成分为软锰矿,杂质为少量褐铁矿、高岭石等。酸介质循环时杂质将逐渐积累,当镁离子质量浓度积累到96.74 g·L?1时,采用水解沉淀法进行除杂。      

Preparation of synthetic rutile by hydrochloric acid leaching of mechanically activated Panzhihua ilmenite

Dissolution of mechanically activated Panzhihua ilmenite in hydrochloric acid for the preparation of synthetic rutile was investigated. Both the dissolution and its coupled titanium hydrolysis were greatly enhanced by the mechanical pretreatment. Increases in the lattice strain and surface area of ilmenite induced by energetic ball milling were responsible for the enhanced dissolution. The rapid hydrolysis led to formation of quantities of the nanosized primary particles, giving rise to a solid/liquid separation problem. The 15 min milled ilmenite, however, yielded an easy-to-filter hydrolysate due to formation of porous, micron-sized, secondary particles during the dissolution. The crystallization and aggregation behaviour of the primary particles were probably related to both the surface property of the un-reacted solid and the ferric ion concentration in solution. The technology for preparation of synthetic rutile was systematically investigated. The optimum milling and dissolution conditions were as follows: milling in air for 15 min, hydrochloric acid concentration 20%, initial reaction temperature 100 °C, ilmenite/20% acid mass ratio 1 g:5.5 g, reaction time ≥ 6 h. The synthetic rutile prepared under the optimum conditions contained 92% TiO₂ and 2.1% Fe₂O₃ as well as combined CaO and MgO of 0.28%. The results demonstrate that the mechanical pretreatment can take the place of the traditional high temperature pretreatment of ilmenite and avoid the dissolution being conducted under pressurized condition.     

Reaction mechanism for the acid ferric sulfate leaching of chalcopyrite

The acid ferric sulfate leaching of chalcopyrite, CuFeS₂ + 4Fe⁺³ = Cu⁺² + 5Fe⁺² + 2S⁰ was studied using monosize particles in a well stirred reactor at ambient pressure and dilute solid phase concentration in order to obtain fundamental details of the reaction kinetics. The principal rate limiting step for this electrochemical reaction appears to be a transport process through the elemental sulfur reaction product. This conclusion has been reached in other investigations and is supported by data from this investigation in which the reaction rate was found to have an inverse second order dependence on the initial particle diameter. Furthermore, the reaction kinetics were found to be independent of Fe⁺³, Fe⁺², Cu⁺² and H₂SO₄ in the range of additions studied. The unique aspect of this particular research effort is that data analysis, using the Wagner theory of oxidation, suggests that the rate limiting process may be the transport of electrons through the elemental sulfur layer. Predicted reaction rates calculated from first principles using the physicochemical properties of the system (conductivity of elemental sulfur and the free energy change for the reaction) agree satisfactorily with experimentally determined rates. Further evidence which supports this analysis includes an experimental activation energy of 20 kcal/mol (83.7 kJ/mol) which is approximately the same as the apparent activation energy for the transfer of electrons through elemental sulfur, 23 kcal/ mol (96.3 kJ/mol) calculated from both conductivity and electron mobility measurements reported in the literature. formerly Metallurgy Graduate Student, University of Utah.     

Simultaneous leaching of a deep-sea manganese nodule and chalcopyrite in hydrochloric acid

Hydrochloric acid leaching of chalcopyrite and a manganese nodule, in combination, was studied using powder samples. Chalcopyrite, which does not dissolve well in HCl, was effectively leached in the presence of a manganese nodule at 3 to 4 M HCl. The rates of dissolution of metal values from the nodule were also enhanced in the presence of chalcopyrite. Dissolution was found to occur through three routes: (1) the galvanic interaction between CuFeS₂ and MnO₂, (2) the action of the Fe³⁺/Fe²⁺ redox couple, and (3) the action of Cl₂ gas generated from the MnO₂-HCl reaction on CuFeS₂. The last route appeared to make the major contribution to the dissolution. The combined leaching of a nodule and chalcopyrite appears to be promising from both technical and economic points of view.     

A kinetic model for the hydrochloric acid leaching of kaolinite clay in the presence of fluoride ions

The dissolution of kaolinite clay in hydrochloric acid solutions has been carried out in the presence of fluoride ions. Leaching in the presence of fluoride ions activates the clay for leaching, making higher extractions possible at lower roasting and leaching temperatures. The activation energy for the leaching of clay calcined at 540°C is decreased from 71 kJ/mol to 23 kJ/mol in the presence of fluoride ions. Dissolution in the presence of fluoride appears to fit a second-order reaction mechanism.     

The selective leaching of uranium,vanadium and phosphorus from phosphate ore with hydrochloric acid

The selective leaching of uranium, vanadium, and phosphorus from phosphate ore may be useful in by-product recovery. Experimental results have shown that it is possible to preferentially remove uranium from phosphate ore using dilute HCl (0.05 M). 93 pct of the uranium is leached within 90 min, leaving 94 pct of the phosphorus and 82 pct of the vanadium unattacked. Phosphorus may then be removed by increasing the pH. The apparent activation energies and orders for the leaching reactions were found. For uranium, the apparent order with respect to H⁺ is 1.05 and the apparent activation energy is 7750 J. The apparent order for the leaching of the vanadium minerals with respect to H⁺ is 1.93 and the apparent activation energy is 12800 J. The phosphorus reaction has an apparent order, with respect to H⁺, of 1.98 and an apparent activation energy of 10200 J. The uranium readsorbs at longer times. The readsorption reaction is a function of temperature, particle size, and H⁺ concentration. Two methods of selectivity analysis were used in the analysis of the data-end point analysis and initial rate analysis.