石煤钠化焙烧料酸浸动力学

研究了石煤钠化焙烧料硫酸浸出过程中,浸出剂初始浓度、搅拌速度和浸出温度对浸出率的影响,并对浸出过程动力学进行了分析.结果表明:浸出剂初始浓度和浸出温度对钒浸出率有显著影响,搅拌速度对钒浸出率影响不大;该浸出过程符合核收缩模型,与化学反应控制动力学方程式相吻合,浸出反应的表观活化能为50.88 kJ·mol^-1,浸出过程控制步骤为化学反应控制.     

磷酸钠-氢氧化钠浸出氟化钙动力学研究

针对现行主流氢氧化钠和碳酸钠浸出低品位白钨矿存在浸出剂用量大,浸出率低的弊端,提出磷酸钠-氢氧化钠-氟化钙协同高效浸出低品位白钨矿的新工艺.氟化钙作为固体氟源,在浸出过程中首先发生溶解并释放出游离的氟离子,再参与白钨矿的浸出反应生成氟磷酸钙沉淀.因此,氟化钙的溶解速率对白钨矿浸出速率有较大影响.文章使用等浸出率法,对磷...     

Kinetics of manganese ore reduction by carbon monoxide

A step has been made in the direction of understanding the fundamental chemical processes taking place inside electric are furnaces producing manganese alloys. The reduction of higher manganese oxides to MnO by carbon monoxide has been studied in the temperature range 700 °C to 1100 °C. A topochemical pattern with a single shrinking core inside the ore particles has been observed in most cases. It has been found that the reduction of some manganese silicates (braunite minerals) is influenced by reaction interface kinetics, whereas the reduction rate of manganese oxides (bixbyite and hausmannite) is mostly determined by product shell pore diffusion. Sintering kinetics and the extent of natural porosity determine the product shell pore diffusivity. As the melting point of the reaction product is approached, rapid sintering leads to a decrease in diffusivity.     

Kinetics of reductive leaching of manganese oxide ore with molasses in nitric acid solution

The kinetics of reductive leaching of manganese from a low grade manganese ore in dilute nitric acid in the presence of molasses is examined. The rate is controlled by diffusion through the “product” layer composed of the associated minerals. The leaching process follows the kinetic model 1 − 2/3X − (1 − X)^2/3 = kt with an apparent activation energy of 25.7 kJ/mole. It is concluded that the concentration of HNO₃ and molasses as well as temperature are the main factors influencing the leaching rate. The results indicate a reaction order of 1.2 for HNO₃ concentration and 1.9 for molasses concentration.     

Kinetics of leaching of lead sulfate in sodium chloride solutions

A powder dissolution method is used to study the kinetics of leaching of lead sulfate from a lead cake in the temperature range 298–353 K and the concentration range 50–270 g/l of a sodium chloride solution. The results obtained are used to calculate the apparent activation energy (E _a = 12.41 kJ/mol); its value indicates that the process of leaching of lead sulfate from a lead cake occurs in a diffusion zone. This process is shown to be described by a first-order kinetic equation.     

Kinetics of the solid-state carbothermic reduction of wessel manganese ores

Reduction of manganese ores from the Wessel mine of South Africa has been investigated in the temperature range 1100 °C to 1350 °C with pure graphite as the reductant under argon atmosphere. The rate and degree of reduction were found to increase with increasing temperature and decreasing particle sizes of both the ore and the graphite. The reduction was found to occur in two stages: (1) The first stage includes the rapid reduction of higher oxides of manganese and iron to MnO and FeO. The rate control appears to be mixed, both inward diffusion of CO and outward diffusion of CO₂ across the porous product layer, and the reaction of carbon monoxide on the pore walls of the oxide phase play important roles. The values of effective CO-CO₂ diffusivities generated by the mathematical model are in the range from 2.15 x 10⁻⁵ to 6.17 X 10⁻⁵ cm².s⁻¹ for different ores at 1300 °C. Apparent activation energies range from 81. 3 to 94.6 kJ/kg/mol. (2) The second stage is slower during which MnO and FeO are reduced to mixed carbide of iron and manganese. The chemical reaction between the manganous oxide and carbon dissolved in the metal phase or metal carbide seems to be the rate-controlling process The rate constant of chemical reaction between MnO and carbide on the surface of the impervious core was found to lie in the range from 1.53 x 10⁻⁸ to 1.32 x 10⁻⁷ mol . s⁻¹ . cm⁻². Apparent activation energies calculated are in the range from 102.1 to 141.7 kJ/kg/mol. Formerly Doctoral Student, Department of Metallurgy and Materials Engineering, University of the Witwatersrand, Johannesburg,     

甲醛-硫酸亚铁协同还原浸出低品位氧化锰矿

在硫酸介质中以甲醛-硫酸亚铁为还原剂协同还原软锰矿,考察了甲醛-硫酸亚铁摩尔比、温度、反应时间、转速、硫酸浓度等因素对锰、铝的浸出率及溶液中铁和有机残留甲酸的影响.采用单因素实验获得较佳的还原工艺条件,采用HPLC测定溶液中的甲酸.结果表明,在固定转速为200 r/min、液固比为8 ml/g时,最佳反应条件为:甲醛-硫酸亚铁摩尔比1∶3(甲醛1.5 mol)、浸出时间3 h、硫酸浓度3 mol/L、温度90℃.在该条件下重复实验,锰的平均浸出率为93.51%,铝的平均浸出率为33.08%,铁的浓度为23.07 mol/L,甲酸浓度为0.001 g/L.     

Kinetics of manganese oxide reduction from basic slags by carbon dissolved in liquid iron

The reduction of manganese oxide from a basic slag by carbon dissolved in liquid iron is a slow reaction, failing to approach equilibrium closely in 20 hr. Furthermore, the rate of stirring has no apparent effect on the reaction rate. This identifies the rate-controlling step as a chemical reaction at the interface. Only the model for the reactionO ²⁻ =O + 2e ⁻ gave a consistent interpretation as the melt geometry, and concentration of manganese oxide and carbon were varied. The rate constant for this reaction was found to be 1.28 × 10⁻⁵ mole per sq cm per min at 155O°C. The effect of temperature is substantial with a calculated energy of activation for the system of 25 kcal per mole. Formerly Graduate Student, The University of Michigan This paper is based on a portion of a thesis submitted by W. L. DAINES in partial fulfillment of the requirements for the degree Doctor of Philosophy at The University of Michigan.     

锌锰干电池的同槽电解及锰还原浸出研究

同槽电解回收锌锰电池中金属发现二氧化锰回收率很低,提出了在中性条件用锌粉还原浸出二氧化锰的方法.实验结果表明,液固比对浸出率影响最大,液固比为2:1最好,浸出时间为120 min最好,50℃后浸出温度对浸出率影响不大,搅拌速度对浸出率几乎无影响.锌粉还原浸出MnO2动力学特征符合收缩球模型.     

降低金精矿浸出氰化钠用量的研究

通过对氰化浸出过程氰化钠消耗情况分析,在进一步研究金精矿氰化浸出机理的基础上,提出了采用碳酸铵、氢氧化铵混合试剂降低氰化钠用量的新工艺.工业应用实践表明,该工艺节约氰化钠用量50%以上,生产成本降低1/3,浸渣银品位降低到15g/t,银浸出率提高11%.     

Extraction of manganese from low-grade manganese ores by FeSO4 leaching

An investigation is reported on manganese recovery from the low-grade ores of Orissa, India by ferrous sulphate leaching. Various process parameters including agitation rate, quantity of ferrous sulphate, solid-liquid ratio, temperature, leaching time and particle size were studied in detail and optimum conditions were established on a laboratory scale. The effect of acid additions during ferrous sulphate leaching of the ore were also studied. The data were compared with the results obtained when the ore was treated with synthetic pickle liquors and the effect of prereduction of the ore was also investigated.     

利用双氧水为还原剂湿法浸出电解锰阳极泥中锰的研究

电解锰阳极泥是生产电解金属锰时阳极产生的副产物,其中含有大量锰、铅等资源.如何高效浸出电解锰阳极泥中的锰是实现其资源化利用的关键,本研究提出了一种H₂SO₄-H₂O₂浸出体系强化电解锰阳极泥中锰浸出的新方法,研究了H₂O₂和H₂SO₄用量、反应温度、反应时间以及固液比对电解锰阳极泥中锰浸出率的影响.研究结果表明,在阳极泥与H₂O₂质量比1∶0.8、阳极泥与H₂SO₄质量比1∶0.9、反应温度45℃、固液质量比1∶10条件下浸出反应15 min,锰的浸出率可达97.23%,浸出渣中Pb的质量分数高达53.71%.浸出机理分析表明,酸性条件下电解锰阳极泥中锰氧化物被H₂O₂还原浸出,浸出液中Mn主要以MnSO₄存在,浸出渣中Pb主要以PbSO₄...     

Separation and recovery of scandium and titanium from red mud leaching liquor through a neutralization precipitation-acid leaching approach

Selective recovery of scandium and titanium from red mud leaching solution was achieved through neutralization precipitation followed by acid leaching approach.In the neutralization precipitation section,the effects of pH,temperature and reaction time on metal precipitation efficiency were investigated.Under the optimal co ndition,the precipitation efficiencies of scandium and titanium were 93.74% and 99.47%,respectively.In the acid leaching section,the effects of acidity,temperature,reaction time,and raffinate to acid solution ratio on leaching efficiency were investigated.Under the optimal condition,the leaching efficiencies of scandium and titanium were 99.97% and 5.44%,respectively.The loss of scandium and titanium were 6.3% and 5.9%,respectively.Compared with the traditional extraction procedure of scandium in red mud,this method could separate titanium from scandium effectively,which is beneficial for the purification of the products and improvement of value of the metal.     

Extraction of scandium from red mud by acid leaching with CaF_2 and solvent extraction with P507

The extraction of Sc by acid leaching with CaF_2 and solvent extraction with P507 from red mud was proposed.The influence of acid leaching and solvent extraction on recovery of Sc was investigated.The CaF_2 can obviously improve the leaching efficiency of Sc and reduce the acid consumption.The leaching efficiency of Sc increases from 74% to 92% and the dosage of acid reduces under suitable conditions by adding 5% CaF_2.The minerals in red mud can easily be decomposed and leached into the acid solution with CaF_2 through analysis of XRD pattern.The particles of red mud become smaller and multihole.The Sc can be selectively extracted with 10% P507 at the pH value of 0.1 from the acid leaching solution.More than 98% of Sc and less than 10% of Al and Fe are extracted.The SC_2O_3 with purity of 99% is obtained after the process of reverse extraction with NaOH,H_2SO_4 dissolution,precipitation by oxalic acid and roasting at 750℃.     

Sulphuric acid leaching of cobalt-bearing manganese wad

Manganese wad from the Mt. Tabor region of Queensland, Australia, contains in excess of 1% cobalt with lesser amounts of nickel and copper. As such, it presents a significant cobalt resource, particularly in terms of Australia's cobalt requirements. Because of the remote location and the lack of infrastructural facilities, potential processing routes need to be kept as simple as possible. There are many similarities between terrestrial manganese wads and deep-sea manganese nodules, and so routes developed for the processing of the latter should be applicable to the wad. Of these routes, sulphuric acid leaching at atmospheric pressure and ambient temperature appears to have considerable potential.This paper describes the results obtained with sulphuric acid leaching using the pug, agitated tank and column techniques. Even with excess acid, dissolution of the cobalt, copper and nickel from the wad was found to be extremely slow, with reaction times of weeks and months required for more than 50% dissolution. No conditions could be found where selective dissolution of cobalt, copper and nickel over manganese occurred. Comparative tests were carried out with Pacific Ocean deep-sea manganese nodules.Because of the low leaching rate found with sulphuric acid at atmospheric pressure and ambient temperature, commercial exploitation of the Mt. Tabor wad with this leachant under these conditions seems unlikely. Previously reported work indicates that a better approach will be to use sulphur dioxide as the leachant, even though this is a considerably more expensive leachant.     

Kinetics of leaching selenium from Ni-Mo ore smelter dust using sodium chlorate in a mixture of hydrochloric and sulfuric acids

The kinetics of leaching selenium from Ni-Mo ore smelter dust in H₂SO₄-HCl-H₂O system was investigated. The effects including leaching temperature and time, particle size of the smelting dust, stirring speed, acid concentration and the coefficient β (the molar ratio of sodium chlorate to selenium in the smelter dust) on leaching of selenium were studied. The results indicated that the leaching of selenium increased sharply with the increase of temperature. The leaching of selenium reached 98% at 95 °C and stirring speed of 350 rpm for 150 min with the particle size of − 0.15 mm, initial [H⁺] concentration of 8 mol/L, the solid/liquid ratio of 1:5 g/mL and the coefficient β of 3.33. The leaching process was controlled by the surface chemical reaction and the kinetics of leaching selenium from Ni-Mo ore smelter dust followed the model of “shrinking core”. The apparent activation energy of leaching selenium was determined to be 44.4 kJ/mol, which was consistent with the values of activation energy reported for the surface chemical reaction control. The kinetics equation of leaching selenium from Ni-Mo ore smelter dust was expressed as , which coincided with the experimental results.     

Cobalt extraction and concentration from manganese wad by leaching and precipitation

Manganese wad often contains significant amounts of nickel, cobalt and copper, and it may represent a future source of these metals. The mineralogy and extractive metallurgy of a number of Australian wads are being examined with regard to producing a high-grade non-ferrous metal sulphide suitable for further processing. This paper describes the general approach taken, and some specific results obtained when a cobalt-rich wad from Queensland was leached with sulphur dioxide. The rate of dissolution is affected by the feed composition, particle size and sulphur dioxide flowrate. Under the conditions used, > 90% Co, Cu and Ni extraction is achieved in about 8 h. The Co, Cu and Ni can be readily precipitated from the clarified leachate by injection of hydrogen sulphide. The product contains about 60% (Co + Cu + Ni) with less than 0.1% Mn. As leaching is non-selective with respect to manganese, it will probably be necessary to recover and recycle the sulphur dioxide associated with the dissolved manganese when the ground wad is processed on a commercial scale.     

Separation of zinc from spent zinc-carbon batteries by selective leaching with sodium hydroxide

Physical methods such as crushing and sieving followed by magnetic separation steps were applied to separate non-magnetic material (88.4 wt.%) from zinc-carbon spent batteries. The oversize material was processed by eddy current separation to recover zinc sheet, carbon rods, and plastics. The undersize fraction (− 2.36 mm) with a metal composition of 15.5% Zn, 17.5% Mn, and 1.4% Fe was used for the leaching experiments under different conditions such as concentration of sodium hydroxide, temperature, agitation speed, and pulp density. Selective leaching using 4 mol dm^− 3 NaOH at 100 g dm^− 3 solid/liquid ratio, 80 °C and 200 rpm gave a zinc extraction of 82% and a manganese extraction of less than 0.1%. An overall zinc recovery was about 88.5% by the processes described in this study.     

Process optimization of rare earth and aluminum leaching from weathered crust elution-deposited rare earth ore with compound ammonium salts

In order to intensify the leaching process of rare earth(RE) and reduce the impurities in the leachate, ammonium chloride(NH4Cl) and ammonium nitrate(NH4NO3) were mixed as a compound leaching agent to treat the weathered crust elution-deposited RE ore. Effects of molar ratio of NH~+_4Cl and NH_4NO_3, ammonium(NH_4) concentration, leaching agent pH and flow rate on the leaching process of RE were studied and evaluated by the chromatographic plate theory. Leaching process of the main impurity aluminium(Al) was also discussed in detail. Results showed that a higher initial ammonium concentration in a certain range could enhance the mass transfer process of RE and Al by providing a driving force to overcome the resistance of diffusion. pH almost had no effects on the mass transfer efficiency of RE and Al in the range of 4 to 8. The relationship between the flow rate and height equivalent to a theoretical plate(HETP) could fit well with the Van Deemter equation, and the flow rate at the lowest HETP was determined. The optimum conditions of column leaching for RE and Al were 1:1(molar ratio) of NH_4Cl and NH_4NO_3, 0.2 mol/L of ammonium concentration, pH 4–8 of leaching agent and 0.5 mL /min of flow rate. Under this condition, the mass transfer efficiency of RE was improved, but no change was observed for Al compared with the most widely used ammonium sulfate. Moreover, the significant difference value(around 20 mL) of retention volume at the peak concentration between RE and Al provided a possibility for their separation. It suggested the potential application of the novel compound leaching agent(NH_4Cl/NH_4NO_3). It was found that the relative concentration of RE in the leachate could be easily obtained by monitoring the pH of leachate.