Investigation of Efficient Conditions for Oxidative Roasting of Chromite Ore to Produce Potassium Chromate with KOH Additive

A method of producing potassium Chromate directly from chromite ore and KOH was investigated. The chromium recovery was >90% after roasting at 800°C for 3 hours. The KOH‐to‐ore weight ratio was 1.2:1 and the ore‐to‐recycling residue weight ratio was 1:1. Production from the roasting process was easy to leach and about 100% of Chromate could be recovered through a treatment at 30°C for 30~40min with a water/solid ratio of 5:1. The obtained Chromate solution was purified by addition of conventional compounds.     

Green metallurgical processing of chromite

A new green metallurgical process for chromite ore has been proposed and developed to solve the serious environmental problem in the traditional production process of chromate. In the new process, the oxidation of chromite ore is carried out in sub-molten potassium hydroxide at 300 °C. Compared with the traditional roasting process in a rotary kiln at 1200 °C, the oxidation and mass transfer are greatly intensified in the new process. Chemical conversion of chromium is above 99% and the recovery of chromium is raised by more than 20% while the reaction temperature is dropped by 900 °C. Thus the energy consumption is decreased. Furthermore, as no limestone or dolomite additives are required, the amount of chromium-containing residue is decreased from 2.5 tonnes to 0.5 tonnes with the production of 1.0 tonne of product in the new process. As all reactions and separating operations are performed in liquid media, pollution due to toxic dust is prevented.With the change of reaction medium, the new process achieves higher resource and energy utilization efficiency. At the same time, the new process also changes the physical and chemical properties of other components co-existing in the chromite ore. The process provides the possibility for virtually complete utilization of chromite and zero emission of residue. Thus, the clean production of chromate can be realized.     

Co-extraction of tungsten and molybdenum from refractory scheelite–powellite blend concentrates by roasting with Na2CO3 and SiO2 and leaching with water

Studies on the recovery of tungsten and molybdenum from refractory scheelite–powellite blend concentrates – mainly composed of scheelite, powellite, and fluorite – were performed using soda-silicon roasting and water leaching processes. However, significant amounts of scheelite and powellite ore are present in an intergrowth state and the application of mineral dressing is not suitable for the separation and extraction of tungsten and molybdenum from this refractory ore. The effects of roasting parameters including sodium carbonate addition, roasting temperature, roasting time, and mass ratio of SiO₂/concentrate (w_SiO2/w_conc.) and the effects of leaching parameters including leaching temperature, leaching time, and liquid-to-solid ratio on the leaching efficiency of tungsten and molybdenum were investigated. The results demonstrated the efficiency of this process for the extraction of tungsten and molybdenum from the ore. Under the optimum experimental conditions where soda-silicon roasting is performed for 2?h at 850°C with three times, the stoichiometric ratio of Na₂CO₃ (Na₂CO₃:WO₃ and Na₂CO₃:MoO₃) and w_SiO2/w_conc of 12%, and water leaching is subsequently performed for 1?h at 70°C with a liquid-to-solid ratio of 3:1, the leaching ratios of W and Mo are 98.89% and 99.41%, respectively.     

The effect of salt-phase composition on the rate of soda-ash roasting of chromite ores

The formation of a liquid phase during the early stages of the roasting reaction is a common problem in the sodium chromate manufacturing process. The molten salt phase, which is primarily constituted of a binary mixture of Na₂CrO₄ and Na₂CO₃, creates major operational problems such as the granulation and blocking of the kilns. In addition to the operational problems, it was observed that the molten salt also affects the transport of oxygen toward the reaction interface. The mechanism of the soda-ash roasting reaction has been analyzed for improving the yield of sodium chromate. It was observed that the conversion efficiency of the roasting process changed dramatically, depending on the origin and the type of the chromite ores used. Thermal and scanning electron microscopic analyses of the products of roasting were carried out to establish the reaction mechanism. It was observed that the presence of silicates in the chromite ores interferes with the formation of sodium chromate involving the binary Na₂CO₃-Na₂CrO₄ liquid. The roasting reaction proceeds in a certain crystallographic direction in the chromite spinel in the presence of a nonsilicate molten salt, whereas a complete dissolution of chromite appears to take place in the binary liquid containing silicate phases present in the ore. This article is based on a presentation given in the Mills Symposium entitled “Metals, Slags, Glasses: High Temperature Properties & Phenomena,” which took place at The Institute of Materials in London, England, on August 22–23, 2002.     

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.     

Leaching behavior of metals from a limonitic nickel laterite using a sulfation–roasting–leaching process

The leaching behavior of metals from a limonitic laterite was investigated using a sulfation–roasting–leaching process for the recovery of nickel and cobalt. The ore was mixed with water and concentrated sulfuric acid followed by roasting and finally leaching with water. Various parameters were studied including the amount of acid added, roasting temperature and time, sample particle size, addition of Na₂SO₄ and solid/liquid ratio in leaching process. More than 88% Ni, 93% Co and < 4% Fe are extracted under the determined conditions. Simultaneously, about 90% Mn and Cu, 70% Mg, 45% Al, 25% Zn, 4% Cr and Ca are extracted respectively. The pH of the leach solution is about 2. The leaching efficiency is independent of sample particle size due to decomposition of ferric sulfate formed during roasting. The roasted mass was characterized by various physico-chemical techniques such as DSC/TGA, XRD and SEM. This process provides a simple and effective way for the extraction of nickel and cobalt from laterite ore.     

Recovery of valuable metals from jarosite by sulphuric acid roasting using microwave and water leaching

In this paper, jarosite residue (JR) blended with concentrated H₂SO₄ was subjected to a process comprising microwave roasting and water leaching. The effects of H₂SO₄/JR weight ratio, microwave roasting temperature and time, water leaching conditions on the recovery of Fe, Zn, In, Cu, Cd, Ag and Pb were investigated utilising a series of experiments.

Based on energy conservation and environmental protection, optimum conditions for metals recovery from JR were determined as: H₂SO₄/JR weight ratio?=?0.36, microwave roasting temperature, 250°C; roasting time, 30?min; leaching temperature, 50°C; leaching time, 1?h; and liquid–solid ratio, 4:1 (mL/g), thus, the extraction of Fe, Zn, In, Cu, Ag and Cd were 89.4, 80.7, 85.1, 90.7, 61.3 and 48.8% respectively, while the Pb was concentrated in the final residue. Scanning electron microscope-energy dispersive spectrometer (SEM-EDS) patterns were used to characterise and analyse the transformation of valuable metals in the residue after roasting and leaching.     

硬质合金磨削废料中钨的回收利用研究

采用氧化焙烧-常压(高压)碱浸和苏打焙烧-常压水浸工艺,对硬质合金磨削废料中WC的焙烧浸出工艺进行了研究。XRD分析氧化焙砂,发现生成了难以分解破坏其结构的Fe(Al,Cu,Ti,Co)O.xSiO.2yWO3的多组分复杂固溶体物相,即使采用高温高压碱浸工艺,钨的浸出率也只有33%左右。苏打焙烧,增加碳酸钠量,提高焙烧温度,均可提高WO3浸出率(最高可达99.35%),降低其渣中含量(最低可至0.2%)。XRD分析可知,苏打焙烧可以抑制Fe(Al,Cu,Ti,Co)O.xSiO.2yWO3固溶体结构的形成。     

Selective Leaching-Recovery of Re and Mo from Out-Gas Dust of Molybdenite Roasting Furnace

Selective leaching and recovery of Mo and Re present in out-gas dust of molybdenite roasting furnace was quantitatively investigated. The effects of solid to liquid (S/L) phase ratio, sulfuric acid concentration and solution temperature on the selective extraction were determined. Considerable Mo and small amount of Re were detected in the out-gas dust of the industrial roasting furnaces. Recovery percentage of both Re and Mo increased by increasing acid concentration and solution temperature. The results also showed that the recovery percentage of Re decreased by increasing the S/L ratio. The best conditions for selective leaching was obtained as 20 min mixing in water (0 g/L sulfuric acid) with S/L ratio of 0.2 in the temperature range of 40–60 °C.     

广西某低品位软锰矿还原焙烧—酸浸工艺研究

针对广西某低品位软锰矿,探究以煤为还原剂进行还原焙烧,焙烧矿采用硫酸浸出的工艺条件,通过单因素实验,考察了还原剂用量、焙烧温度与时间的组合、浸出酸量、时间、液固比、搅拌强度及浸出温度对浸出结果的影响。实验表明:在取300g软锰矿进行实验时,配煤11%、在温度750℃下焙烧60min的焙烧矿,在理论酸量、固液比5∶1、搅拌强度300r/min、常温下浸出45min后,锰的浸出率可达到95.57%的良好指标。     

The soda-ash roasting of chromite ore processing residue for the reclamation of chromium

Sodium chromate is produced via the soda-ash roasting of chromite ore with sodium carbonate. After the reaction, nearly 15 pct of the chromium oxide remains unreacted and ends up in the waste stream, for landfills. In recent years, the concern over environmental pollution from hexavalent chromium (Cr⁶⁺) from the waste residue has become a major problem for the chromium chemical industry. The main purpose of this investigation is to recover chromium oxide present in the waste residue as sodium chromate. Cr₂O₃ in the residue is distributed between the two spinel solid solutions, Mg(Al,Cr)₂O₄ and γ-Fe₂O₃. The residue from the sodium chromate production process was analyzed both physically and chemically. The compositions of the mineral phases were determined by X-ray diffraction (XRD), scanning electron microscopy (SEM), and electron probe microanalysis (EPMA). The influence of alkali addition on the overall reaction rate is examined. The kinetics of the chromium extraction reaction resulting from the residue of the soda-ash roasting process under an oxidizing atmosphere is also investigated. It is shown that the experimental results for the roasting reaction can be best described by the Ginstling and Brounshtein (GB) equation for diffusion-controlled kinetics. The apparent activation energy for the roasting reaction was calculated to be between 85 and 90 kJ·mol⁻¹ in the temperature range 1223 to 1473 K. The kinetics of leaching of Cr³⁺ ions using the aqueous phase from the process residue is also studied by treating the waste into acid solutions with different concentrations.     

铬铁矿球团焙烧固结特性

本文系统研究铬铁矿球团的焙烧固结特性.结果表明:预热时间对于预热球强度影响不大,在预热时间为10 min时,随着预热温度的提高,预热球强度和氧化率呈直线型增加,适宜温度为1050℃,此时预热球强度可达每个400 N以上;与传统铁矿球团相比,铬铁矿球团焙烧所需的温度高,焙烧时间为10 min时,焙烧温度从1250℃提高到1350℃,球团强度从每个1078 N提高到1973 N.在铬铁矿球团预热和焙烧过程中,铬尖晶石(Fe,Mg)(Cr,Fe,Al)₂O₄氧化生成富镁的(Fe,Mg)(Cr,Fe,Al)₂O₄和铬铁铝复合氧化物(Cr,Fe,Al)₂O₃,当温度高于1000℃时,(Cr,Fe,Al)₂O₃新相生成,其主要以环状分布在颗粒外层,颗粒内部为针状与(Fe,Mg)(Cr,Fe,Al)₂O₄形成交织结构,降低Cr/Fe比或升高焙烧温度均有助于(Cr,Fe,Al)₂O₃向颗粒外层富集和再结晶长大,有利于球团的固结,提高球团强度.      

低品位红土镍矿焙烧-碱浸过程中硅的转化

针对目前红土镍矿碱法处理过程中存在的问题提出工艺改进,研究低品位红土镍矿焙烧活化-碱浸过程中含硅矿物的转化。考察了焙烧温度对红土镍矿活性的影响,探索了红土镍矿经焙烧后碱浸过程中温度、时间、搅拌强度、液固比以及碱初始质量浓度对硅转化的影响。结果表明,红土镍矿经650 °C焙烧2 h后,活性得到明显提高,红土镍矿经焙烧后采用初始质量浓度为60 g/L的碱溶液,在搅拌速度为400 r/min、浸出温度为140 °C、液固比为5∶1的条件下浸出120 min,硅的转化率可达89.42%。     

Dissolution kinetics of malachite in ammonia/ammonium carbonate leaching

The leaching of oxide copper ore containing malachite, which is the unique copper mineral in the ore, by aqueous ammonia solution has been studied. The effect of leaching time, ammonium hydroxide, and ammonium carbonate concentration, pH, [NH₃]/[NH₄⁺] ratio, stirring speed, solid/liquid ratio, particle size, and temperature were investigated. The main important parameters in ammonia leaching of malachite ore are determined as leaching time, ammonia/ammonium concentration ratio, pH, solid/liquid ratio, leaching temperature, and particle size. Optimum leaching conditions from malachite ore by ammonia/ammonium carbonate solution are found as ammonia/ammonium carbonate concentrations: 5 M NH₄OH+0.3 M (NH₄)₂CO₃; solid/liquid ratio: 1:10 g/mL; leaching times: 120 min; stirring speed: 300 rpm; leaching temperature: 25 °C; particle size finer than 450 μm. More than 98% of copper was effectively recovered. During the leaching, copper dissolves as in the form of Cu(NH₃)₄⁺² complex ion, whereas gangue minerals do not react with ammonia. It was determined that interface transfer and diffusion across the product layer control the leaching process. The activation energy for dissolution was found to be 15 kJ mol⁻¹.     

Leaching Ni and Co from Saprolitic Laterite Ore by Employing Atmospheric Acid Leaching Solution with High Concentration of FeCl 3 at Mild Conditions

Limonitic and saprolitic laterite ores are used to produce Ni by employing different kinds of processes. Acid leaching is an energy-saving process for pure Ni metal production. The high concentration of Fe in atmospheric acid leaching solution caused difficulties in metal-ions separation. In this work, saprolitic laterite ore was leached by atmospheric acid leaching solution of limonitic laterite ore at moderate temperatures. Conditions affecting the leaching of valuable metals and the conversion ratio of Fe were investigated. The results showed that optimal output was obtained after 1.5 h of leaching at 150°C with 1.0 liquid/solid ratio (volume/weight). X-ray diffraction analysis and mineral liberation analysis indicate that some nickel was adsorbed by leached residues, resulting in the loss of nickel. The physiochemical properties of leached residues could be changed due to the presence of Al and Cr. The decrease in zeta potential and the increase in specific surface area resulted in the increase of Ni adsorption.     

某碳质金精矿富氧焙烧—氯化浸出试验研究

某碳质金精矿直接氰化浸出金浸出率很低,小于30%,为进一步提高金浸出率,针对碳质金精矿性质,进行了富氧焙烧—氯化浸出试验研究。结果表明:与常规氧化焙烧相比,富氧焙烧降低了焙烧温度,缩短了焙烧时间;富氧焙烧最佳焙烧温度550℃~600℃,氧气体积分数50%,焙烧时间2.0 h,在此条件下,碳、硫去除率均在95%以上;焙砂采用M-NaCl氯化浸出,在最佳浸出条件为固液比1∶6,浸液pH=3,浸出剂用量8 kg/t,试样粒度62~75μm,浸出时间4 h时,金浸出率可达92.50%,相对于试样直接氯化浸出时有显著提高;表明富氧焙烧—氯化浸出工艺是可行的。     

Extraction of chromium from Egyptian chromite ores

Abstract

The roasting process of Egyptian chromite middling product with lime and soda was successfully performed to produce sodium chromate with a chromium recovery of 97% and an aluminum recovery of 81%. The optimum conditions were determined. Separation of alumina from the chromatealuminate solution was achieved by controlling the pH of the solution at 7.6 by carbon dioxide gas. Separation of sodium dichromate crystals was feasible and the production of .chromium oxide from the dichromate through solid state reduction by carbon is studied.

Résumé

Un produit mixte de chromite égyptienne a été grillé avec de la chaux et du carbonate de sodium pour donner du chromate de sodium avec un recouvrement en chrome de 97%et en aluminium de 81% Les conditions optimales ont été déterminées. L'extraction de l'alumine à partir de la solution chromate-aluminite a ete effectuee en maintenant le pH de la solution a 7.6 par du gaz carbonique. Il a été possible d'ioder les cristaux de dichromate de sodium et la préparation de l'oxyde de chrome à partir du dichromate par carbothermie a été étudiée.     

基于响应曲面法的含锆废盐中ZrO2的回收工艺

含锆废盐是粗四氯化锆提纯工艺产生的主要固废,含有大量氧化锆和可溶性氯化物,通过水浸-焙烧处理可回收其中的氧化锆。采用响应曲面法优化水浸工艺,当液固比为8∶1 mL/g,搅拌时间为60 min,浸出次数为3次时,浸出渣中氧化锆含量的预测值为95.2%。同时对浸出渣进行焙烧处理,当焙烧温度为600 ℃,时间为60 min时,焙烧产物中氧化锆含量为96.23%。采用SEM、XRD、XRF对浸出渣和焙烧产物的微观形貌和成分进行表征分析,研究结果表明,浸出渣和焙烧产物的主要成分为氧化锆,焙烧产物中氧化锆的含量相比浸出渣提高约1%,且晶粒相比浸出渣表现更优。      

Studies on Sulfating Roasting Process for Mianning Bastnasite and Baotou Mixed RE Concentrate Ore

Some processes of sulfating roasting and water leaching of crude Mianning RE concentrate ore, of fine Mianning RE concentrate ore, of Baotou RE concentrate ore and of their mixture were investigated.The result shows that the mixture of Mianning and Baotou RE concentrate ore has the optimum leaching rate and rate of recovery when the mixture ratio is 1:4.The recovery rate of the mixture is higher by 14.76% than that of crude Mianning RE concentrate ore, by 5.0 % than that of Mianning fine RE concentrate ore and by 2.4 % than that of Baotou RE concentrate ore.     

某难浸金矿低温焙烧预处理试验

贵州某难浸金矿原矿直接浸出率约15%,通常需在650℃以上进行焙烧预处理。为提高该金矿的浸出率,在加入氧化钙的条件下进行低温焙烧预处理试验,考察焙烧温度、氧化钙用量、焙烧时间对金浸出率的影响。结果表明,在下述最佳预处理条件下金浸出率可达85%以上:焙烧温度600℃,氧化钙用量17.5%,焙烧时间1.5h。