Molybdenite-limestone oxidizing roasting followed by calcine leaching with ammonium carbonate solution

Oxidizing roasting of molybdenite with lime can significantly reduce SO2pollution compared with the traditional roastingwithout lime.However,the calcine is subsequently leached by sulfuric acid,resulting in serious equipment corrosion and abundantnon-recyclable CaSO4slag.In this work,a novel process,in which the molybdenite was roasted with CaCO3followed by(NH4)2CO3solution leaching,was proposed to improve the art of lime roasting-sulfuric acid leaching.Oxidizing roasting of molybdenite withCaCO3was investigated through thermodynamic calculation,thermogravimetric analysis and roasting experiments.The results showthat the products of the oxidizing reaction of MoS2in the presence of CaCO3and O2are CaSO4,CaMoO4and CO2at573-1000K.The MoS2conversion rate achieves approximately99%and the sulfur-retained rate attains approximately95%with aCaCO3-to-MoS2molar ratio of3.6at500°C for1h by adding5%mineralizer A(mass fraction).The leaching results show that theleaching rate of Mo reaches98.2%at85°C for7h with a(NH4)2CO3concentration of600g/L and a liquid-solid ratio of10mL/g.The results presented are potential to develop a novel cleaner technique for ammonium molybdate production.

Oxidation roasting of molybdenite concentrate

In order to investigate the oxidation roasting of molybdenite concentrate in pure oxygen atmosphere, experiments at 673, 723, 773, 873 and 973 K were carried out. The phase transitions and morphology evolutions of the samples obtained at different temperatures after reacting for different time were analyzed by X-ray diffraction（XRD） and scanning electron microscopy（SEM）. The results showed that molybdenite concentrate was oxidized directly to Mo O3 in pure oxygen atmosphere. There were remarkable changes of the morphologies of products with the increase of the roasting temperature. It was also found that sintering phenomenon occurred during the roasting process in pure oxygen when the temperature was above 873 K. The composition of sintered sample was mainly comprised of Mo O3 and some unreacted Mo S2.

Removal of magnesium and calcium from electric furnace titanium slag by H_3PO_4 oxidation roasting-leaching process

H3PO4 oxidation roasting followed by HCl acid leaching was proposed to remove magnesium and calcium from electric furnace titanium slag containing 3.12% MgO and 0.86% CaO. XRF, XRD and SEM techniques were used to characterize the composition, mineral phase component and microstructure of the titanium slag. The H3PO4 oxidation thermodynamic, mineral phase transformation, microstructure, element distribution in titanium slag during H3PO4 oxidation process and leaching process were investigated. The thermodynamic analysis indicated that H3PO4 could promote the decomposition of MgTi2O5 and CaSiO3. The results indicated that H3PO4 could effectively promote the transformation of titanium-bearing mineral to rutile and enrich the impurities in MxTi（3-x）O5 into phosphate which could be removed by acid leaching process. Under the studied conditions, the leaching rates of magnesium and calcium reached 94.68% and 87.19%, respectively. The acid leached slag containing 0.19% MgO and 0.13% CaO（mass fraction） was obtained.

Response surface optimization of process parameters for removal of F and Cl from zinc oxide fume by microwave roasting

Microwave was applied to roasting the zinc oxide fume obtained from fuming furnace for the removal of F and Cl. The effects of important parameters, such as roasting temperature, holding time and stirring speed, were investigated and the process conditions were optimized using response surface methodology (RSM). The results show that the effects of roasting temperature and holding time on the removal rate of F and Cl are the most significant, and the effect of stirring speed is the second. The defluorination rate reaches 92.6% while the dechlorination rate reaches 90.2%, under the process conditions of roasting temperature of 700 °C, holding time of 80 min and stirring speed of 120 r/min. The results indicate that the removal of F and Cl from fuming furnace production of zinc oxide fumes using microwave roasting process is feasible and reliable.

Extracting reaction mechanism analysis of Zn and Si from zinc oxide ore by NaOH roasting method

The orthogonal test was used to optimize the reaction conditions of roasting zinc oxide ore with NaOH aiming to comprehensively utilize zinc oxide ore.The optimized reaction conditions were molar ratio of NaOH to zinc oxide ore 6:1,roasting temperature 450°C,holding time 150 min.The molar ratio of NaOH to zinc oxide ore was the most predominant factor affecting the extraction ratios of zinc oxide and silica.The mineral phase transformations were investigated by testing the phases of specimens obtained at different temperatures.The process was that silica reacted with molten NaOH to form Na_2SiO_3 at first,then transformed into Na_4SiO_4 with temperature rising.ZnCO_3 and its decomposing product ZnO reacted with NaOH to form Na_2ZnO_2.Na_2ZnSiO_4was also obtained.The reaction rate was investigated using unreacted shrinking core model.Two models used were chemical reaction at the particle surface and diffusion through the product layer.The results indicated that the reaction rate was combine-controlled by two models.The activation energy and frequency factor were obtained as 24.12 k J/mol and 0.0682,respectively.

Kinetics of oxidation roasting of molybdenite with different particle sizes

The kinetics of oxidation roasting of molybdenum concentrate was studied by differential thermal−gravimetric experiments and non-isothermal analysis methods.The results show that high temperature is beneficial for oxidation of molybdenum concentrate.The initial oxidation temperature of the molybdenum concentrate is 450℃,and the rapid oxidation occurs above 500℃.The oxidation process conforms to the unreacted shrinking nucleus model.The early stage of the oxidation is controlled by chemical reaction with the apparent activation energy of 123.180 kJ/mol,while the later stage is controlled by internal diffusion with the apparent activation energy of 80.175 kJ/mol.Moreover,the oxidation rate is closely related to particle size of the concentrate.The smaller the particle size is,the larger the oxidation rate is.

Thermodynamics of chromite ore oxidative roasting process

To explicate the thermodynamics of the chromite ore lime-free roasting process, the thermodynamics of reactions involved in this process was calculated and the phrases of sinter with different roasting times were studied. The thermodynamics calculation shows that all the standard Gibbs free energy changes of the reactions to form Na2CrO4, Na2O-Fe2O3, Na2O·Al2O3 and Na2O3 SiO2 via chromite ore and Na2CO3 are negative, and the standard Gibbs free energy changes of the reactions between MgO, Fe2O3 and SiO2 released from chromite spinel to form MgO-Fe2O3 and MgO·SiO2 are also negative at the oxidative roasting temperatures （1 173 1 473 K）. The phrase analysis of the sinter in lime-free roasting process shows that Na2O·Fe2O3, Na2O·Al2O3 and Na2O·SiO2 can be formed in the first 20 min, but they decrease in contents and finally disappear with the increase of roasting time. The final phase compositions of the sinter are Na2CrO4, MgO·Fe2O3, MgO·SiO2 and MgO. The results indicate that Na2CrO4 can be formed easily via the reaction ofNa2CO3 with chromite ore. Na2O·Fe2O3, Na2O-Al2O3 and Na2O·SiO2 can be formed as intermediate compounds in the roasting process and they can further react with chromite ore to form Na2CrO4. MgO released from chromite ore may react with iron oxides and silicon oxide to form stable compounds of MgO·Fe2O3 and MgO·SiO2, respectively.

Roasting oxidation behaviors of ReS_2 and MoS_2 in powdery rhenium-bearing, low-grade molybdenum concentrate

The oxidation roasting process of molybdenum concentrate has significant advantages in industrial applications.However,utilization of low-grade mineral has many problems because it is more complex than the standard concentrate.In this study,the oxidation behaviors of powdery rhenium-bearing low-grade molybdenum concentrate were investigated through thermodynamic calculation,roasting experiments,thermogravimetric analysis,and phase analysis.The results obtained show that oxidation of MoS2 begins at 450℃,and MoO3 reacts with metal-oxide forming molybdate at 600℃.Finally,MoO3 can be dissolved in ammonia with a maximum content of approximately 80%.The volatile speed of Re was considerably slower than the oxidation speed of MoS2 because the nonvolatile products ReO2 and ReO3 were generated in reactions among MoS2,SO2,and Re2O7.The final volatilization rate of Re was almost 70%.This study determined the problems related to the roasting of low-grade molybdenum concentrate,which lays the scientific foundations for subsequent enhancement of molybdenum and rhenium extraction.

Features and mechanisms of self-sintering of molybdenite during oxidative roasting

Roasting experiments were carried out with pure reagents as raw materials.The self-sintering behaviors of molybdenite(MoS_(2))during oxidation were investigated by thermodynamic calculation,XRD,SEM-EDS and high-temperature in situ analysis.The results indicate that the desulfurization efficiency of MoS_(2) pellet decreases with the increase of roasting temperature from 600 to 700℃,owing to the expansion of sintered area.At the very beginning of roasting,sintered layer can be rapidly formed and cover the pellet surface on the windward side,and meanwhile,MoO_(2)and Mo_(4)O_(11) intensively appear and constitute the sintered layer together with MoO_(3).Moreover,it is proven that MoO_(3)-rich products containing low-valence molybdenum oxides have low melting points,thus easy to be melted during the occurrence of exothermic reaction between MoS_(2) and O_(2),which leads to the sintering of materials.

Oxidation and roasting characteristics of artificial magnetite pellets

Compared with natural magnetite concentrate, artificial magnetite with more lattice defects and higher activity tends to be oxidized. And the artificial magnetite pellet at the temperature of 400℃ has the oxidation degree approaching to natural magnetite concentrate pellet fired at 1000℃. Besides, two kinds of pellets displayed quite different roasting characteristics. When preheated at the same temperature for the same period of time, natural magnetite concentrate pellet and artificial magnetite concentrate pellet need to be roasted at the temperature of 1100℃ and 1250℃, respectively, for 25 min to reach the compressive strength of 3000 N per pellet. When roasted at the same temperature of 1200℃, natural magnetite pellet and artificial magnetite pellet need to be roasted for 15 min and 30 min, respectively, to reach the compressive strength over 3000 N per pellet. It can be seen from the test that artificial magnetite pellet has a faster oxidation, resulting in the high porosity in the produced pellet, and it requires a roasting process at higher temperature for a longer time to reach the desired compressive strength for industrial production.

Recovery of Ferric Oxide from Bayer Red Mud by Reduction Roasting-Magnetic Separation Process

A great amount of red mud generated from alumina production by Bayer process was considered as a low-grade iron ore with a grade of 5wt% to 30wt% iron.We adopted the reduction roastingmagnetic separation process to recover ferric oxide from red mud.The red mud samples were processed by reduction roasting,grinding and magnetic separating respectively.The effects of different parameters on the recovery rate of iron were studied in detail.The optimum techqicalparameters were proposed with 700 ℃roasting for 20 min,as 50wt% carbon and 4wt% additive were added.The experimentalresults indicated that the iron recovery and the grade of totaliron were 91% and 60%,respectively.A novelprocess is applicable to recover ferric oxide from the red mud waste fines.

Preparation of sodium molybdate from molybdenum concentrate by microwave roasting and alkali leaching

The preparation process of sodium molybdate has the disadvantages of high energy consumption,low thermal efficiency,and high raw material requirement of molybdenum trioxide,in order to realize the green and efficient development of molybdenum concentrate resources,this paper proposes a new process for efficient recovery of molybdenum from molybdenum concentrate and preparation of sodium molybdate by microwave-enhanced roasting and alkali leaching.Thermodynamic analysis indicated the feasibility of oxidation roasting of molybdenum concentrate.The effects of roasting temperature,holding time,and power-to-mass ratio on the oxidation product and leaching product sodium molybdate (Na_(2)MoO_(4)·2H_(2)O) were investigated.Under the optimal process conditions:roasting temperature of 700℃,holding time of 110 min,and power-to-mass ratio of 110 W/g,the molybdenum state of existence was converted from MoS_(2) to Mo O3.The process of preparing sodium molybdate by alkali leaching of molybdenum calcine was investigated,the optimal leaching conditions include a solution concentration of 2.5 mol/L,a liquid-to-solid ratio of 2 mL/g,a leaching temperature of 60℃,and leaching solution termination at pH 8.The optimum conditions result in a leaching rate of sodium molybdate of 96.24%.Meanwhile,the content of sodium molybdate reaches 94.08wt%after leaching and removing impurities.Iron and aluminum impurities can be effectively separated by adjusting the pH of the leaching solution with sodium carbonate solution.This research avoids the shortcomings of the traditional process and utilizes the advantages of microwave metallurgy to prepare high-quality sodium molybdate,which provides a new idea for the highvalue utilization of molybdenum concentrate.

Zinc extraction from zinc oxidized ore using(NH4)2SO4 roasting−leaching process

An improved method of(NH4)2SO4 roasting followed by water leaching to utilize zinc oxidized ores was studied.The operating parameters were obtained by investigating the effects of the molar ratio of(NH4)2SO4 to zinc,roasting temperature,and holding time on zinc extraction.The roasting process followed the chemical reaction control mechanism with the apparent activation energy value of 41.74 kJ·mol^−1.The transformation of mineral phases in roasting was identified by X-ray diffraction analysis combined with thermogravimetry–differential thermal analysis curves.The water leaching conditions,including the leaching temperature,leaching time,stirring velocity,and liquid-to-solid ratio,were discussed,and the leaching kinetics was studied.The reaction rate was obtained under outer diffusion without product layer control;the values of the apparent activation energy for two stages were 4.12 and 8.19 kJ·mol^−1.The maximum zinc extraction ratio reached 96%while the efficiency of iron extraction was approximately 32%under appropriate conditions.This work offers an effective method for the comprehensive use of zinc oxidized ores.

Comparison of the mechanisms of microwave roasting and conventional roasting and of their effects on vanadium extraction from stone coal

Experiments comparing microwave blank roasting and conventional blank roasting for typical vanadium-bearing stone coal from Hubei Province in central China, in which vanadium is present in muscovite, were conducted to investigate the effects of roasting tempera- ture, roasting time, H2SO4 concentration, and leaching time on vanadium extraction. The results show that the vanadium leaching efficiency is 84% when the sample is roasted at 800℃ for 30 min by microwave irradiation and the H2SO4 concentration, liquid/solid ratio, leaching temperature, and leaching time are set as 20vo1%, 1.5：1 mL.g-1, 95℃, and 8 h, respectively. However, the vanadium leaching efficiency achieved for the sample subjected to conventional roasting at 900℃ for 60 min is just 71% under the same leaching conditions. Scanning electron microscopy （SEM） analysis shows that the microwave roasted samples contain more cracks and that the particles are more porous compared to the conventionally roasted samples. According to the results of X-ray diffraction （XRD） and Fourier-transform infrared （FTIR） analyses, neither of these roasting methods could completely destroy the mica lattice structure under the experimental conditions; however, both methods deformed the muscovite structure and facilitated the leaching process. Comparing with conventional roasting, microwave roasting causes a greater deformation of the mineral structure at a lower temperature for a shorter roasting time.

Mechanochemical sulfidization of a mixed oxide-sulphide copper ore by co-grinding with sulfur and its effect on the flotation efficiency

Mechanochemical sulfidization of a mixed sulfide/oxide copper ore by co-grinding with sulfur and additives including Mg(NO3)2 and Fe(NO3)3 salts and iron,aluminum and magnesium powders was investigated for the first time.Also,the influence of sulfidization during the wet-milling process was examined on the separation efficiency and recovery of copper in detail.The results demonstrated that co-grinding with sulfur solely had the best flotation performance at the value of 0.5 wt.%and it was attributed to the possible existence of S\\O bonding on copper oxides surfaces.In addition,adding magnesium nitrate salt,magnesium powder,iron nitrate salt and aluminum powder as additive associated with 0.5 wt%sulfur into ball milling caused the flotation improvement at the amounts of 0.2 wt%,0.2 wt%,0.5 wt%and 0.5 wt%,respectively.Also,the effect of grinding time and sulfidization pH with 0.5 wt%sulfur solely was determined and pH s of 7.5 to 8.5 gave the best results.The highest recovery(75.76%)and separation efficiency(63.44%)were achieved at pH of 7.5 and 8.5,respectively.

Sulphidization flotation for recovery of lead and zinc from oxide-sulfide ores

A new flowsheet was developed to recover the valuable minerals from oxide or oxide-sulfide ores of lead and zinc. The flowsheet consisted of flotation of sulfide minerals, desliming and sulphidization-flotation of oxide minerals. The corresponding reagent system and techniques to the flowsheet were investigated. Batch and continuous tests show that the dosage of sodium sulfide, temperature, and collector type are main affecting factors on the recovery of smithsonite and cerussite. For the flotation of cerussite, there is an appropriate dosage of sodium sulfide at which the recovery reaches its maximum value. The required sodium sulfide for smithsonite flotation is higher than that for cerussite and the recovery of smithsonite flotation increases with the dosage of sodium sulfide at low level and becomes insensitive at high dosage. The appropriate temperature for smithsonite and cerussite flotation is found to be 2540℃. Amines are found to be the effective collectors for the flotation of smithsonite after sulphidization. Investigation also shows that desliming prior to sulphidization-flotation is essential to the effective recovery of smithsonite and cerussite, and the desliming process of two-stage hydrocyclon is well feasible and effective for the treatment of lead-zinc oxide ores. A further treatment on the cerussite flotation concentrate by shaking table is proposed to obtain higher lead grade.

Single Step Solid-Solid Reaction Scheme for the Synthesis of Cobalt Sulphide-Oxide Nanoparticles in Polymer Matrix

At present, there is considerable interest in polymer-metal chalcogenides/oxides based nano-composites on account of their tunable optical, magnetic, electronic and catalytic properties. Here in, we report a simple single step approach for the in-situ synthesis of combined cobalt sulphide/cobalt oxide in polyphenylenesulphide (PPS) polymer matrix. We have illustrated the suitability of this methodology by reacting commonly available cobalt precursors with engineering thermoplastic, PPS. The cobalt precursor was homogeneously mixed with PPS in the molar ratios of 1:1, 1:5, 1:10, 1:15, respectively, followed by heating the mixtures obtained at the melting temperature of the polymer (285?C) for six hours. The resultant products were characterized by X-ray Diffractometry (XRD), Field-Emission Scanning Electron Microscopy (FESEM), High Resolution Transmission Electron Microscope (HRTEM), Diffuse Reflectance Spectroscopy (DRS) technique and Fourier Transformation Infra red Spectroscopy (FTIR). Formation of mixed phases viz., sulphide and oxide of cobalt within modified PPS matrix was confirmed by XRD. The resultant nanoparticles of cobalt sulphide and cobalt oxide embedded in the PPS matrix showed crooked and chunk morphology. The optical properties of the resultant nanocomposites indicate the shift in the absorption hump due to nanoscale size effect.

QUANTUM CHEMICAL MECHANISM ON THE SURFACE OXIDATION AND FLOTATION OF SULPHIDE MINERALS

The surfaces of galena and pyrite,the adsorption of oxygen,and the reactions of ethyl xanthate on the sulfides were studied using quantum chemical calculations.In addition,the surface electron structures of the minerals were discussed.According to the results,the mechanisms of ionic exchange reaction and dixanthogen adsorption for the xanthate notation of galena and pyrite were explained.The important role of oxygen in collector and collectorless flotation processes was also investigated.

Bioleaching of a kind of alkaline mixed copper oxide and sulphide mineral

We conducted two-stage acidification-bioleaching experiments to probe the feasibility of bioleaching for a kind of mixed alkaline copper oxide and sulphide mineral. We used the uniform design method for data analysis and experimental optimization, with initial pH value, pulp density, inoculation of bacteria and ferrous iron concentration selected as the influential factors. Polynomial regression shows that the four factors sequentially influence the copper recovery by 14.430%, 8.555%, 1.982% and 3.895%. Acid equilibrium in the bioleaching system is mainly influenced by alkaline gangue content, chemical reactions and bacterial activity. A maximal portion of refractory copper extracted reaches 71.08%. The dynamic analysis of copper recovery indicates that bioleaching goes through a lag leaching phase, prime leaching phase and leaching stationary phase corresponding to the growth phases of bacteria. Compared with the predicted value of 80.87%, the confirmatory experiment observes a 78.21% copper recovery under the optimal conditions of pH of 1.5, pulp density of 5%, bacteria inoculation of 30% and initial ferrous iron concentration of 9 g L-1. Results suggest that bioleaching is technically feasible to improving total copper recovery.

The Machanism of Bacterial Oxidation of Sulphides

The bacteria used in the experiment are Thiobacillus Ferrooxidans separated from acidic mine water in sulphide deposits. The chemoautotrophic bacteria can act directly on sulphides and accelerate the oxidation of sulphides. The experiment shows that the bacteria, as an important microbial factor of gold's supergenous enrichment within the oxidized zones of sulphide deposits, are helpful to dissolve gold and silver in ferric sulphate. In the bacterial oxidation process, the precipitation of goethite is concerned both with the lower activity of ferric ions and with the existence of carbonates in solution. Meanwhile, the acid-resisting and oxidizing ability of the bacteria will certainly lead up to a microbial way of treating the acidic mine water.