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.

Reclamation of CO_(2) sodium silicate used sands by steam leaching

The bond film on the surface of the CO_(2) sodium silicate used sands is not easy to decompose,therefore,it is difficult to reclaim used sands.A new reclamation method of CO_(2) sodium silicate used sands was developed by steam leaching,which can reduce the water consumption of reclamation and improve the removal effect of sodium silicate bond film.Firstly,the leaching effect of the sodium silicate sands after 20/200/400/600/800/1,000°C heat preservation treatment was simulated.Furthermore,the influence of the leaching time on the removal effect of the sodium silicate bond film was studied.Finally,the casting properties of the reclaimed sands after the leaching reclamation treatment were tested.The results show for simulated used sands after 30 min of steam leaching,the removal ratio of the alkali exceeds 84.1%,the removal ratio of silicate is 86.2%,and the removal ratio of carbonate is 93.6%.The removal rate of alkali,silicate and carbonate is relatively low in the leaching time of 30-50 min.Considering the reclamation effect and cost,the leaching time is controlled in 30 min.Water consumption is only 60%of the mass of used sands for 30 min steam leaching,while it is 200%for wet reclamation.Morphological analysis shows that most of the hazardous substances in the used sands are removed in 30 min steam leaching,and the reclaimed sands surface after steam leaching in 50 min is as smooth as new sands.After 30 min of steam leaching,the alkali removal effect of the factory used sands can reach 81.5%,the water consumption by the steam leaching reclamation is 58%of the mass of the used sand,which is similar to the result of simulated used sands.The performance of reclaimed sands obtained after 30 min steam leaching is better than that of new sands when the amount of sodium silicate added is 6%of the mass of the reclaimed sands and the CO_(2) blowing time is 15 s:the 24 h ultimate compressive strength of reclaimed sands is 5.6 MPa(equated with new sands),and the collapsibility compressive strength is 5.2 MPa,which is lower than the collapsibility compressive strength of new sands(7.7 MPa).This indicates that the reclamation of CO_(2) sodium silicate used sands by steam leaching is a feasible method.

Leaching kinetics of selenium from selenium-tellurium-rich materials in sodium sulfite solutions

The leaching kinetics of selenium from selenium-tellurium-rich materials in sodium sulfite solutions was studied. The morphologies of selenium-tellurium-rich materials are mainly spheroid and columnar bodies and the range of particle size of selenium-tellurium-rich materials is between 17.77μm and 56.58μm, which contain 41.73%selenium and 40.96%tellurium. The ranges of experimental elements are 126-315 g/L of sodium sulfite concentration, 100-400 r/min of agitation speed, 23-95 ℃ of reaction temperature, 7：1-14：1 of liquid-solid ratio and 17.77-56.58μm of average particle size. The results show that the leaching rate increases with increasing the sodium sulfite concentration, agitation speed, reaction temperature or liquid-solid ratio and the leaching rate decreases with increasing the particle size. The reaction temperature has the significant effects on the selenium leaching rate which increases from 21%to 67%with increasing temperature from 23 ℃ to 95 ℃. The experimental data agree quite well with the Avrami model for leaching, with model parameter of 0.235 and apparent activation energy of 20.847 kJ/mol.

Leaching kinetics of scheelite concentrate with sodium hydroxide in the presence of phosphate

The reaction kinetics for the leaching of low-grade scheelite concentrate was investigated in an autoclave with sodium hydroxide in the presence of phosphate. The effects of stirring speed (300-600 r/min), reaction temperature (353-383 K), sodium hydroxide concentration (1.69-6.76 mol/L) and phosphate concentration (0.68-1.69 mol/L) on the WO3 dissolution ratio were studied. The results showed that the WO3 dissolution ratio was practically independent of stirring speed, while it increased with increasing the reaction temperature, and the concentrations of sodium hydroxide and phosphate. The experimental data were consistent with the shrinking core model, with a surface chemical reaction as the leaching rate-determining step. The apparent activation energy was calculated as 49.56 kJ/mol, and the reaction orders with respect to the concentrations of sodium hydroxide and phosphate were determined as 0.27 and 0.67, respectively. The kinetics equation of the leaching process was established.

Enhancing the leaching effect of an ion-absorbed rare earth ore by ameliorating the seepage effect with sodium dodecyl sulfate surfactant

Ameliorating the problem of low leaching efficiency,long leaching period,and high agent consumption should be studied to efficiently exploit ion-absorbed rare earth ore resources.In this study,the surfactant sodium dodecyl sulfate(SDS) is used to enhance the leaching effect of an ion-absorbed rare earth ore by ameliorating the seepage effect for the first time.The effects of surfactant concentration,leaching agent dosage,solution flow velocity,and solution pH on the leaching rate were explored,and the mechanism of SDS was discussed.Under the optimum conditions,the addition of a small amount of SDS(mass fraction0.04%) can increase the leaching rate by about 5%,shorten the leaching period,and reduce the consumption of the leaching agent.SDS significantly ameliorates the seepage effect of the ore body by reducing the surface tension of the leaching agent and ameliorating the wettability of the mineral surface.This effect is the main factor that improves the leaching efficiency.DFT(density functional theory) calculation results show that SDS can react with rare earth ions,which reduces the adsorption strength on clay mineral surfaces.Hence,rare earth ions are easily exchanged by ammonium ions,and mass transfer is enhanced.

Microemulsion leaching of vanadium from sodium-roasted vanadium slag by fusion of leaching and extraction processes

The fusion of the leaching and purification processes was realized by directly using microemulsion as the leaching agent.The bis-(2-ethyhexyl)phosphoric acid(DEHPA)/n-heptane/NaOH microemulsion system was established to directly leach vanadates from sodium-roasted vanadium slag.The effect of the leaching agent on the leaching efficiency was investigated,in addition to the molar ratio of H_(2)O/NaDEHP(W),DEHPA concentration,solid/liquid ratio,stirring time,and leaching temperature.In optimal situations,the vanadium leaching efficiency reaches 79.57%.The X-ray diffraction characterization of the leaching residue and the Raman spectrum of the microemulsion before and after leaching demonstrate the successful entry of vanadates from the sodium-roasted vanadium slag into the microemulsion.The proposed method successfully realizes the leaching and purification of vanadates in one step,thereby greatly reducing production costs and environmental pollution.It also offers a new way to achieve the green recovery of valuable metals from solid resources.

Kinetics and leaching behaviors of aluminum from pharmaceutical blisters in sodium hydroxide solution

A hydrometallurgical process was developed for recycling pharmaceutical blisters.Leaching aluminum from pharmaceutical blisters using sodium hydroxide(NaOH) solutions was investigated with respect to leaching behaviors and kinetics.AL9(34) orthogonal design of experiments suggests that the most significant factor is NaOH concentration followed by temperature and leaching time.Factorial experiments demonstrate that the leaching rate of aluminum increases with increasing of the factors.The optimum conditions are temperature of 70 °C,leaching time of 20 min,NaOH concentration of 1.25 mol/L,liquid-to-solid mass ratio of 15:1 and agitation speed of 400 r/min.Under optimum conditions,the leaching rate is up to 100%,implying that aluminum and polyvinyl chloride(PVC) plastic in pharmaceutical blisters are separated completely.Kinetics of leaching aluminum is best described by the product layer diffusion control model,and the activation energy is calculated to be 19.26 kJ/mol.

Sodium sulfide leaching of low-grade jamesonite concentrate in production of sodium pyroantimoniate

Sodium sulfide leaching of a low-grade jamesonite concentrate in the production of sodium pyroantimoniate through the air oxidation process and the influencing factors on the leaching rate of antimony were investigated. In order to decrease the consumption of sodium sulfide and increase the concentration of antimony in the leaching solution, two-stage leaching of jamesonite concentrate and combination leaching of high-grade stibnite concentrate and jamesonite concentrate were used. The experimental results show that the consumptions of sodium sulfide for the two-stage leaching process and the combination leaching process are decreased by 20% and 60% compared to those of one-stage leaching process respectively. The final concentrations of antimony in the leaching solutions of both processes are above 100g/L.

Gold-leaching performance and mechanism of sodium dicyanamide

In this work,sodium dicyanamide(SD)was used as a leaching reagent for gold recovery,and the effects of the SD dosage and solution pH on the gold-leaching performance were investigated.A gold recovery of 34.8%was obtained when SD was used as the sole leaching reagent at a dosage of 15 kg/t.In the presence of a certain amount of potassium ferrocyanide(PF)in the SD solution,the gold recovery was found to increase from 34.8%to 57.08%.Using the quartz crystal microbalance with dissipation(QCM-D)technique,the leaching kinetics of SD with and without PF were studied.The QCM-D results indicate that the gold-leaching rate increased from 4.03 to 39.99 ng·cm^(–2)·min^(–1) when the SD concentration was increased from 0 to 0.17 mol/L,and increased from 39.99 to 272.62 ng·cm^(–2)·min^(–1) when 0.1 mol/L of PF was used in combination with SD.The pregnant solution in the leaching tests was characterized by X-ray photoelectron spectroscopy and electrospray mass spectrometry,which indicated that Au and(N(CN)2)–in the SD solution formed a series of metal complex ions,[AuNax(N(CN)2)x+2]–(x=1,2,3,or 4).

Study on Mechanism of Gold Leaching by Thiourea in the Presence of Sodium Sulphite

The linear potential sweep voltammetry and electrochemical methods were used to study the mechanism and kinetics of gold dissolution in thiourea solution in the absence and presence of sodium sulphite.The results show that in the absence of sodium sulfite the dissolution rate of gold in thiourea solution is relatively slow and it increases with the increase of the concentration of thiourea and decreases with pH. Thiourea is easily oxidized to formamidine disulphide, which reduces the leaching efficiency and increases reagent consumption. In the presence of sodium sulphite, the effective concentration of thiourea is increased and the dissolution rate of gold in thiourea solution is greatly increased. The activation energy required for gold dissolution in thiourea solution with sodium sulphite is 20.9 kJ·mol -1 which is much lower than 55.0 kJ·mol -1 in the absence of sodium sulphite. It is evident that sodium sulphite has a significant effect on the electrochemical reaction kinetics of gold in thiourea solution.

Enhanced extraction of scandium and inhibiting of iron from acid leaching solution of red mud by D2EHPA and sodium chloride

D2EHPA(P204),tri-butyl-phosphate(TBP)and sodium chloride(NaCl)were attractive for selective extraction of scandium from acid leaching solution of red mud.The extraction parameters of P204 concentration(X_(P204)),NaCl concentration(C_(NaCl)),pH value,reaction time,stirring speed and O/A were investigated to extract scandium and separate iron from the acid leaching solution.The extraction mechanism was analyzed by Fourier transform infrared spectroscopy(FT-IR)and thermodynamic theory.The single-stage extraction efficiency of scandium,iron andβ(Sc/Fe)were 99.1%,9.4%and 1061.2,respectively,with C_(NaCl) of 75 g/L and XP204 of 0.75 at solution pH value of 1.2 and stirring speed of 200 r/min for 6 min,in which a good separation effect of scandium and iron was obtained.The vibration absorption peak Sc─O was contributed to the extraction of scandium with P204.The complex[FeCln]^(3−n) existed in the solution with adding NaCl into the acid leaching solution.The value of n was higher and the valence state of the complex[FeCln]^(3−n) was lower with an increase of chloride concentration,which restricts the extraction efficiency of iron with P204.The extraction of three stages in the counter-current simulation experiments was carried out according to the McCabe-Thiele diagram.Gibbs free energy change(ΔG)of−5.93 kJ/mol,enthalpy change(ΔH)of 23.45 kJ/mol and entropy change(ΔH)of 98.54 J/(mol·K)were obtained in the solvent extraction proces,which indicate that the extraction reaction is easily spontaneous and endothermic and a proper increase of temperature is conducive to the extraction of scandium.

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time,liquid to solid ratio,leaching temperature,agitation speed,and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum.The optimum process operating parameters were found to be：time,4 h;sodium hydroxide concentration,10%;sodium hypochlorite concentration,8%;liquid to solid ratio,10：1;temperature,50℃;and, agitation speed,500 r/min.Under these conditions the extraction of molybdenum is greater than 99,9%and the extraction of copper is less than 0.01%.A shrinking particle model could be used to describe the leaching process.The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.

Leaching kinetics of molybdenum from Ni-Mo ore in sulfuric acid solution with sodium peroxodisulfate as oxidant

The leaching kinetics of molybdenum from Ni-Mo ore in sulfuric acid solution with sodium peroxodisulfate was studied.The effects including leaching temperature, reaction time, particle size, stirring speed, and concentrations of sulfuric acid and sodium peroxodisulfate were investigated. The leaching process of molybdenum from Ni-Mo ore is controlled by the chemical reaction through the solid layer across the unreacted shrinking core. The apparent activation energy of the leaching of molybdenum is calculated to be 41.0 k J/mol and the leaching kinetics equation of molybdenum from Ni-Mo ore is expressed as1-(1-a)1/3=3405.7exp[-41030.0/(RT)]t.

Leaching of vanadium,sodium,and silicon from molten V?Ti-bearing slag obtained from low-grade vanadium-bearing titanomagnetite

The water leaching process of vanadium, sodium, and silicon from molten vanadium-titanium-bearing（V-Ti-bearing） slag obtained from low-grade vanadium-bearing titanomagnetite was investigated systematically. The results show that calcium titanate, sodium aluminosilicate, sodium oxide, silicon dioxide and sodium vanadate are the major components of the molten V-Ti-bearing slag. The experimental results indicate that the liquid-solid（L/S） mass ratio significantly affects the leaching process because of the respective solubilities and diffusion rates of the components. A total of 83.8% of vanadium, 72.8% of sodium, and 16.1% of silicon can be leached out via a triple counter-current leaching process under the optimal conditions of a particle size below 0.074 mm, a temperature of 90°C, a leaching time of 20 min, an L/S mass ratio of 4：1, and a stirring speed of 300 r/min. The kinetics of vanadium leaching is well described by an internal diffusion-controlled model and the apparent activation energy is 11.1 kJ/mol. The leaching mechanism of vanadium was also analyzed.

Leaching kinetics of low-grade copper ore containing calcium-magnesium carbonate in ammonia-ammonium sulfate solution with persulfate

The leaching kinetics of copper from low-grade copper ore was investigated in ammonia-ammonium sulfate solution with sodium persulfate. The effect parameters of stirring speed, temperature, particle size, concentrations of ammonia, ammonium sulfate and sodium persulfate were determined. The results show that the leaching rate is nearly independent of agitation above 300 r/min and increases with the increase of temperature, concentrations of ammonia, ammonium sulfate and sodium persulfate. The EDS analysis and phase quantitative analysis of the residues indicate that bornite can be dissolved by persulfate oxidization. The leaching kinetics with activation energy of 22.91 kJ/mol was analyzed by using a new shrinking core model （SCM） in which both the interfacial transfer and diffusion across the product layer affect the leaching rate. A semi-empirical rate equation was obtained to describe the leaching process and the empirical reaction orders with respect to the concentrations of ammonia, ammonium sulfate and sodium persulfate are 0.5, 1.2 and 0.5, respectively.

Oxidative leaching behavior of metalliferous black shale in acidic solution using persulfate as oxidant

The oxidative dissolution of metalliferous black shale in sulfuric acid solution using sodium persulfate as an oxidant was investigated. The effects of leaching factors including leaching temperature, leaching time, stirring speed, initial concentration of sodium persulfate and sulfuric acid and particle size on the leaching rate were studied as well. The leaching kinetics of molybdenum, nickel and iron from metalliferous black shale shows that the leaching rate is controlled by a chemical reaction through a layer on the unreacted shrinking core. The leaching process follows the kinetics model 1-（1-a）^1/3=kt with apparent activation energies of 34.50, 43.14 and 71.79 kJ/mol for Mo, Ni and Fe, respectively. The reaction orders in sodium persulfate are 0.80, 1.01 and 0.75 for molybdenum, nickel and iron, respectively, while in sulfuric acid, these orders are 0.45, 0.75 and 0.50 for molybdenum, nickel and iron, respectively. In addition, the reaction mechanism for the dissolution of the metalliferous black shale was discussed.

Molybdenite alkali fusion and leaching:reactions and mechanism

The production of MoO3 from Sarcheshmeh molybdenite concentrate via a pyro-hydrometallurgical process was studied.The molybdenite concentrate and sodium carbonate were premixed and fused under air atmosphere.Then the fused products were leached in water and the dissolved molybdenum was recovered as ammonium molybdate.The ammonium molybdate was then calcined to produce mo-lybdic oxide.At the fusion stage,the effect of the mass ratio of carbonate to sulfide on the reaction products and the solubility of the products was investigated.The results show that during the fusion,sodium molybdate and sodium sulfate are the final reaction products and sodium sulfide is detected as an intermediate reaction product.By melting at 850℃with 5wt%excess carbonate,the maximum solubility of the products is obtained.The molybdenum is recovered from the solutions as ammonium molybdate.

Process for recycle of spent lithium iron phosphate battery via a selective leaching-precipitation method

Applying spent lithium iron phosphate battery as raw material,valuable metals in spent lithium ion battery were effectively recovered through separation of active material,selective leaching,and stepwise chemical precipitation.Using stoichiometric Na2S2O8 as an oxidant and adding low-concentration H2SO4 as a leaching agent was proposed.This route was totally different from the conventional methods of dissolving all of the elements into solution by using excess mineral acid.When experiments were done under optimal conditions(Na2S2O8-to-Li molar ratio 0.45,0.30 mol/L H2SO4,60℃,1.5 h),leaching efficiencies of 97.53% for Li^+,1.39%for Fe^3+,and 2.58% for PO4^3−were recorded.FePO4 was then recovered by a precipitation method from the leachate while maintaining the pH at 2.0.The mother liquor was concentrated and maintained at a temperature of approximately 100℃,and then a saturated sodium carbonate solution was added to precipitate Li2CO3.The lithium recovery yield was close to 80%.

ONE-STEP LEACHING OF SOME REFRACTORY GOLD CONCENTRATECONTAINING ARSENIC AND THEORY ANALYSIS

A new technology of one-step leaching of refractory gold concentrate containing arsenic and sulfur was studied and 96. 8% of gold was extracted in an hour by this technology. Themodynamic possibility of reactions of arsenopyrite. pyrite and gold with sodium hypochlorite in NaOH medium, and solution chemistry of H3AuO3 at different pH values were analyzed. The reason why the gold wrapped by FeAsS, FeS2 can be leached out rapidly by grnding of porcelain ball mill. is explained.

Pressure leaching of bismuth sulfide concentrate containing molybdenum and tungsten in alkaline solution

The leaching results of bismuth sulfide concentrate containing molybdenum and tungsten in air-H2O2-NaOH system, pressure-O2-Na2CO3 system and pressure-O2-NaOH system were investigated. The results show that the extraction of molybdenum, tungsten and sulfur goes up with the increase of NaOH concentration, oxygen partial pressure and reaction time. The extraction of molybdenum and tungsten also rises up with temperature, but the leaching ratio of sulfur increases initially to a peak of 98% at 150℃ and then decreases with the increase of temperature. Under the optimal conditions, the extraction of molybdenum, tungsten and sulfur is more than 95.6%, 93.8% and 96.0%, respectively, and the main phases of residue are Bi2O3 and Fe2O3. Therefore, the method of pressure leaching in alkaline solution is provided as an effective separation of molybdenum, tungsten and sulfur from bismuth and a beneficial pretreatment for consequent process.