Process mineralogy characteristics of acid leaching residue produced in low-temperature roasting-acid leaching pretreatment process of refractory gold concentrates

To provide a theoretical basis for a suitable process to extract gold from refractory gold concentrates, process mineralogy on the acid leaching residue of gold calcine was studied by chemical composition, X-ray diffraction, scanning electron microscopy-energy spectrum, and mineral dissociation analysis. The results showed that the acid leaching residue contained Au 68.22 g/t, Ag 92.71 g/t, Fe 0.44%, As 0.10%, and S 0.55%. Gold and silver minerals existed as native gold, argentite, and proustite. Quartz, the main gangue mineral, accounted for 78.33 wt/%. The dissociation degree analysis showed that the proportions of monomer and exposed gold in acid leaching residue were 96.66 wt%. The cyanidation results showed that the cyanide gold leaching rate of acid leaching residues was close to 100 wt%. However, the maximum cyanide gold leaching rate of gold calcine was only 85.31 wt%. This suggests that acid leaching can increase the gold dissolution rate in the cyanide process.

Manganese extraction from high-iron-content manganese oxide ores by selective reduction roasting-acid leaching process using black charcoal as reductant

Reduction roasting-acid leaching process was utilized to process high-iron-content manganese oxide ore using black charcoal as reductant. The results indicate that, compared with the traditional reductant of anthracite, higher manganese extraction efficiency is achieved at lower roasting temperature and shorter residence time. The effects of roasting parameters on the leaching efficiency of Mn and Fe were studied, and the optimal parameters are determined as follows: roasting temperature is 650 °C, residence time is 40 min, and black charcoal dosage is 10%(mass fraction). Under these conditions, the leaching efficiency of Mn reaches 82.37% while that of Fe is controlled below 7%. XRD results show that a majority of MnO2 and Fe2O3 in the raw ore are reduced to MnO and Fe3O4, respectively.

Calcified roasting-acid leaching process of vanadium from low-grade vanadium-containing stone coal

The low-grade vanadium-containing stone coal used in this experiment was collected from Wuxi Country, Chongqing City, China. The experiment focused on the vanadium recovery from roasted residue through optimizing the process conditions of an effective and environmentally-friendly technology, named calcified roasting-sulfuric acid leaching technology. By single-factor experiments and orthogonal experiments, the effects of roasting temperature, roasting time, sulfuric acid concentration and leaching time on the leaching ratio of vanadium were analyzed. The results showed that the leaching ratio of vanadium reached 85.5% under the proper technological conditions of roasting temperature=950℃, roasting time=4 h, 40% concentration of sulfuric acid and leaching time=6 h.

Leaching of WO_(3)from sulfuric acid converted product of scheelite in NH_(3)·H_(2)O−NH_(4)HCO_(3)solution

The leaching of sulfuric acid converted product of scheelite in NH_(3)·H_(2)O−NH_(4)HCO_(3)solution was systemically studied to improve sulfuric acid conversion−ammonium salts leaching technology route for ammonium paratungstate production.The results showed that the WO_(3)leaching efficiency was about 99%under optimal conditions of 350 r/min,liquid-to-solid ratio of 3 mL/g,1 mol/L NH_(4)HCO_(3),4 mol/L NH_(3)·H_(2)O,25℃,and 15 min.During the leaching,CaSO_(4)almost had no change and was still in a banding or rod-like shape in short leaching time,while conglobate CaCO_(3)was gradually formed on the CaSO_(4)surface.A secondary reaction might occur between CaSO_(4)and WO_(4)^(2−),which could be restrained by a certain amount of CO_(3)^(2−)in the solution.There was no CaCO_(3)phase determined by XRD in leaching residue of converted product for scheelite concentrate under optimal conditions,which was different from that for synthetic scheelite.The leaching process could be explained by neutralization reaction of H_(2)WO_(4)and solid transformation of CaSO_(4)in NH_(3)·H_(2)O−NH_(4)HCO_(3)solution.

Recovery of Li, Ni, Co and Mn from spent lithium-ion batteries assisted by organic acids: Process optimization and leaching mechanism

The proper recycling of spent lithium-ion batteries(LIBs)can promote the recovery and utilization of valuable resources,while also negative environmental effects resulting from the presence of toxic and hazardous substances.In this study,a new environmentally friendly hydro-metallurgical process was proposed for leaching lithium(Li),nickel(Ni),cobalt(Co),and manganese(Mn)from spent LIBs using sulfuric acid with citric acid as a reductant.The effects of the concentration of sulfuric acid,the leaching temperature,the leaching time,the solid-liquid ratio,and the reducing agent dosage on the leaching behavior of the above elements were investigated.Key parameters were optimized using response surface methodology(RSM)to maximize the recovery of metals from spent LIBs.The maxim-um recovery efficiencies of Li,Ni,Co,and Mn can reach 99.08%,98.76%,98.33%,and 97.63%.under the optimized conditions(the sulfuric acid concentration was 1.16 mol/L,the citric acid dosage was 15wt%,the solid-liquid ratio was 40 g/L,and the temperature was 83℃ for 120 min),respectively.It was found that in the collaborative leaching process of sulfuric acid and citric acid,the citric acid initially provided strong reducing CO_(2)^(-),and the transition metal ions in the high state underwent a reduction reaction to produce transition metal ions in the low state.Additionally,citric acid can also act as a proton donor and chelate with lower-priced transition metal ions,thus speeding up the dissolution process.

Selective leaching of lithium from spent lithium-ion batteries using sulfuric acid and oxalic acid

Traditional hydrometallurgical methods for recovering spent lithium-ion batteries(LIBs)involve acid leaching to simultaneously extract all valuable metals into the leachate.These methods usually are followed by a series of separation steps such as precipitation,extraction,and stripping to separate the individual valuable metals.In this study,we present a process for selectively leaching lithium through the synergistic effect of sulfuric and oxalic acids.Under optimal leaching conditions(leaching time of 1.5 h,leaching temperature of 70°C,liquid-solid ratio of 4 mL/g,oxalic acid ratio of 1.3,and sulfuric acid ratio of 1.3),the lithium leaching efficiency reached89.6%,and the leaching efficiencies of Ni,Co,and Mn were 12.8%,6.5%,and 21.7%.X-ray diffraction(XRD)and inductively coupled plasma optical emission spectrometer(ICP-OES)analyses showed that most of the Ni,Co,and Mn in the raw material remained as solid residue oxides and oxalates.This study offers a new approach to enriching the relevant theory for selectively recovering lithium from spent LIBs.

Effect of Na_(2)CO_(3)Roasting Activation on Acid Leaching Property of Fly Ash

By using high-alumina fly ash as raw material,a process was proposed for activating the fly ash with Na_(2)CO_(3)calcination and extracting aluminum from activated clinker with sulfuric acid leaching.The feasibility of roasting process of activated fly ash by Na_(2)CO_(3)was discussed based on thermodynamic analysis.The experimental results showed that Na_(2)CO_(3)gradually reactes with mullite over 700 K to produce NaAlSiO_(4).The optimal process conditions for the activation stage are:a material ratio of 1:1 between sodium carbonate and fly ash,a calcination temperature of 900℃,and a calcination time of 2.5 hours.Under these conditions,the leaching rate of aluminum is 90.3%.By comparing the SEM and XRD analysis of raw and clinker materials,it could be concluded that the mullite phase of fly ash is almost completely destroyed and transformed into sodium aluminosilicate with good acid solubility.

Hydrometallurgical processing of chalcopyrite:A review of leaching techniques

Copper,an essential metal for the energy transition,is primarily obtained from chalcopyrite through hydrometallurgical and pyrometallurgical methods.The risks and harmful impacts of these processes pose significant concerns for environmental and human safety,highlighting the need for more efficient and eco-friendly hydrometallurgical methods.This review article emphasizes current pro-cesses such as oxidative leaching,bioleaching,and pressure leaching that have demonstrated efficiency in overcoming the complicated chalcopyrite network.Oxidative leaching operates under benign conditions within the leaching media;nevertheless,the introduction of oxidizing agents provides benefits and advantages.Bioleaching,a non-aggressive method,has shown a gradual increase in copper extrac-tion efficiency and has been explored using both primary and secondary sources.Pressure leaching,known for its effectiveness and se-lectivity in copper extraction,is becoming commercially more viable with increased research investments.This research also provides im-portant data for advancing future research in the field.

Study on synergistic leaching of potassium and phosphorus from potassium feldspar and solid waste phosphogypsum via coupling reactions

To achieve the resource utilization of solid waste phosphogypsum(PG)and tackle the problem of utilizing potassium feldspar(PF),a coupled synergistic process between PG and PF is proposed in this paper.The study investigates the features of P and F in PG,and explores the decomposition of PF using hydrofluoric acid(HF)in the sulfuric acid system for K leaching and leaching of P and F in PG.The impact factors such as sulfuric acid concentration,reaction temperature,reaction time,material ratio(PG/PF),liquid–solid ratio,PF particle size,and PF calcination temperature on the leaching of P and K is systematically investigated in this paper.The results show that under optimal conditions,the leaching rate of K and P reach more than 93%and 96%,respectively.Kinetics study using shrinking core model(SCM)indicates two significant stages with internal diffusion predominantly controlling the leaching of K.The apparent activation energies of these two stages are 11.92 kJ·mol^(-1)and 11.55 kJ·mol^(-1),respectively.

Effect of Saline Water on Soil Acidity, Alkalinity and Nutrients Leaching in Sandy Loamy Soil in Rwamagana Bella Flower Farm, Rwanda

The necessity to saline and sodic waters is sometimes used for irrigating agricultural activities under certain circumstances, but it is important to note that the use of these waters comes with specific considerations and limitations. One way to decrease undesirable effects of sodic waters on the physical and chemical properties of soils is to apply organic and chemical amendments within the soil. This study aimed to assess the effectiveness of saline water on soil acidity, alkalinity and nutrients leaching in sandy loamy soil at Bella flower farm, in Rwamagana District, Rwanda. The water used was from the Muhazi Lake which is classified as Class I (Saline water quality). Column leaching experiments using treated soils were then conducted under saturated conditions. The soil under experimental was first analyzed for its textural classification, soil properties and is classified as sandy loamy soil. The t-test was taken at 1%, 5% and 10% levels of statistical significance compared to control soil. The results indicated that the application of saline water to soils caused an increase in some soil nutrients like increase of Phosphorus (P), Potassium (K+), Magnesium (Mg2+), Sulphur (S), CN ratio and Sodium (Na+) and decreased soil texture, physical and chemical properties and remained soil nutrients. Consequently, the intensive addition of saline water leachates to soil in PVC pipes led to decreased of soil EC through leaching and a raiser Soluble Sodium Percentage (SSP). The rate of saline water application affected the increase accumulation of SAR and Na% in the top soil layers. The study indicated that saline water is an inefficient amendment for sandy soil with saline water irrigation. The study recommends further studies with similar topic with saline water irrigation, as it accentuated the alkalinity levels.

Experimental investigation of dynamic characteristics of leaching tubing for solution mining of salt cavern carbon and energy storage

Salt caverns are extensively utilized for storing various substances such as fossil energy,hydrogen,compressed air,nuclear waste,and industrial solid waste.In China,when the salt cavern is leached through single-well water solution mining with oil as a cushion,engineering challenges arise with the leaching tubing,leading to issues like damage and instability.These problems significantly hinder the progress of cavern construction and the control of cavern shape.The primary cause of this is the flowinduced vibration instability of leaching tubing within a confined space,which results in severe bending or damage to the tubing.This study presents a model experimental investigation on the dynamic characteristics of leaching tubing using a self-developed liquid-solid coupling physical model experiment apparatus.The experiment utilizes a silicone-rubber pipe(SRP)and a polycarbonate pipe(PCP)to examine the effects of various factors on the dynamic stability of cantilevered pipes conveying fluid.These factors include external space constraint,flexural rigidity,medium outside the pipe,overhanging length,and end conditions.The experiments reveal four dynamic response phenomena:water hammer,static buckling,chaotic motion,and flutter instability.The study further demonstrates that the length of the external space constraint has a direct impact on the flutter critical flow velocity of the cantilevered pipe conveying fluid.Additionally,the flutter critical flow velocity is influenced by the end conditions and different external media.

Extraction of Valuable Metals from Titanium-bearing Blast Furnace Slag by Acid Leaching

To realize the resource utilization of the valuable metals in the titanium-containing blast furnace slag,the process route of “hydrochloric acid leaching-electrolysis-carbonization and carbon dioxide capture-preparation of calcium carbonate” was proposed.In this study,the influences of process conditions on the leaching rates of calcium,magnesium,aluminum,and iron and the phases of the leaching residue were investigated for the leaching process.The experimental results show that the HCl solution could selectively leach the elements from the titanium-containing blast furnace slag.The better leaching conditions are the HCl solution concentration of 4 mol/L,the leaching time of 30 min,the ratio of liquid volume to solid gas of 10 mL/g,and the stirring paddle speed of 300 r/min.Under the conditions,the leaching rates of calcium,magnesium,aluminum,and iron can reach 85.87%,73.41%,81.35%,and 59.08%,and the leaching rate of titanium is 10.71%.The iron and the aluminum are removed from the leachate to obtain iron-aluminum water purification agents,and the magnesium is removed from the leachate to obtain magnesium hydroxide.The leaching residue phase is dominated by perovskite,followed by magnesium silicate and tricalcium aluminate,and the titaniumrich material could be obtained from the leaching residue by desiliconization.

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.

Process engineering of demineralisation of moderate to high ash Indian coals through NaOH‑HCl leaching and HF leaching

Chemical leaching of coals would be required to produce cleaner coals for some special applications where physical benefi-ciation may not be effective enough.This would also help in recovering Li and rare earth metals besides in the sequestration of CO_(2).About 20 Indian coals having complexly distributed moderate to high ash contents were sequentially treated with various alkali–acid such as NaOH-HCl,HF,HCl,HCl-HF,and NaOH-HCl-HF leaching.This aimed to establish and design the best stepwise sequential process for the highest degree of demineralisation through a chemical leaching process.Kinetics and process intensification studies were carried out.More than 80%demineralisation of Madhaipur and Neemcha coals was observed using the best sequential treatment designed presently.The repeated stepwise treatment of the alkali and the acid was also studied,which was found to significantly enhance the degree of demineralisation of coals.The integrated process of alkali–acid leaching followed by solvent extraction(Organo-refining)and vice versa of the treated coal was also studied for producing cleaner coals.

Recovery of vanadium from vanadium slag by composite additive roasting-acid leaching process

To minimize the vanadium content in the vanadium extraction tailings, composite additive roasting with (CaO + MgO) and subsequent acid leaching process was carried out dealing with vanadium-bearing converter slag. The effect of additive with different MgO/(CaO + MgO) molar ratios on the roasting and leaching behaviours of vanadium slag was investigated, and the optimum process conditions were obtained. The results show that in the roasting experiment, under the conditions of roasting temperature of 850 ℃ and roasting time of 2 h, the main kinds of vanadate transformed from Ca_(2)V_(2)O_(7) to Ca_(5)Mg_(4)V_(6)O_(2)4 and then to Mg_(2)V_(2)O_(7) with the increase in the MgO/(CaO + MgO) molar ratio. In the leaching experiment, under the conditions of particle size less than 75 μm, leaching temperature of 50 ℃, pH of 2.5, liquid–solid ratio of 20:1, and MgO/(CaO + MgO) molar ratio of 1:3, the leaching efficiency of vanadium is increased by about 5%, but the substitution of MgO for most or all of CaO will significantly reduce the leaching efficiency of vanadium. Furthermore, the leaching efficiency of impurities (P and Cr) can also be decreased by a composite addictive (CaO + MgO) roasting process. The X-ray diffraction, scanning electron microscopy, energy dispersive spectrometry, and X-ray photoelectron spectroscopy of the original vanadium slag and solid products of both roasting and leaching processes were also evaluated.

Study on metal recovery process and kinetics of oxidative leaching from spent LiFePO_(4)Li-batteries

A green environmental protection and enhanced leaching process was proposed to recover all elements from spent lithium iron phosphate(LiFePO_(4)) lithium batteries.In order to reduce the influence of Al impurity in the recovery process,NaOH was used to remove impurity.After impurity removal,the spent LiFePO_(4) cathode material was used as raw material under the H_(2)SO_(4) system,and the pressure oxidation leaching process was adopted to achieve the preferential leaching of lithium.The E-pH diagram of the Fe-P-Al-H_(2)O system can determine the stable region of each element in the recovery process of spent LiFePO_(4)Li-batteries.Under the optimal conditions(500 r·min^(-1),15 h,363.15 K,0.4 MPa,the liquid-solid ratio was 4:1 ml·g^(-1)and the acid-material ratio was 0.29),the leaching rate of Li was 99.24%,Fe,Al,and Ti were 0.10%,2.07%,and 0.03%,respectively.The Fe and P were precipitated and recovered as FePO_(4)·2H_(2)O.The kinetic analysis shows that the process of high-pressure acid leaching of spent LiFePO_(4) materials depends on the surface chemical reaction.Through the life cycle assessment(LCA)of the spent LiFePO_(4) whole recovery process,eight midpoint impact categories were selected to assess the impact of recovery process.The results can provide basic environmental information on production process for recycling industry.

Environmental availability and ecological risk assessment of heavy metals in zinc leaching residue

Four different methods,namely mineralogical analysis,three-stage BCR sequential extraction procedure,dynamic leaching test and Hakanson Potential Ecological Risk Index Method were used to access the environmental activity and potential ecological risks of heavy metals in zinc leaching residue.The results demonstrate that the environmental activity of heavy metals declines in the following order：CdZnCuAsPb.Potential ecological risk indices for single heavy metal are CdZnCuAsPb.Cd has serious potential ecological risk to the ecological environment and contributes most to the potential toxicity response indices for various heavy metals in the residue.

Leaching behavior of metals from high-arsenic dust by NaOH-Na_2S alkaline leaching

Arsenic is selectively extracted from high-arsenic dust by NaOH-Na2S alkaline leaching process. In the leaching arsenic process, the effects of alkali-to-dust ratio, sodium sulfide addition, leaching temperature, leaching time and liquid-to-solid ratio on metals leaching efficiencies were investigated. The results show that the arsenic can be effectively separated from other metals under the optimum conditions of alkali/dust mass ratio of 0.5, sodium sulfide addition of 0.25 g/g, leaching temperature of 90 ℃, leaching time of 2 h, and liquid-to-solid ratio of 5：1 （mL/g）. Under these conditions, the average leaching efficiencies of arsenic, antimony, lead, tin and zinc are 92.75%, 11.68%, 0.31%, 29.75% and 36.85%, respectively. The NaOH-Na2S alkaline leaching process provides a simple and highly efficient way to remove arsenic from high-arsenic dust, leaving residue as a suitable lead resource.

Technological mineralogy and environmental activity of zinc leaching residue from zinc hydrometallurgical process

Chemical, physical, structural and morphological properties of zinc leaching residue were examined by the combination of various detection means such as AAS, XRF, XRD, M？ssbauer spectrometry, SEM-EDS, TG-DSC, XPS and FTIR. The toxicity characteristic leaching procedure （TCLP） was used to investigate the environmental activity of zinc leaching residue for a short contact time. The phase composition analysis indicated that the zinc leaching residue mainly consists of super refined flocculent particles including zinc ferrite, sulfate and silicate. The physical structural analysis showed that it has a thermal instability and strong water absorption properties. The results of TCLP indicated that the amounts of Zn and Cd in the leaching solution exceed 40 and 90 times of limit, respectively, which demonstrate that this residue is unstable in weak acidic environment for a short contact time.

Leaching of heavy metals from Dexing copper mine tailings pond

AThe wastewater source of 4# tailing pond in Dexing copper mine consists of alkaline flotation pulp and acid mine drainage （AMD） from the nearby opencast mine. Therefore, the heavy metals in tailing ore are very likely to be released due to acidification from AMD. The leaching behaviors ofZn, Cu, Fe and Mn in mine tailings from Dexing copper mine were investigated by a series of laboratory batch experiments. The effectcs ofpH, temperature, particle size and contact time on the leachability of such heavy metals were examined. It was evident that Zn, Cu, Fe and Mn were major heavy metals in the tailings while gangue minerals like quartz were major constituents in examined tailings. The tailing dissolution reaction was controlled by the acid, whose kinetics could be expressed according to the heterogeneous reaction models and explained by a shrinking core model with the surface chemical reaction as the rate-controlling step. The leachability of all metals examined depended on pH and contact time. The batch studies indicated that the maximum leaching ratios ofZn, Cu, Fe and Mn at pH 2.0 were 5.4%, 5.8%, l 1.1% and 34.1%, respectively. The dissolubility of all metals examined was positively correlated to the temperatures. The particle size would not change dissolution tendency of those heavy metals, but decrease the concentrations of leached heavy metals.