Recovery of zinc and cadmium from industrial waste by leaching/cementation

Metallic zinc production from sulfide zinc ore is comprised by the stages of ore concentration, roasting, leaching, liquor purification, electrolysis and melting. During the leaching stage with sulfuric acid, other metals present in the ore in addition to zinc are also leached. The sulfuric liquor obtained in the leaching step is purified through impurities cementation. This step produces a residue with a high content of zinc, cadmium and copper, in addition to lead, cobalt and nickel. This paper describes the study of selective dissolution of zinc and cadmium present in the residue, followed by the segregation of those metals by cementation. The actual sulfuric solution, depleted from the electrolysis stage of metallic zinc production, was used as leaching agent. Once the leaching process variables were optimized, a liquor containing 141 g/L Zn, 53 g/L Cd, 0.002 g/L Cu, 0.01 g/L Co and 0.003 g/L Ni was obtained from a residue containing 30 wt.% Zn, 26 wt.% Cd, 7 wt.% Cu, 0.35 wt.% Co and 0.32 wt.% Ni. The residue mass reduction exceeded 80 wt.%. Cementation studies investigated the influence of temperature, reaction time, zinc concentration in feeding solution, pH of feeding solution and metallic zinc excess. After that such variables were optimized, more than 99.9% of cadmium present in liquor was recovered in the form of metallic cadmium with 97 wt.% purity. A filtrate (ZnSO₄ solution) containing 150 g/L Zn and 0.005 g/L Cd capable of feeding the electrolysis zinc stage was also obtained.     

Development of a process for the separation of zinc and copper from sulfuric liquor obtained from the leaching of an industrial residue by solvent extraction

This paper describes a study of the separation of zinc and copper from the leach liquor generated in the treatment of the zinc residue (29.6 g/L Zn and 37.4 g/L Cu) by liquid–liquid extraction. In it, the influence of the extractant type and concentration, aqueous phase acidity, contact time and stripping agent concentration were investigated. Organophosphorus extractants (D2EHPA, IONQUEST®801 and CYANEX®272) and the chelating extractants (LIX®63, LIX®984N and LIX®612N-LV) were also investigated. The organophosphorus reagents are selective for zinc, while the chelating extractants are selective for copper. In the experiment, D2EHPA was found to be the best extractant. A sulfuric acid solution was used in the stripping study. Five continuous experiments were carried out until an optimal condition for the separation of the metals Zn and Cu was achieved. Experiment 5 was carried out in three extraction steps, three scrubbing stages and five stripping stages. In this experiment, a pregnant strip solution containing 125 g/L Zn and 0.01 g/L Cu was obtained and the concentration of the metals in the raffinate was 28.3 g/L Cu and 0.49 g/L Zn.     

Formation of sulphur oxyanions and their influence on antimony electrowinning from sulphide electrolytes

Antimony electrowinning from synthetic alkaline sulphide electrolytes has been studied in a nondiaphragm electrolytic cell. The electrodes were constructed in such a way that the anode produces ten times higher current density than the cathodic current density to promote sulphide oxidation to sulphate at the anode; and simultaneously decreasing the tendency of hydrogen evolution at the cathode. The result revealed that at an anodic current density lower than 1500 A/m², minute amounts of sulphate ions were formed but when the anode current density increased beyond 1500 A/m², sulphate formation was promoted. The initial molar concentration ratio between hydroxide and free sulphide ions should be ⩾10.3 to avoid thiosulphate formation at 2000 A/m² anodic current density under the conditions used in these experiments. The highest anodic current efficiency obtained based on the amount of sulphate formed was 89%. An increase in the anode current density as well as NaOH concentration enhances the cathodic and anodic current efficiencies with respect to the antimony metal deposited and sulphate ions produced, respectively. Despite the high anodic current densities used, the specific energy of this process ranges from 0.6 to 2.3 kW h/kg which is significantly lower than values reported previously due to the prevention of undesirable sulphur species from being formed. The tests revealed that the concentration of thiosulphate formed during the electrolysis decreased with increasing anode current density and NaOH concentration. Addition of polysulphide from 0 to 30 g/L to the electrolyte decreases the current efficiency from 83% to 32% and correspondingly increases the specific energy from 1.7 to 4.8 kW h/kg. Results showed that a build-up of sulphite and sulphate ions in the solution does not have any detrimental effect on the current efficiency of antimony deposition.     

Uranium recovery from strong acidic solutions by solvent extraction with Cyanex 923 and a modifier

Uranium stripping with strong acid solution is always highly desired due to its simple operation and less pollution. However, intensive acid neutralisation for uranium precipitation in the subsequent step limited its application. A new solvent extraction process has been developed to transfer uranium from strong to weak sulphuric acid solutions suitable for uranium precipitation without intensive neutralisation. An organic system consisting of 10% Cyanex 923 and 10% isodecanol as the modifier in ShellSol D70 was optimised for the process. It was found that uranium was extracted efficiently from 4 to 6 M H₂SO₄ solutions with the organic system, and it could be efficiently stripped with 0.2–0.5 M H₂SO₄ solutions. Both extraction and stripping kinetics of uranium were very fast, reaching the equilibrium within 0.5 min. Temperature between 30 and 60 °C has slight effect on uranium extraction and stripping. Four theoretical stages could effectively extract more than 98% uranium from a solution containing 17.5 g/L U and 6.0 M H₂SO₄ at an A/O ratio of 1:1.5, and it could generate a loaded organic solution containing about 12 g/L U. More than 99% U could be stripped from the loaded organic solution containing 14.6 g/L U with 0.5 M H₂SO₄ using five stages at an A/O ratio of 1:3. As a result, the loaded strip liquor containing more than 40 g/L U would be obtained which is suitable for uranium recovery by precipitation using hydrogen peroxide. A conceptual process has been proposed for uranium transfer from strong to weak sulphuric acid solutions for its recovery.     

Process development for the recovery of europium and yttrium from computer monitor screens

This work describes the development of a process for the recovery of Eu and Y from cathode ray tubes (CRTs) of discarded computer monitors with the proposition of a flow sheet for the metals dissolution. Amongst other elements, europium and yttrium are presented in the CRTs in quantities – 0.73 w/w% of Eu and 13.4 w/w% of Y – that make their recovery worthwhile. The process developed is comprised of the sample acid digestion with concentrated sulphuric acid followed by water dynamic leaching at room temperature. In the CRTs, yttrium is present as oxysulphide (Y₂O₂S) and europium is an associated element – Y₂O₂S:Eu³⁺ (red phosphor compound). During the sulphuric acid digestion, oxysulphide is converted into a trivalent Eu and Y sulphate, in solid form, with the liberation of H₂S. In the second step, metals are leached from the solid produced in the acid digestion step by dynamic leaching with water. This study indicates that a proportion of 1250 g of acid per kg of the sample is enough to convert Eu and Y oxysulphide into sulphate. After 15 min of acid digestion and 1.0 h of water leaching, a pregnant sulphuric liquor containing 17 g L⁻¹ Y and 0.71 g L⁻¹ Eu was obtained indicating yield recovery of Eu and Y of 96% and 98%, respectively. Both steps (acid digestion and water leaching) may be performed at room temperature.     

Thorium and uranium extraction from rare earth elements in monazite sulfuric acid liquor through solvent extraction

This paper describes the process of extraction of thorium and uranium from the sulfuric liquor generated in the chemical monazite treatment through a solvent extraction technique. The influence of the extractant type and concentration, contact time between phases, type and concentration of the stripping solution and aqueous/organic volumetric ratio were investigated. The results indicated the possibility of extracting, simultaneously, thorium and uranium from a solvent containing a mixture of Primene JM-T and Alamine 336. The stripping was carried out with a hydrochloric acid solution. After selecting the best conditions for the process, a continuous experiment was carried out in a mixer-settler circuit using four stages in the extraction step, five stages of stripping and one stage of the solvent regeneration. A loaded stripping solution containing 29.3 g/L of ThO₂ and 1.27 g/L of U₃O₈ was obtained. The metals content in the raffinate was below 0.001 g/L, indicating a thorium extraction of over 99.9% and a uranium extraction of 99.4%. The rare earths content in the raffinate was 38 g/L of RE₂O₃.     

Zinc removal in anaerobic sulphate-reducing liquid substrate process

Zinc and sulphate removal from synthetic wastewater was investigated by using four laboratory parallel upflow-mode reactors (referred as R1 to R4; R1 contained carriers to retain biomass, whereas R2–R4 were operated as suspended reactors). All reactors were inoculated with anaerobically digested cow manure. R1 and R2 were first fed with glucose- and sulphate-containing feed for 48 days after which all four reactors were fed with wastewater containing 50 mg l⁻¹ of zinc in R1–R3 and 200 mg l⁻¹ in R4 and operated for 96 days. In all reactors, hydraulic retention time, organic loading rate, and sulphate load were 5–6 d, 0.2–0.4 kg COD m⁻³ d⁻¹ and 3.3–3.8 g SO₄ l⁻¹ d⁻¹, respectively, whereas the zinc load in R1–R3 was 0.074–0.077 and in R4 0.282 g Zn l⁻¹ d⁻¹. During the runs, 30–40% of sulphate and over 98% of zinc was removed, and up to 150–200 mg H₂S was produced in all reactors. Effluent pH dropped in all reactors (feed pH 6.5) to 3–5 by the end of the experiment. No significant effects on zinc removal were observed, despite differences in operating conditions and feed. It was only in the latter part of the runs (i.e. between experiment days 120–142) that zinc removal began to fluctuate, showing a negligible decrease in R3 and R4, whereas in R1 and R2 zinc was removed below the limit of detection (<0.01 mg Zn l⁻¹). Qualitative X-ray diffraction analysis of the reactor sludge at the end of the runs indicated that the compounds precipitated were most probably ZnS (Code 05-0566 Sphalerite), suggesting metal removal through sulphide precipitation; this was supported by the fact that sulphate was reduced and zinc removed simultaneously.     

Alkaline leaching of zinc from electric arc furnace steel dust

Electric arc furnaces (EAF) generate about 10–20 kg of dust per metric ton of steel, which constitute a hazardous waste, known as EAF dust. This dust contains a remarkable amount of non-ferrous metals, which include zinc, cadmium, lead, chromium and nickel that could be recovered, reducing the environmental impact of the leachable toxic metals, and generating revenue. In this paper, different alkaline leaching techniques were tested in order to dissolve the zinc present in an EAF dust: (i) conventional agitation leaching; (ii) pressure leaching; (iii) conventional leaching following a microwave pretreatment and (iv) leaching with agitation provided by an ultra-sonic probe. Temperature and sodium hydroxide concentration were the variables tested. The highest zinc recovery from the EAF dust, containing about 12% of zinc, was about 74%. This was achieved after 4 h of leaching in a temperature of 90 °C and with a sodium hydroxide concentration of 6 M of the leaching agent.     

Recovery of zinc from an industrial zinc leach residue by solvent extraction using D2EHPA

A hydrometallurgical treatment involving solvent extraction of zinc using di-2-ethylhexyl phosphoric acid (D2EHPA) has been investigated to recover zinc from an industrial leach residue. The residue was leached with sulfuric acid producing leach liquor which was subjected to solvent extraction for enrichment of zinc and removal of impurities. Operating variables, such as pH, D2EHPA concentration, temperature, aqueous/organic (A/O) phase ratio, tri-butyl phosphate (TBP) concentration and sodium sulfate (Na₂SO₄) concentration in aqueous phase were studied. Practically, all zinc was extracted from the aqueous solution at pH 2.5 with 20% w/w D2EHPA in kerosene. Increasing either TBP concentration up to 5%, or Na₂SO₄ concentration up to 0.2 M, increased the zinc extraction. Zinc could be extracted at one theoretical stage at A/O of 1/1, as calculated by McCabe–Thiele method.     

Process flowsheet development for recovering antimony from Sb-bearing copper concentrates

The technical feasibility, on laboratory scale, of hydro- and electrometallurgical processes of recovering metallic antimony from an antimony-bearing copper sulphide concentrate has been investigated. The influence of Na₂S concentration, temperature and solid concentration was studied during the leaching test while the effect of current density, Na₂S concentration, electrolyte temperature and NaOH concentration on antimony electrowinning from alkaline sulphide solutions was investigated. The leaching results showed that antimony dissolution is strongly dependent on the concentration of the leaching reagent as well as the leaching temperature. The antimony content in the concentrate was reduced from 1.7% to less than 0.1% Sb which is desirable for copper metallurgy. Cathode current efficiency is one of the important parameters to evaluate the performance of an electrolytic process. It is revealed in this study that current efficiency of antimony deposition from sulphide electrolytes is highly dependent on the concentration of sodium hydroxide and the current density used. The results illustrate that the combined effect of increasing anode current density (which was 10 times higher than the cathode current density) and NaOH concentration enhanced the current efficiency of the electrolytic process. It was demonstrated that excess free sulphide ions impacts the current efficiency of the process detrimentally. An integrated hydro-/electrometallurgical process flowsheet for antimony removal and recovery from a sulphide copper concentrate was developed.     

Metals recovery from galvanic sludge by sulfate roasting and thiosulfate leaching

The purpose of this work is the selective recovery of Au, Ag, Cu, and Zn from two types of galvanic sludge using a mixed process of sulfate roasting and sodium thiosulfate leaching. In the experiments, the sludge was mixed with a sulfate promoter (sulfur, iron sulfate, or pyrite) and treated by pyrometallurgical processes at temperatures up to 750 °C. At this stage, this agent is thermally oxidized, turning the furnace atmosphere into a reducing one and the metallic oxides into water-soluble sulfates. Afterward, the sulfates can be treated by leaching with water for recovery of Ag, Cu, and Zn. The gold does not form sulfates in this reaction and was recovered through a second leaching stage using sodium thiosulfate, an effective reagent and less harmful to the environment than cyanide. Different parameters such as the sulfate promoter that achieves the highest recovery of metals, the proportion of galvanic sludge to sulfating agent, the temperature, the heating time in the oven, and the leaching time were evaluated. Additionally, a comparison of gold recovery using cyanide versus sodium thiosulfate was performed. The configuration that showed the best metal recovery included a 1:0.4 ratio of sludge to sulfur, an oven temperature of 550 °C, a roasting time of 90 min, and a water leaching time of 15 min. Using these parameters, recovery rates of 80% of the silver, 63% of the copper, and 73% of the Zn were obtained. The sodium thiosulfate leaching resulted in a recovery of 77% of the Au, close to the values obtained using cyanide.     

Solvent extraction separation and recovery of cobalt and nickel from sulphate medium using mixtures of TOPS 99 and TIBPS extractants

We report in this paper the solvent extraction separation of cobalt and nickel from synthetic sulphate solutions using TOPS 99 and TIBPS mixtures diluted in kerosene. The feed contains 1.061 g/L Co and 1.187 g/L Ni. Extraction experiments with synergistic mixture of extractants showed highest separation factor of 12,245 with 0.1 M TOPS 99 and 0.05 M TIBPS at pH 1.1. McCabe–Thiele plot for Co extraction with 0.1 M TOPS 99 and 0.05 M TIBPS extractants mixture indicated the necessity of three theoretical stages for >99% Co extraction at an aqueous to organic phase (A/O) ratio of 2. A three stage counter current extraction simulation test conducted at pH 1.1 with 0.1 M TOPS 99 and 0.05 M TIBPS mixture, confirmed Co extraction of 99.5% with Ni co-extraction of 0.02%. The results demonstrated that the addition of TIBPS–TOPS 99 acts as a synergist for Co extraction and antagonist for Ni.     

Selective pressure leaching of Fe (II)-rich limonitic laterite ores from Indonesia using nitric acid

The selective extraction of nickel and cobalt over iron from an Indonesian limonitic laterite was investigated using nitric acid pressure leaching (NAPL). The mineralogical analysis showed that the major minerals were goethite and magnetite, and the content of the divalent iron was as high as 7.06%. Nickel and cobalt were mainly distributed in these two minerals; however, the distribution was non-uniform. A series experiments were conducted to examine the basic parameters and propose the optimal conditions for the extraction. When the ore was treated via HPAL under the optimal condition, the extracted nickel and cobalt were less than 75%, and the iron concentration in the leach liquor was over 12.5 g/L. By contrast, over 85% of nickel and cobalt were extracted and about 1.8 g/L iron was achieved using NAPL. The loss of nickel and cobalt can be mainly attributed to the undissolved magnetite and manganese minerals. The leaching process of NAPL is a dissolution–oxidation–precipitation mechanism, and in this process nitric acid acts as both a lixiviant and an oxidant. The formation of hematite results in a low iron concentration in the leach liquor without oxygen injected. Meanwhile, the oxidation and the precipitation of dissolved divalent iron results in a calculated savings in acid consumption of about 120 kg nitric acid per ton of ore can be obtained, which is equal to over 93 kg of sulfuric acid per ton of ore. Moreover, lower residual acid (20 g/L nitric acid) is a significant advantage of NAPL. The iron residues had a high iron content (>56 wt%) with no sulfur, making it suitable as raw materials for ironmaking.     

Ferric sulphate/chloride leaching of zinc and minor elements from a sphalerite concentrate

Atmospheric leaching of a sphalerite concentrate in sulphate and chloride media was performed and the effect of several variables, such as solid/liquid ratio and oxidant (Fe(III)) concentration were investigated. The behaviour of minor elements, such as Cu, In, As, Sb, Bi, Sn and Pb, was also studied under different conditions. The results showed that using a solid/liquid ratio of 5% (w/v) it was possible to leach 95% of zinc after 2 h, with a solution of 0.5 M H₂SO₄ and Fe₂(SO₄)₃ at 80 °C. The minor elements As, Sb and Bi were also completely leached whereas copper leaching was favoured by the use of chloride medium. The oxidation of Fe(II) during the leaching tests was studied and an improvement of 20% zinc extraction was observed in an oxygenated system. Cross-current leach tests using two/three stages and a solid/liquid ratio of 10% (w/v) were performed to achieve 90% of zinc extraction. The electron microprobe analysis of the leaching residues showed no change on the sphalerite composition after the leaching, which indicates that the leaching of sphalerite involves the break down of the sulphide structure.     

Characterizing hydrocyclone performance for grit removal from wastewater treatment activated sludge plants

Typically, 15–45% of the mixed liquor (sludge) in biological wastewater treatment plants (WWTPs) consists of inorganic (fixed) suspended solids. A portion of these inorganic compounds is grit (sand) originating from the influent. Grit accumulation impacts WWTP design and operating costs as these unbiodegradable solids reduce the effective treatment capacity of the bioreactor and other unit operations that must be sized to carry this material.The goal of this study was to characterize the performance of a hydrocyclone to selectively separate grit from activated sludge. Laboratory experiments were conducted with a 13 mm diameter Krebs hydrocyclone treating sludge from eight WWTPs. Reduced efficiencies of 17 ± 7% on fixed suspended solids and 9 ± 6% on volatile suspended solids were obtained. Grade efficiency curves enabled the development of a modified definition for cut size useful for this application. The characterization of hydrocyclone performance for grit removal from activated sludge will enable modelling of the process for integration into wastewater treatment simulators used for process performance prediction and design.     

A novel method to recover zinc and iron from zinc leaching residue

A novel method to recover zinc and iron from zinc leaching residue (ZLR) by the combination of reduction roasting, acid leaching and magnetic separation was proposed. Zinc ferrite in the ZLR was selectively transformed to ZnO and Fe₃O₄ under CO, CO₂ and Ar atmosphere. Subsequently, acid leaching was carried out to dissolve zinc from reduced ZLR while iron was left in the residue and recovered by magnetic separation. The mineralogical changes of ZLR during the processes were characterized by XRF, TG, XRD, SEM–EDS and VSM. The effects of roasting and leaching conditions were investigated with the optimum conditions obtained as follows: roasted at 750 °C for 90 min with 8% CO and CO/CO + CO₂ ratio at 30%; leached at 35 °C for 60 min with 90 g/l sulfuric acid and liquid to solid ratio at 10:1. The iron was recovered by magnetic separation with magnetic intensity at 1160 G for 20 min. Under the optimum operation, 61.38% of zinc was recovered and 80.9% of iron recovery was achieved. This novel method not only realized the simultaneous recovery of zinc and iron but also solved the environmental problem caused by the storage of massive ZLR.     

Simulated small-scale pilot plant heap leaching of low-grade oxide zinc ore with integrated selective extraction of zinc

The leaching of low-grade oxide zinc ore and simultaneous integrated selective extraction of zinc were investigated using a small-scale leaching column and laboratory scale box mixer-settlers. Di-2-ethylhexyl phosphoric acid (D2EHPA) dissolved in kerosene was used as an extractant. The results showed that it was possible to selectively leach zinc from the ores by heap leaching. The zinc concentration of the leach liquor in the first leaching–extraction circuit was 32.57 g/L, and in the 16th cycle the zinc concentration was 8.27 g/L after the solvent extraction. The leach liquor was subjected to solvent extraction, scrubbing and selective stripping for the enrichment of zinc and the removal of impurities. The pregnant zinc sulfate solution produced from the stripping cycle was suitable for zinc electrowinning.     

Sulphate and molybdate ions uptake by chitin-based shrimp shells

The removal of sulphate and molybdate anions (among other anions) from mining liquid effluents is attracting much interest because of the strict environmental legislation world-wide and the need for water recycling and reuse. In this work, adsorption of sulphate and molybdate ions on chitin-based materials was investigated. Chitin flakes with various deacetylation degrees (DD) were produced from an industrial shrimp shell waste after demineralisation, deproteinisation and deacetylation steps, without further purification, immobilisation or grinding. Batch adsorption experiments were carried out as a function of pH and the best adsorbent material was selected following its chemical stability in acidic medium, degree of anions uptake and time needed for the deacetylation reaction stage. Thus, detailed sulphate adsorption studies were conducted with a chitin having a 25% DD, at various adsorbent concentrations, medium pH and other operating conditions. Best sulphate removal values (92%) were obtained at equilibrium pH 4.5, 8.5 mg mg⁻¹ chitin/ions ratio and 15 min contact time. The adsorption data followed the Langmuir model and showed saturation values of the order of 3.2 mEq g⁻¹. Chitin also proved to adsorb molybdate ions in the presence of sulphate ions, but reaction required longer equilibration time (60 min for the same 92% removal). Practical examples of removal of these anions were studied in actual mining effluents, attaining values of the order of 71% sulphate and 85% Mo from a Cu–Mo flotation mill effluent and 80% sulphate removal from a coal AMD––acid mines drainage (Mo free). The regeneration of the adsorbent material was possible through the anions desorption in alkaline medium. All results are discussed in terms of solution and interfacial phenomena and the practical aspects of the process, in the mining and metallurgy fields, are envisaged.     

An innovative process for extracting boron and simultaneous recovering metallic iron from ludwigite ore

Ludwigite ore has not yet been utilized on an industrial scale due to its complex mineralogy and fine mineral dissemination in China. Boron–iron separation and dissolution activity of boron-bearing minerals in alkaline liquor are the two key issues in the utilization of ludwigite ore, governing the boron recovery as well as operating cost. This paper proposes an innovative process for extraction of boron and iron from ludwigite ore based on coal-based direct reduction process with sodium carbonate (Na₂CO₃). The novel process involves reduction roasting, combined leaching and grinding of reduced ludwigite ore, followed by magnetic separation of leach residue, and experimental validation for each of the processing steps is demonstrated. Alkali-activation of boron and metallization of iron were synchronously achieved during carbothermic reduction of ludwigite ore in the presence of Na₂CO₃. Consequently, boron was readily extracted in the form of sodium metaborate (NaBO₂) with water at room temperature during ball mill grinding, and metallic iron powder was recovered from the leaching-filtering residue by magnetic separation. Boron extraction of 72.1% and iron recovery of 95.7% with corresponding iron grade of 95.7% in the magnetic concentrate were achieved when ludwigite ore was reduced with 20% sodium carbonate at 1050 °C for 60 min.     

Leaching and CIL processes in gold recovery from refractory ore with thiourea solutions

In this study, the applicability of leaching and CIL processes in gold recovery with thiourea method, alternative to the cyanidation from the refractory Gümüşhane-Kaletaş/Eastern Black Sea Region (Turkey) ore was investigated.The experiments were conducted at laboratory conditions using ore samples of which approximately 80% were ground to −0.038 mm. The grade of the ore samples was 6.8 g Au/ton. At the first part of the experimental studies, assuming that the gold could be recovered with CIC and CIP processes, the effects of pH, thiourea, oxidizing agent consumption, and leaching time on leaching were investigated. Then, on the basis of the optimum pH and reagent consumption values obtained in the first part (pH = 1.5, 15.2 kg thiourea/ton ore, 140.9 kg iron(III) sulfate/ton ore and 46.2 kg sulfuric acid consumption/ton ore) and adding 50 kg activated carbon/ton ore at the beginning of experiments, the gold leaching extents were obtained for the same leaching times. At this part, the applicability of CIL process in gold recovery with thiourea was investigated for the first time. As a result of the experiments, although higher gold leaching extents were obtained in CIL process, the increase in extent was about maximum 8% and the highest gold leaching extent was obtained as 75% at the end of the 5th hour.