Use of quaternary ammonium salts to remove copper–cyanide complexes by solvent extraction

Cyanidation is one of the most common methods for the extraction of precious metals. In this process, effluents frequently contain relatively high concentrations of copper, which may react with cyanide to form cuprocyanide complexes adversely affecting the process.In this preliminary work, the use of solvent extraction to remove the copper–cyanide species from a synthetic solution similar to that of gold mill effluents was studied in order to permit the recycling of the solution into the process. For the extraction of these anions, the quaternary ammonium salts Quartamin TPR, Adogen 464 and Aliquat 336 were studied as extractants. The experimental results showed that for a synthetic solution of 710 mg/L copper and 1100 mg/L cyanide, it is possible to obtain a copper extraction of 99% when using 0.033 mol/L of the extractant Adogen 464 (organic/aqueous volume ratio (O/A) = 1) in the range of pH of 9–11. Up to 99% of the copper can be stripped from the organic solution after three contact times (5 min each) with 50 mL of sodium hydroxide 0.5 M (O/A = 1).     

Copper removal from cyanide solutions by acidification

Cyanidation is the most used method to recover gold and silver from their ores. In this process, other metals besides gold and silver may dissolve under certain circumstances and interfere with the efficiency of extraction. Copper is one of these metals, being able to reach concentrations as high as 1000 mg/L. When the cyanide solution contains a high copper concentration, the extraction of precious metals decreases and the operating costs increase. In addition, the control of the process is more complicated because the free cyanide analysis performed by the operators by titration does not represent the actual cyanide available to dissolve gold and silver. This experimental work has two objectives: to evaluate the amount of copper–cyanide complexes that are measured when titration is used for free cyanide analysis, and to develop and propose a method for copper removal from cyanidation solutions. The method consists in the acidification of the solution with sulfuric acid, and the separation of the precipitated solid (CuCN) by filtration. The thermodynamics of the copper–cyanide system is discussed, and the hydrogen cyanide evolution (HCN) at pH acid is evaluated.One of the main operating variables in cyanidation plants is the “free cyanide” concentration, that is, the amount of cyanide available to dissolve the precious metals. The usual method to estimate free cyanide is titration. When copper is present in the cyanide solution, the titration method not only measures the free cyanide but also part of the cyanide that is forming complexes with copper. It was demonstrated from the tests performed in this work, that for cyanide/copper ratios of 1–10, typically found in cyanidation solutions, around 10% of the free cyanide measured by titration corresponds to copper–cyanide complexes and is not available for gold and silver dissolution.In order to recycle the solution to the process, it is necessary to remove part of the copper. A method of copper removal is proposed, based on the precipitation of CuCN when the copper–cyanide solution is acidified. The precipitated solid is separated from the solution by filtration and finally the clear solution is neutralized. The acidification/filtration/alkalinization of cyanidation solutions containing copper permits the removal of most of the copper present in the solution, thus allowing the recycling of the solution. For synthetic solutions containing 200–730 mg/L Cu at different cyanide/copper ratios, it was found that 93–98% of the copper can be removed as CuCN at pH 2.5, releasing free cyanide to the solution. If a flotation stage is considered to remove the solid formed, the HCN formed by acidification should not represent a problem: the amount of HCN gas stripped at pH 2.5 when using the usual flotation gas flow rates (0.17 cm/s, 0.28 L/min, 1 h for our experimental design) was only 6%, which can be easily controlled with conventional equipment.     

Separation of nickel from copper in ammoniacal/ammonium carbonate solution using ACORGA M5640 by selective stripping

Separation of nickel from copper in ammoniacal/ammonium carbonate solution using ACORGA M5640 by selective stripping was carried out. The influence of equilibration time, equilibrium pH and extractant concentration on the extraction of both the metals was studied. It was found that the copper extraction equilibrium was reached in a shorter time than the nickel extraction equilibrium. Nickel extraction decreases above an equilibrium pH of 9.0, while the extraction of copper remains unaffected by the changes in the equilibrium pH range of 7–10. Co-extraction, ammonia scrubbing and the selective stripping of copper and nickel were performed for a solution containing 3 g/l each of copper and nickel and 60 g/l ammonium carbonate. The extraction and the percentage stripping of copper and nickel were almost quantitative.     

Complex treatment of mining and metallurgical wastes for recovery of base metals

A complex process for the recovery of copper and zinc from mining and metallurgical wastes has been investigated and proposed. It includes sulfuric acid leaching of old pyrite flotation tailings to produce ferric containing leach solution; followed by ferric leaching of copper converter slag flotation tailings with the leach solution. A sample of old pyrite flotation tailings from the concentrator containing 0.36% of copper and 0.23% of zinc was leached with 10% sulfuric acid in the column. Recovery of copper and zinc reached 47.1% and 47.2%, respectively. The pregnant leach solutions contained 15.9 g/L of ferric iron. The subsequent ferric leaching of copper converter slag flotation tailings containing 0.53% copper and 2.77% zinc with the pregnant leach solution was conducted. The effects of various process parameters on the leaching dynamics of metals under batch conditions were investigated. Under the best conditions (temperature 70 °C, pulp density 30%, ferric iron concentration 15.9 g/L, initial pH of the pulp 0) the recovery of copper and zinc reached 79.6% and 43.7%, respectively. It was concluded that acid leaching of base metals from old pyrite flotation tailings with pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective and promising technique for the complex treatment of mining and metallurgical wastes.     

Chloride assisted leaching of chalcocite by oxygenated sulphuric acid via Cu(II)–OH–Cl

The beneficial effect of the addition of sodium chloride upon the leaching kinetics of complex iron–nickel–copper sulphides at elevated temperatures and oxygen pressures has been widely reported since the late 1970s, but the role of chloride is still being investigated or debated. Previous researchers have considered chloride as: (i) a complexing agent for cuprous ions; (ii) a surfactant that disperses the molten sulphur and thus removes passivation of the mineral surface by elemental sulphur during pressure leaching; and (iii) a reagent which increases the surface area and the porosity of the insoluble product layer on the surface. A proper understanding of the role of chloride based on the leaching of individual sulphides of known composition in the absence of host minerals at low pulp densities would be useful for the development of chloride assisted sulphate leaching processes for complex sulphide ores, concentrates, and mattes. In the present study evidence for the formation of basic salts of Cu(II) and Fe(III) during leaching are presented. The published rate data are analysed for the leaching of copper from mono-sized chalcocite particles in oxygenated sulphuric acid solutions maintained at 85 °C, a temperature lower than the melting point of sulphur. The initial leaching follows a shrinking particle (sphere) model, and the apparent rate constants are first order with respect to the concentration of dissolved oxygen and chloride. The intrinsic rate constant for the surface reaction (0.2 m s⁻¹) is two orders of magnitude larger than the calculated mass transfer coefficient of oxygen (3 × 10⁻³ m s⁻¹). The proposed reaction mechanism considers the formation of an interim Cu(II)(OH)Cl⁰ species which facilitates the leaching process.     

季铵盐捕收剂对铝硅矿物的浮选行为

通过单矿物浮选试验、动电位测定及红外光谱分析研究了十二烷基三甲基氯化铵、十六烷基三甲基溴化铵和十八烷基二甲基苄基氯化铵3种季铵盐捕收剂对铝硅矿物一水硬铝石、高岭石、叶蜡石和伊利石的浮选行为和作用机理.结果表明:在碱性条件下,以季铵盐为捕收剂可实现一水硬铝石与3种硅酸盐矿物的反浮选分离;一水硬铝石、高岭石、叶蜡石及伊利石的等电点分别为pH6.0、3.4、2.3、3.2,随着矿浆pH值提高,这些矿物的表面动电位均呈负增加;季铵盐捕收剂主要靠静电作用吸附在一水硬铝石、高岭石、叶蜡石及伊利石表面.     

Mechanisms of sulfide ion oxidation during cyanidation. Part II: Surface catalysis by pyrite

The mechanisms and the reaction products for the oxidation of sulfide ions in the presence of pyrite have been established. When the leach solution contains free sulfide ions, oxidation occurs via electron transfer from the sulfide ion to dissolved oxygen on the pyrite mineral surface, with polysulfides being formed as an intermediate oxidation product. In the absence of cyanide, the polysulfides are further oxidised to thiosulfate, whilst with cyanide present, thiocyanate and sulfite are also formed from the reaction of polysulfides with cyanide and dissolved oxygen. Polysulfide chain length has been shown to affect the final reaction products of polysulfide oxidation by dissolved oxygen.The rate of pyrite catalysed sulfide ion oxidation was found to be slower in cyanide solutions compared to cyanide free solutions. Mixed potential measurements indicated that the reduction of oxygen at the pyrite surface is hindered in the presence of cyanide. The presence of sulfide ions was also found to activate the pyrite surface, increasing its rate of oxidation by oxygen. This effect was particularly evident in the presence of cyanide; in the presence of sulfide the increase in total sulfur from pyrite oxidation was 2.3 mM in 7 h, compared to an increase of <1 mM in the absence of sulfide over 24 h.     

Copper leaching from chalcopyrite concentrate in Cu(II)/Fe(III) chloride system

This study was conducted to develop a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper leaching from chalcopyrite concentrate in ferric and cupric chloride system was investigated using a two-stage countercurrent leach circuit under a nitrogen atmosphere at 97 °C to minimize the concentrations of cupric and ferric ions in pregnant leach solution for subsequent copper solvent extraction while maintaining a maximum copper extraction. A high calcium chloride concentration (110–165 g/L) was used to maintain a high cuprous solubility and enhance copper leaching. With 3–4 h of leaching time for each stage, the copper extraction reached 99% or higher while that of iron was around 90%. With decreasing concentrate particle size from p80 of 26 to 15 μm, the copper extraction increased by about 0.2% while the iron extraction increased by about 2.0%. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was able to be reduced to 0.04 M. When the cupric concentration fell below the above limiting value, the elemental sulfur present was reduced by cuprous ions to form copper sulfide, eventually stopping the leaching of copper. Under this condition, only iron was leached. A very small amount of sulfur (1.2–1.4%) was oxidized to sulfate, resulting in an increase from 3 to 9 g/L in HCl concentration. The extractions of trace metals (Cr, Pb, Ni, Ag and Zn) were 96–100%.     

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.     

Copper–gold ore processing with ion exchange and SART technology

Anglo Asian Mining has developed a 50,000 oz Au/yr open pit gold mine at Gedabek in Western Azerbaijan. The deposit at Gedabek is a copper–gold porphyry, comprising both oxide and sulphide ore mineralisation, which is being mined at the rate of about 1 million tons of ore per year. Ore processing is by conventional cyanide heap leaching, which produces a pregnant leach solution (PLS) containing 1–2 ppm of gold, together with 1000 ppm or more of copper. The PLS is treated by column ion exchange, using Dow’s gold-selective MINIX resin. Loaded resin is stripped with an acidic thiourea solution, from which gold and silver are electrowon on to stainless steel mesh cathodes. Copper concentrations in the leach solutions are controlled by passing part of the PLS flow through a SART process, where the acronym stands for “Sulphidisation, Acidification, Recycling and Thickening”. The product from the SART process is a copper/silver sulphide precipitate, which is thickened, filtered and dried and then sold for copper smelting.     

Selective elution of the gold cyanide complex from anion exchange resin using mixed solvents

This study is concerned with the use of mixed solvents for the elution of the cyanide complexes of copper and gold from Purolite A500, a strong-base anion exchange resin. The mixed solvents investigated include acetone + water, dimethylsulfoxide + water and N-methyl-2-pyrrolidone + water. Three types of counterions are employed in each of the mixed solvents: CN⁻, Cl⁻ and OH⁻. The effects of counterion concentration and mixed solvent composition on the elution of the complexes are examined. High recoveries of the gold cyanide complex are achieved in the mixed solvents at relatively low counterion concentrations. In contrast, the recoveries of the copper cyanide complexes are 1–3 orders of magnitude lower for the given initial loading of the metals on the resin. The selectivity of the elution process for gold is discussed in terms of the degree of solvation of the various anions in the mixed solvents. The results of this study point to the possibility of using mixed solvents to develop an elution process that is selective for gold over multivalent cyanide complexes.     

Salt-tolerant microorganisms potentially useful for bioleaching operations where fresh water is scarce

Anions and cations can accumulate in process waters due to the source water, to evaporation and because of gangue mineral dissolution. Common salts that can accumulate from mineral gangues are the anions, chloride and sulfate, both of which impact on microorganisms used in bioleaching processes. The search for salt-tolerant acidophilic microorganisms able to tolerate high concentrations of these salts, as well as high concentrations of metals, has been underway for at least 20 years because their application would considerably improve bioleaching process efficiency in areas where fresh water is scarce.A thorough search of microorganisms from saline and acidic drains, lakes and sediments in the South West of Western Australia and a CSIRO culture collection was undertaken to bio-prospect for salt-tolerant bioleaching cultures. Pure strains of common bioleaching acidophiles did not tolerate seawater salinities, however, enrichment cultures of mesophilic acidophilic microorganisms that could tolerate up to 70 g/L sea salts and 350 g/L MgSO₄⋅7H₂O were established. The salt tolerance of acidophiles was less in thermophilic temperature ranges, compared with mesophilic and moderately thermophilic temperature ranges. Tolerance to sulfate salts was greater than chloride with magnesium ions likely limiting maximum sulfate tolerance. Iron oxidising cultures were more sensitive than sulfur oxidising cultures to higher chloride concentrations. The addition of pyrite to enrichment cultures increased salt tolerance. The efficacy of the salt tolerant cultures to extract copper will be determined in bioleaching experiments with chalcopyrite ore and salty process water.     

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.     

Oxidation of cyanide in effluents by Caro’s Acid

Caro’s Acid (peroxymonosulphuric acid: H₂SO₅) is a powerful liquid oxidant made from hydrogen peroxide that has been adopted for the detoxification of effluents containing cyanides in gold extraction plants in recent years.The present work reports the findings of a study on the kinetics of aqueous cyanide oxidation with Caro’s Acid. Experiments were conducted in batch mode using synthetic solutions of free cyanide. The experimental methodology employed involved a sequence of two 2³ factorial designs using three factors: initial [CN⁻]: 100–400 mg/L; H₂SO₅:CN⁻ molar ratio: 1–1.5–3–4.5; pH: 9–11; each one conducted at one level of Caro’s Acid strength which is obtained with the H₂SO₄:H₂O₂ molar ratio used in Caro’s Acid preparation of 3:1 and 1:1. The objective was the evaluation of the effect of those factors on the reaction kinetics at room temperature. Statistical analysis showed that the three investigated variables were found to be significant, with the variables which affected the most being the initial [CN⁻] and the H₂SO₅:CN⁻ molar ratio. The highest reaction rates were obtained for the following conditions: H₂SO₅:CN⁻ molar ratio = 4.5:1; pH = 9; and Caro’s Acid strength produced from the mixture of 3 mol of H₂SO₄ with 1 mol of H₂O₂. These conditions led to a reduction of [CN⁻] from an initial value of 400 mg/L to [CN⁻] = 1.0 mg/L after 10 min of batch reaction time at room temperature. An empirical kinetic model incorporating the weight of the contributions and the interrelation of the relevant process variables has been derived as: −d[CN⁻]/dt = k [CN⁻]^1.8 [H₂SO₅]^1.1 [H⁺]^0.06, with k = 3.8 (±2.7) × 10⁻⁶ L/mg min, at 25 °C.     

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.     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).     

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.     

Zinc precipitation by heavy-metal tolerant sulfate-reducing bacteria enriched on phosphogypsum as a sulfate source

In the present study, heavy-metal tolerance and precipitation by a mixed culture of sulfate-reducing bacteria (SRB) were evaluated. These bacteria have been enriched during a previous study from a sewage sludge using phosphogypsum as sulfate source. Taking into account that both sulfate and zinc are naturally occurring in phosphogypsum, zinc tolerance of SRB was tested in synthetic media containing 20 mM sulfate and zinc chloride at concentrations ranging from 0 to 200 mg L⁻¹. Zinc tolerance was determined by bacterial growth susceptibility and zinc removal monitoring. Bacterial growth and sulfate reduction were possible between 10 and 150 mg L⁻¹ of initial zinc concentration. Zinc concentrations more than 150 mg L⁻¹ were lethal to SRB. Zinc was removed effectively by SRB to less than 5% from medium containing 150 mg L⁻¹ initial zinc concentrations or less. Energy-dispersive X-ray analysis showed that precipitation of zinc occurred in the form of sulfide. The results presented in this paper have shown that this mixed culture might be of use for bioremediation of sulfate and heavy-metals containing wastewaters.     

Removal of cyanide in aqueous solution by oxidation with hydrogen peroxide in presence of activated carbon prepared from olive stones

This work is dedicated to the removal of the very toxic free cyanide from aqueous solution by oxidation with hydrogen peroxide H₂O₂ in the presence of activated carbon prepared from olive stones. Effects of the initial molar ratio [H₂O₂]₀/[CN^?]₀, the initial cyanide concentration, the activated carbon concentration and the temperature on cyanide removal have been examined. The removal of free cyanide in absence of activated carbon showed very slow kinetics. The presence of activated carbon has increased the reaction rate showing thus a catalytic activity. The kinetics of cyanide removal has been found to be of pseudo-first-order with respect to cyanide and the rate constants have been determined for different values of the aforementioned parameters. The apparent activation energy has been determined from tests carried out at three different temperatures. It was found equal to 46.2 kJ/mol in the presence of activated carbon, which is about half of the 82.7 kJ/mol found for the oxidation in absence of the activated carbon.This process can be interesting for the cyanide removal from processed solutions because it does not use soluble metal catalyst and it consumes only hydrogen peroxide as chemical product.     

萃取-电沉积处理含铜氰化废水回收铜和氰化物

以季铵盐N263为萃取剂,采用萃取—电沉积工艺对铜氰废液中的铜和氰化物进行回收。结果表明,N263对含氰溶液中的铜氰配合离子有良好的萃取能力,在高碱性条件下其对铜的单级萃取率仍超过90%;饱和负载有机相经反萃可为后续电沉积提供高浓度含铜溶液;提高电沉积温度有利于铜的回收与氰化物的保护;处理后尾液可直接用于氰化浸出。通过萃取—电沉积工艺实现了废水中铜和氰化物的综合回收利用。