Acid leaching and rheological behaviour of a siliceous goethitic nickel laterite ore: Influence of particle size and temperature

This study investigates the isothermal, batch, H₂SO₄ acid leaching behaviour of siliceous goethitic (SG) nickel (Ni) laterite ore and its links to pulp rheology. Specifically, the effect of feed ore particle size (−0.2 vs −2.0 mm), leaching temperature (70 vs 95 °C) and pulp rheology on Ni and pay metal, cobalt (Co) extraction kinetics and yield was studied for 4 h on 40 wt.% solid dispersions at pH 1. The leaching behaviour was distinctly incongruent, reflecting the disproportionate proliferation of major gangue mineral’s constituent elements (e.g., Fe, Al, Mg, Na, Si) alongside Ni and Co in the pregnant leach solution. At 70 °C, Ni/Co extraction rates were notably lower (<20%) in contrast with 95 °C where a significant increase in Ni/Co extraction to 78/77% and 74/77%, respectively, for the −0.2 and −2.0 mm feeds occurred. The slurries displayed a non-Newtonian, shear thinning Bingham plastic rheological behaviour of which the viscosity and shear yield stress increased markedly in the course of 4 h leaching. The pulp viscosity and shear yield stress were greater at lower temperature than at higher temperature and they were also greater in slurries with finer than coarser feed particles. The dynamic pulp rheology, however, had no marked effect on the overall Ni/Co extraction rates. Whilst the feed ore particle size had no remarkable impact on overall Ni/Co extraction, it led to noticeably higher acid consumption and enhanced slurry rheology in the finer sized ore. The mechanism of leaching the SG ore followed a two-stage, first order chemical reaction-controlled shrinking core model, the kinetics of which gave higher rate constants and lower activation energies for the release of Ni, Co, Fe and Mg in the first stage. A faster leaching process involving more reactive minerals during the first 30 min is envisaged to be followed by leaching of the more refractory minerals.     

Determination of ferrosilicon medium rheology and stability

Ferrosilicon (FeSi) has a fast settling rate in dense suspension, attributed to its very high solids density, coarse particle size, more spherical particle shape and low medium viscosity. The fast-settling nature in dense suspension is a challenge to acquire reliable rheological data. A testing rig was set up to maintain a constant medium density during the rheology measurement by circulating the medium to keep FeSi particles well suspended. A rotational rheometer with a modified concentric bob-cup measuring system was incorporated in the testing rig. Taylor number (T_a) was calculated and a threshold T_a = 41.3 was used to identify the onset of unstable laminar flow owning to the Taylor vortices and turbulence formation in the measuring system. The data with T_a > 41.3 were excluded in the determination of rheological flow curves. Evaluation of the modified measuring system with Newtonian silicone oils of known viscosities confirmed that the system can produce true flow curves over the entire tested shear rate range for a stable laminar flow (T_a < 41.3). This data reduction procedure was applied to the FeSi medium rheology measurement. It demonstrates that FeSi medium exhibits a pseudoplastic trend with a yield stress. The apparent viscosities were calculated at two shear rates, 10 s⁻¹ that is assumed to be a typical shear rate in dense medium bath separators and 75 s⁻¹ for dense medium cyclones. A characteristic curve between apparent viscosity and medium density was established, which can be used in FeSi selection for dense medium separation. Medium stability was determined from the FeSi sedimentation rate measurement. It shows that medium stability was closely correlated with medium viscosity. A trade-off between stability and viscosity for optimal dense medium separation should be established.     

Reactivity and mineralogical evolution of an underground mine sulphidic cemented paste backfill

The reactivity of highly sulphidic tailings (≈53% pyrite) used in a cemented paste backfill (CPB) was investigated at the Laronde mine (Quebec, Canada). Oxygen consumption (OC) tests were performed directly in a mine stope filled with CPB. In situ OC test results showed a high oxidation rate at the beginning of the testing period (mean value of 2.4 mol O₂/m²/day). However, the observed oxidation rate decreased progressively to reach a flux value of 0.2 mol O₂/m²/day after 80 days. The physical and mineralogical characterization of the CPB and the pore water quality evolution indicated that this reduction in the oxidation rate for the Laronde CPB can be related to the high degree of saturation maintained in the material, the possible coating of the pyrite grains, and the formation of a thin oxidized layer having a lower porosity (and lower diffusion coefficient) than the bulk CPB. For the Laronde CPB, the addition of binder to the sulphidic tailings appears to have a positive environmental effect on reactive backfill environmental behaviour.     

Study of silver leaching with the thiosulfate–nitrite–copper alternative system: Effect of thiosulfate concentration and leaching temperature

Thiosulfate system is considered an interesting alternative leaching process for precious metals. Nevertheless, most of the literature published on these conventional thiosulfate leaching solutions has been focused on the use of ammonia and copper to generate the cupric tetraamine complex, which acts as a catalytic oxidant for silver. However, ammonia toxicity is also a detrimental issue in terms of the process sustainability. For that reason, thiosulfate–nitrite–copper solutions were studied as an alternative less toxic system for silver leaching.In this work, the effect of the thiosulfate concentration (0.07 M, 0.1 M and 0.15 M) and temperature (room temperature, 30, 35, 40 and 45 °C) on the metallic silver leaching kinetics is presented for the S₂O₃–NO₂–Cu system. The results show that the thiosulfate concentration plays an important role in the S₂O₃–NO₂–Cu–Ag system since it controls the silver leaching kinetics. On the other hand, an increase in temperature favors the silver recovery.Finally, the SEM–EDS analysis, the X-ray mapping and the X-ray diffractograms show that the solid silver particles are coated by a Cu, S and O layer for the 0.07 M and 0.1 M thiosulfate experiments, which is consistent with the formation of antlerite (Cu₃(SO₄)(OH)₄); while the 0.15 M thiosulfate scenario produced a layer composed only of Cu and S, revealing the formation of stromeyerite (CuAgS). The UV–Visible technique confirmed the in-situ generation of copper–ammonia complexes for the 0.07 M leaching condition; however, these complexes are not formed at the 0.15 M condition.     

Single and binary component sorption of the fission products Sr2+, Cs+ and Co2+ from aqueous solutions onto sulphate reducing bacteria

This study investigates the removal of the fission products Sr²⁺, Cs⁺ and Co²⁺ in single and binary metal solutions by a sulphate reducing bacteria (SRB) biomass. The effect of initial concentration and pH on the sorption kinetics of each metal was evaluated in single metal solutions. Binary component equilibrium sorption studies were performed to investigate the competitive binding behaviour of each metal in the presence of a secondary metal ion. Results obtained from single metal equilibrium sorption studies indicated that SRB have a higher binding capacity for Sr²⁺ (q_max = 416.7 mg g^?1), followed by Cs⁺ (q_max = 238.1 mg g^?1), and lastly Co²⁺ (q_max = 204.1 mg g^?1). Among the binary systems investigated, Co²⁺ uptake was the most sensitive, resulting in a 76% reduction of the sorption capacity (q_max) in the presence of Cs⁺. These findings are significant for future development of effective biological processes for radioactive waste management under realistic conditions.     

Agglomeration and column leaching behaviour of goethitic and saprolitic nickel laterite ores

Processability of complex, low-grade nickel (Ni) laterite ores via heap leaching is very limited due to some intractable geotechnical and hydrological challenges such as poor heap porosity/permeability and structural stability. This work presents some investigations on laboratory batch drum agglomeration and continuous column leaching behaviour of saprolitic (SAP) and goethitic (G) Ni laterite ores as part of the quest for an effective ore pre-treatment process for enhanced heap leaching. As a focus, the effect of ore mineralogy/chemistry on the agglomeration and column leaching behaviour of −2 mm (crushed from −15 mm run-of-mine) G and SAP Ni laterite ores was examined. To produce ∼5–40 mm agglomerates in <15 min, the SAP ore required a higher H₂SO₄ (30 wt.%) binder dosage compared with the G ore, although both ores displayed substantially similar, coalescence-controlled agglomeration mechanism. The resulting G agglomerates were more robust than the SAP ones based upon their compressive strength and acidic solution soak test measurements. However, over 100 days of continuous column leaching, the structural stability of the SAP agglomerate bed was slightly greater than that of G agglomerates, reflecting a lesser slump of the former. The pregnant leach solution analysis revealed greater Ni/Co extraction rates from the SAP than the G agglomerates. Whilst the total mass of acid consumed per ton dry ore processed was greater for the SAP ore, the total kg acid per kg Ni extracted was markedly lower. Incongruent leaching of gangue minerals’ constituent elements (e.g., Fe, Mn, Mg, Al and Si) occurred and contributed significantly to the overall acid consumption. The findings show the relevance of agglomeration and column leaching tests for providing useful information for plant designing and optimization of Ni laterite heap leaching operations.     

Adsorption of copper sulphate on PGM-bearing ores and its influence on froth stability and flotation kinetics

Copper sulphate is used as an activator in the flotation of base metal sulphides as it promotes the interaction of collector molecules with mineral surfaces. It has been used as an activator in certain platinum group mineral (PGM) flotation operations in South Africa although the mechanisms by which improvements in flotation performance are achieved are not well understood. Some investigations have suggested these changes in flotation performance are due to changes in the froth phase rather than activation of minerals by true flotation in the pulp zone. In the present study, the effect of copper sulphate on froth stability was investigated on two PGM containing ores, namely Merensky and UG2 (Upper Group 2) ores from the Bushveld Complex of South Africa. Froth stability tests were conducted using a non-overflowing froth stability column. Zeta potential tests and ethylenediaminetetraacetic acid (EDTA) tests were used to confirm the adsorption of reagents onto pure minerals commonly found in the two ores. The results of full-scale UG2 concentrator on/off copper sulphate tests are also presented. The UG2 ore showed a substantial decrease in froth stability in the order of reagent addition: no reagents > copper > xanthate > copper + xanthate, while Merensky ore showed a slight decrease. It was shown through zeta potential measurements that copper species were to be found on plagioclase, chromite, talc and pyrrhotite surfaces and through EDTA extraction that this copper was in the form of almost equal amounts of Cu(OH)₂ and chemically reacted copper ions on the Merensky and UG2 ore surfaces. In certain cases, the presence of copper sulphate and xanthate substantially increased the recovery, and therefore the implied hydrophobicity, of pure minerals in a frothless microflotation device. It was, therefore, proposed that increases in hydrophobicity beyond an optimum contact angle for froth stability, were the cause of instabilities in the froth phase and these were found to impact grade and recovery in a full-scale concentrator. Differences in the extent of froth phase effects between the different ores can be attributed to differences in mineralogy.     

Modeling breakage rates in mills with impact energy spectra and ultra fast load cell data

A new era in modeling particle size distribution in grinding mills started at the beginning of 2000s. A direct estimation of breakage parameters became possible via computation of collision energy by discrete element method (DEM) and material breakage data.The material breakage data can be obtained for primary modes of breakage. In this study, impact and abrasion are assumed to be the primary modes of particle breakage, which are readily studied in the laboratory. The impact breakage mode is studied in a drop-weight apparatus and in a specialized device known as the ultra fast load cell. The abrasion mode of breakage is studied in a laboratory scale ball mill. Next, the particle breakage versus energy data is converted into breakage rates via impact energy spectra of the grinding mill computed by a DEM code. The fundamental material breakage information is converted into energy based breakage distribution function.The verification of the modeling concepts is shown for a 90 cm laboratory scale ball mill. In the batch mill, approximately a 10 kg mass of limestone in the 30 mm size is ground with around 100 kg of 50 mm steel ball charge. The breakage rate and the breakage distribution functions constitute the parameters of the energy based batch population balance model. It is shown that accurate particle size distribution predictions are possible with this modeling approach for different grinding regimes.     

On the nanobubbles interfacial properties and future applications in flotation

Nanobubbles, generations forms, basic studies and applications constitute a growing research area, included their usage in advanced mineral flotation. Yet, there are investigation needs for sustainable generation procedures, stability and understanding the nanobubbles interfacial properties and structures. Results proved that a reduction in pressure makes the super-saturated liquid suffers cavitation and nanobubbles were generated. Medium pH and solutions tested were adjusted, in the air saturation vessel, before the nanobubbles were formed, and this allowed to control (in situ) the surface charge/zeta potential-size of the forming nanobubbles. Measurements obtained with a ZetaSizer Nano equipment showed zeta potential values, in the presence of 10⁻² mol L⁻¹ NaCl, displaying sigmoidal pH behaviour between pH 2 (+26 mV) and 8.5 (−28 mV); an isoelectric point was attained at pH 4.5 and were positively charged (up to 23 mV) in acidic medium, a phenomenon which has not been previously observed. In alkaline medium, bubbles were highly negative zeta potential (−59 mV) at pH 10. The double layers appear to play a role in the formation of stable nanobubbles providing a repulsive force, which prevents inter-bubble aggregation and coalescence. Accordingly, the sizes of the nanobubbles depended on their charge and increased with pH, reaching a maximum (720 nm) around the isoelectric point (±5 mV). Highly charged and small nanobubbles (approximately 150–180 nm) were obtained in the presence of surfactants (10⁻⁴ mol L⁻¹ of alkyl methyl ether monoamine or sodium dodecyl sulphate); the zeta potential values in these experiments followed a similar trend of other reported values, validating the technique used with the nanobubbles sizes varying with pH from 150 to 400 nm. Thus, charged and uncharged stable nanobubbles can be tailor-made with or without surfactants and it is expected that their use will broaden options in mineral flotation especially if collectors coated nanobubbles (“bubble-collectors”) were employed. A detailed and updated review on factors involving stability, longevity and coalescence of nanobubbles was made. It is believed that future trend will be on sustainable formation and application of nanobubbles at industrial scale contributing to widen applied research in mineral, materials processing and liquid effluent treatment by advanced flotation.     

Adsorption of polysaccharide onto talc

The interaction of polysaccharide with talc has been investigated through adsorption, flotation, and electrokinetic measurements. The adsorption densities are independent of pH and the isotherms exhibit Langmuirian behavior. The order of adsorption density of several polysaccharides investigated in the paper onto talc is CMC > CMS > dextrin. Pretreatment of talc with EDTA (ethylenediaminetetraacetic acid), leaching and calcinations result in a decrease in the adsorption density, highlighting the importance of metallic magnesium sites for the adsorption process. An increase in the surface face-to-edge ratio leads to increase in adsorption density. The flotation recoveries are independent of pH. CMC exhibit best depressant ability, followed by CMS and then dextrin among the three depressants, complementing the adsorption results. However, polysaccharide depressant ability is reduced in the case of leached and calcined talc sample in comparison to that on talc. Electrokinetic measurements portray conformational rearrangements of macromolecules with the loading, resulting in the shift of the shear plane, further away from the interface. The adsorption process is governed by hydrogen bonding as well as chemical interaction between the polysaccharides and the surface metal hydroxide groups of talc. The best depressant activity of CMC may be attributed to its more favorable carboxyl groups as opposed to the hydroxyl groups of dextrin, apart from its higher molecular weight.     

Dissolution of gold during pyrite oxidation reaction

Numerous papers discuss the mechanism of alkaline oxidation of pyrite but there is limited information available describing the actual kinetics of the pyrite sulphide to thiosulphate reaction. A previous investigation in this series determined the rate of sulphide sulphur oxidation and thiosulphate yield in the reaction of pyrite with sodium hydroxide under various testing conditions. The goal of the current study is to validate these rates using two different gold-containing pyrite concentrates. A further objective of the current work is to investigate the simultaneous dissolution of gold with in situ formed thiosulphate during pyrite oxidation.It was found that at 20 psi oxygen overpressure and a temperature of 80 °C, the initial rate of sulphide oxidation and thiosulphate yield were close to 0.08 mol/h and 0.0155 mol/h, respectively. These rates are in agreement with previously published data. However, a shift from linearity occurred when the pH decreased below 12. A rapid decay of thiosulphate was evidenced at pH 8.3–9.2 while E_H was in the range of 22–141 mV. Based on relevant thermodynamic analysis of metastable thiosalts system, such rapid decomposition is not expected at these pH and E_H values. It is believed that the presence of unreacted pyrite acting as a catalyst caused this behaviour. It appears that under mildly alkaline conditions, the rate of oxidation of sulphide to thiosulphate becomes slower than the rate of thiosulphate degradation, which causes a net loss of thiosulphate in the system. The maximum extraction of gold and silver (96% and 75% respectively) was achieved under conditions of pH < 12.     

Modification of ilmenite surface properties by superficial dissolution method

Acid surface dissolution as a pretreatment method converts Fe²⁺ ions on the ilmenite surface to Fe³⁺ ions. XPS analysis showed that the content of Fe³⁺ increases from 48.5% to 59.8% after surface dissolution for 15 min in a solution of sulfuric acid with a concentration of 10%. This conversion, without any phase transformation, decreases the zeta potential of ilmenite in a wide pH range, resulting in a shift in IEP (Iso-Electric Point) from a pH of 5.4 to 2.3. FTIR spectra and zeta potential measurements showed that the increase of oleate ions adsorption on the ilmenite surface, resulting from the surface dissolution process, is insignificant. After surface dissolution, the formation of more ferric iron oleate species (Ksp = 10^−29.7) being more stable than ferrous iron oleate (Ksp = 10^−15.5) compounds yields an increase of ilmenite hydrophobicity and floatability in a wide pH range. Using 3.65 × 10⁻⁴ M sodium oleate at a pH of 6.3, the maximum flotation recoveries are obtained as 73.5% and 92% for non-treated and acid pretreated ilmenite, respectively.     

Adsorption of cationic collectors and water on muscovite (0 0 1) surface: A molecular dynamics simulation study

A series of molecular dynamics simulations have been performed employing PCFF_phyllosilicates force field. The interaction energies between water molecules/ammonium ions and the muscovite (0 0 1) surface are computed. Results drawn from the calculation show that ammonium ions have thermodynamic advantages to resist the hydration layer sufficiently for effective flotation to occur. Adsorption sites of water molecules and amine ions have also been considered in detail.     

Biosorptive flotation of copper ions from dilute solution using BSA-coated bubbles

Copper adsorption was carried out using the novel material known as air-filled emulsion (AFE). AFE is a stable colloidal system containing microscopic protein-coated bubbles (<10 μm) dispersed through an aqueous solution, resulting in an increased specific surface area and contact time between extractant and metal ions. Bovine serum albumen (BSA) generated emulsion concentration had a significant impact on copper removal, with maximum metal uptake obtained at 2.5 g/l of BSA-coated bubbles. It was shown that copper sorption was rapid over the first 10 min, and equilibrium conditions were reached within 40 min. Separation of the copper-loaded microcells from the aqueous solution was also investigated. Micro-flotation was employed to remove the microbubbles by means of attachment to the surface of larger air bubbles. In absence of a cationic surfactant, approximately 0.5% copper recovery was obtained at pH ranging from 5 to 8 due to the lack of hydrophobic groups on the surface of Cu-loaded BSA emulsions. Due to the fine sizes of the emulsion bubbles (<10 μm) a cationic flocculant was used to induce coagulation of the bubbles leading to easier phase separation. A combination of collector and flocculant at a concentration of 3 × 10⁻⁴ M and 0.025 g/l, respectively, led to an increase in copper recovery to nearly 35% at pH 7.     

Liberated 0–10 μm particles from sulphide ores,their production and separation—Recent developments and future needs

The general types of industrial behaviour of liberated valuable sulphide minerals in the 0–10 μm size fraction are described and tests to explore on-site at a concentrator the general nature of any deficient behaviour are discussed. In addition, the general types of behaviour of the gangue minerals in this size fraction are also described in terms of their mechanisms for recovery.These behaviours are also discussed in terms of the type of grinding device and grinding environment in which they were produced and the target grinding product size. Major recent improvements in the industrial capabilities of grinding devices are discussed.Present difficulties in characterization of particles in the 0–10 μm range are described and desirable improvements for the future are outlined. Further, desirable improvements for industrial processing of valuable minerals in this size range are discussed along with some remedial actions.     

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.     

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.     

Dynamic modelling and simulation of semi-autogenous mills

This paper presents a dynamic simulator of the semi-autogenous grinding operation deduced from first principles coupled to an on-line parameter estimation scheme able to simulate industrial operations for control purposes. Model equations are based on conventional non-stationary population balance approach to develop the necessary dynamic model of the semi-autogenous mill operation. The model is able to predict the time-evolution of key operating variables such as product flow rate, level charge, power-draw and load position, as function of other important variables such as mill rotational speed and fresh feed characteristics. The set of ordinary differential equations was solved using MATLAB/SIMULINK as a graphic programming platform, a very helpful tool for understanding the grinding phenomenon.Using an on-line parameter estimator, the simulator was tested and validated at an 1800 t/h copper–ore mill showing the effectiveness of the system to track the dynamic behaviour of the variables. It is now being used for the study of advanced control schemes.     

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%.