The influence of pyrite pre-oxidation on gold recovery by cyanidation

The influence of pyrite pre-oxidation in alkaline solutions on gold recovery by cyanidation from Twin Creek refractory gold ore in which pyrite was identified as the major sulfide mineral has been investigated with the aid of electrochemical measurements, leaching experiments, and direct analysis of reaction products for selected residues. It was found that gold recovery by cyanidation in bottle roll experiments mainly depended on the extent of pyrite pre-oxidation. The rate of pyrite oxidation in alkaline solutions measured by electrochemical measurements, including chronoamperometry and linear sweep voltammetry, increased with an increase in pH, potential, and temperature. All alkaline reagents used for the electrochemical measurements, NaOH, NH₄OH, Na₂CO₃ and Ca(OH)₂, showed a similar effect on pyrite oxidation kinetics. However, the results of alkaline pre-oxidation for pyrite of the Twin Creek refractory gold ore suggested that NaOH and Na₂CO₃/Ca(OH)₂ were superior to Ca(OH)₂. Without pre-oxidation, cyanide leachable gold was found to be only 20% which could be increased to 70% under appropriate pre-oxidation conditions. At the same time, cyanide consumption decreased from 2.5 kg/t ore to 1.5 kg/t ore.Selected residues after pre-oxidation and cyanidation were examined by X-ray diffraction. Backscattered electron images of pyrite particles in these residues were taken. The reaction products at the surface of pyrite particles were found to be iron-, silicon-, and calcium-bearing compounds with variable amounts of sulfur as determined by X-ray energy dispersion analysis. Additionally, some mineral fines, such as aluminum and/or potassium-bearing minerals, were found to be present at the partially oxidized pyrite surface.     

Thiosulphate leaching of gold in the presence of ethylenediaminetetraacetic acid (EDTA)

The addition of low levels of ethylenediaminetetraacetic acid (EDTA) in the ammoniacal thiosulphate gold leach system lowered the catalytic cupric/cuprous redox equilibrium potential, hence the mixed solution potential and reduced the consumption of thiosulphate. In the leaching of pure gold, gold dissolution was enhanced in the presence of EDTA at a relatively low concentration, but excessive EDTA decreased gold dissolution. Raman analysis of the leached gold foil indicated that the stabilisation of thiosulphate by EDTA decreased the formation of the passivation layers of elemental sulphur and copper sulphide at the gold surface. In the leaching of a sulphide ore, the leaching kinetics and overall extractions of gold and silver were enhanced substantially, while the consumption of ammonium thiosulphate was reduced from 9.63 kg/t to 3.85 kg/t in the presence of 2.0 mM EDTA after 24 h leaching. This beneficial effect became more pronounced at a higher EDTA concentration. The enhanced gold and silver extractions by EDTA were attributed to the increase in the dissolution of gold and silver bearing sulphides, the stabilisation of copper and thiosulphate in leach solutions, the prevention of leaching passivation and the decrease in the interference of foreign heavy metal ions. The use of EDTA in the ammoniacal thiosulphate leaching system makes it practical to achieve satisfactory gold extraction over extended periods of leaching under low reagent concentrations, where the consumption of thiosulphate is low.     

Review of rate constants for thiosulphate leaching of gold from ores,concentrates and flat surfaces: Effect of host minerals and pH

A review of literature data for different types of sulphide concentrates and gold ores has been carried out to examine the impact of host minerals and pH upon gold leaching. Analysis of initial rate data over the first 30–60 min of gold leaching from sulphide concentrates or silicate ores over a range of ammonia, thiosulphate, and copper(II) concentrations, pH (9–10.5) and temperatures up to 70 °C shows the applicability of a shrinking sphere kinetic model with an apparent rate constant of the order k_ss = 10⁻⁶–10⁻³ s⁻¹. The dependence of apparent rate constant on pH and initial concentrations of copper(II) and thiosulphate is used to determine a rate constant k_Au(ρr)⁻¹ of the order 1.0 × 10⁻⁴–7.4 × 10⁻⁴ s⁻¹ for the leaching of gold over the temperature range 25–50 °C (ρ = molar density of gold, r = particle radius). These values are in reasonable agreement with rate constants based on electrochemical and chemical dissolution of flat gold surfaces: k_Au = 1.7 × 10⁻⁴–4.2 × 10⁻⁴ mol m⁻² s⁻¹ over the temperature range 25–30 °C. The discrepancies reflect differences in surface roughness, particle size and the effect of host minerals.     

Self-heating activation energy and specific heat capacity of sulphide mixtures at low temperature

Energy of activation (E_a) and specific heat capacity (C_p) for mixtures of sulphide minerals that on their own do not self-heat (SH), sphalerite/pyrite, pyrite/galena, chalcopyrite/galena and sphalerite/galena, were determined using a self-heating apparatus at temperatures below 100 °C in the presence of moisture. The mixtures all gave E_a ranging from 22.0 to 27.8 kJ mol⁻¹, similar to the range reported for Ni- and Cu-concentrates. The E_a is close to that for partial oxidation of H₂S which adds to the contention that the partial oxidation of H₂S contributes to SH of sulphides at low temperature. The C_p values ranged from 0.152 to 1.071 JK⁻¹ g⁻¹ as temperature rose from 50 °C to 80 °C, similar to the reported findings on Ni- and Cu-concentrates. The role of galvanic interaction in promoting SH is tested by examining correlations with the rest potential difference of the sulphides in the mixture.     

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.     

Simultaneous sulfide oxidation and gold leaching of a refractory gold concentrate by chloride–hypochlorite solution

In this research, oxidation of sulfide and leaching of gold from a gold–bearing sulfide concentrate using chloride–hypochlorite solution has been investigated. Effects of calcium hypochlorite concentration, initial pH and sodium chloride concentration on the recovery of gold were examined. Two conditions were considered; the stability range of the gold complex (Eh > 900 mV) and formation of chlorine gas (pH < 3.5). During leaching, due to oxidation of sulfide and generation of acid, pH dropped. About 82% of gold was extracted from 200 g/L concentrate after 2 h using 200 g/L Ca(OCl)₂, 200 g/L NaCl at initial pH of 11, stirring speed of 600 rpm and temperature of 25 °C.     

Two-step oxidation of a refractory gold-bearing sulfidic concentrate and the effect of organic nutrients on its biooxidation

Methods for improving the treatment efficiency of a refractory gold-bearing sulfidic concentrate are proposed. These methods consist of the oxidation of the concentrate during a two-step process, which includes a high temperature ferric leaching step and a subsequent biooxidation step, and the use of organic nutrients during the biooxidation step. The concentrate contained 34.7% pyrite and 7.9% arsenopyrite. The biooxidation of the concentrate (for a one-step process) was conducted at 45 °C in two bioreactors that were connected in series under continuous conditions. The pyrite and arsenopyrite oxidation levels after 240 h were 60.2% and 92.0%, and the gold recovery level by carbon-in-pulp cyanidation was 65.7%. The two-step process included the leaching of the concentrate by a biologically generated Fe³⁺-containing solution and the subsequent biooxidation of the leach residue. In this case, the pyrite and arsenopyrite oxidation levels after 240 h of biooxidation were 65.7% and 94.1%, and the gold recovery level was 71.7%.The effect of an organic nutrient (yeast extract) on biooxidation during the two-step process was studied. The pyrite and arsenopyrite oxidation levels after 240 h of biooxidation under mixotrophic conditions were 73.5% and 95.1%, and the gold recovery level was 77.9%. The effect of the organic nutrient on the microbial population was determined. Sulfobacillus thermosulfidooxidans and Acidithiobacillus caldus were the predominant microorganisms studied under both autotrophic and mixotrophic conditions. Archaeon Acidiplasma sp. MBA-1 was a minor component of the microbial community under autotrophic conditions but was one of the predominant microorganisms studied under mixotrophic conditions. These results suggest that the organic nutrient changed the composition and increased the activity of the microbial population.Thus, a two-step process with organic nutrients added during biooxidation may be considered as an effective strategy for treating refractory pyrite–arsenopyrite concentrates.     

Evaluating waste rock mineralogy and microtexture during kinetic testing for improved acid rock drainage prediction

This study integrates detailed mineralogical and microtextural analyses of waste rock with the results of standard kinetic test procedures to identify the mineralogical changes that influence leachate chemistry over time. The integration of mineralogy and texture provides the opportunity for improved mine waste management strategies and acid rock drainage (ARD) prediction.Waste rock material from an abandoned gold mine in northern Queensland, Australia, was subjected to column leach kinetic testing over a 30 week period. The column feed comprised of a range of waste rock lithologies (porphyritic rhyolite, massive arsenopyrite, massive pyrite ± galena, and semi-massive polysulphide). In total, 12 individual columns were established to represent six lithologies prepared to two different size fractions (−10 mm and −4 mm). The mineralogy and microtextural characteristics of the column feed material was defined using quantitative X-ray diffractometry (QXRD), scanning electron microscopy and laser ablation (LA-ICPMS) at the start of kinetic tests, and at 5 week intervals during the length of the tests. These data were directly correlated with leachate chemistry (i.e., pH, SO₄ and select elements).Results of this study indicated that sulphide oxidation was strongly influenced by the morphology of sulphide minerals, their trace element contents, the presence of mineral micro-inclusions and galvanic interactions with other sulphide minerals. Waste rock with abundant arsenopyrite was consistently the most acid forming, and oxidised to scorodite (enriched in Zn, Pb and Cu). Pyrite was commonly As-rich as indicated by LA-ICPMS mapping. QXRD results indicated that the abundance of rhomboclase, jarosite, alunite and hydrous ferric oxides increased over time. Galena weathered rapidly to porous anglesite, particularly when in direct physical contact with pyrite. Sphalerite contents decreased consistently over the 30 weeks implying its oxidation, however few reaction products were directly observed. By week 30, the −4 mm fraction material generated lower pH leachate, higher mass release of elements and sulphate for the majority of samples. This indicates that the particle size used in kinetic tests can exert a significant control on leachate chemistry, especially in the absence of abundant neutralising minerals. This contribution demonstrates the value of integrating mineralogy and microtextural analyses during kinetic testwork to improve the interpretation of sulphide oxidation for better prediction of ARD.     

Bulk flotation of auriferous pyrite and arsenopyrite by using tertiary dodecyl mercaptan as collector in weak alkaline pulp

Laboratory and industrial scale experiments were conducted to investigate the effect of tertiary dodecyl mercaptan (TDM) as a collector on the flotation of auriferous pyrite and arsenopyrite. The optimum recovery of gold associated with auriferous sulphides was obtained by adding a mixture of TDM and sodium butyl xanthate, together with only a little CuSO₄ as an activator in a weak alkaline pulp adjusted by NaOH. A two-month industrial trial at the Liumei plant in Guangxi, China showed that an average gold recovery of 90.8% into a concentrate assaying 81.1 g/t Au from a feed assaying 2.9 g/t Au could be achieved at pH 8–8.5 using TDM as a collector.     

The leaching and adsorption of gold using low concentration amino acids and hydrogen peroxide: Effect of catalytic ions,sulphide minerals and amino acid type

The leaching of gold using alkaline amino acids–hydrogen peroxide solutions at low concentrations has been studied. The application of alkaline amino acid–hydrogen peroxide system may offer an alternative and environmentally benign process for gold leaching, particularly in the context of leaching low grade gold ores in an in-situ or in heap leach processes. In the presence of an oxidant or oxidants, it was found that amino acids can dissolve gold at alkaline condition at low and moderate temperature. Heating the leach solution between 40 and 60 °C was found to enhance the gold dissolution significantly in alkaline amino acid–peroxide solutions. It was also found that gold dissolution increases by increasing amino acid concentration, peroxide and pH. Amino acids acts synergistically to dissolve gold. Although glycine showed the highest gold dissolution as a single amino acid compared to histidine and alanine, histidine was found to enhance gold dissolution when used in equimolar amounts with glycine. The presence of Cu²⁺ ion enhances gold dissolution in the glycine–peroxide solutions. The process will propose an environmentally benign process for gold treatment in order to replace the use of cyanide in heap or in-situ leaching. In the presence of pyrite, the amount of gold leached was lower due to the peroxide consumption in sulphide oxidation.     

Post-regrind selective depression of pyrite in pyritic copper–gold flotation using aeration and diethylenetriamine

The present study investigates the effect of aeration and diethylenetriamine (DETA) on the selective depression of pyrite in a porphyry copper–gold ore, after regrinding (at grind sizes, d₈₀ = 38 and 8 μm) with respect to Au recovery and grade using oxygen demand tests, flotation, QEMSCAN, X-ray spectroscopy (XPS) and EDTA extraction analysis. It was found that pyrite depression increases after aeration and with decreasing grind size. This was observed to be due to the markedly higher oxygen consumption rate of pyrite at the 8 μm (kla = 0.10 min⁻¹) than at the 38 μm grind size (kla = 0.02 min⁻¹). The addition of DETA improved pyrite depression (9% with aeration only versus 39% with aeration + DETA) at the 38 μm grind size. Gold and copper flotation recovery followed pyrite recovery for the two grind sizes using XD5002 in the presence of air and DETA.The surface analysis (XPS and EDTA extraction) revealed that the significant pyrite depression at the 8 μm grind size was due to increased amount of surface iron oxides, oxy-hydroxides (FeO/OH), sulphate species and increased liberation of mineral phases (QEMSCAN analysis), whilst the poorer pyrite depression at the 38 μm grind size was due to insufficient liberation of mineral phases and the persistence of activating Cu on the pyrite surface. The addition of DETA increased pyrite depression at the coarser grind size due to a significant reduction in Cu(I)S and increased Cu(II)O species, correlating with the flotation results of pyrite under this test condition. Two-stage copper and pyrite flotation, followed by Au cleaning after regrinding to 38 μm grind size, under high pH or aerated condition is proposed as the recommended route to optimise Au flotation.     

Kinetics of high-sulphur and high-arsenic refractory gold concentrate oxidation by dilute nitric acid under mild conditions

On one hand, high-sulphur and high-arsenic refractory gold concentrate (HGC) leads to regional ecological damage. On the other hand, it contains a lot of valuable elements. So utilization of it can bring social and environmental benefits. In this paper, the kinetics of HGC oxidation by dilute nitric acid under mild conditions was investigated. The effects of particle size (50–335 μm), reaction temperature (25–85 °C), initial acid concentration (10–30 wt.%) and stirring speed (400–800 rpm) on the iron extraction rate (C_r) were determined. It is obvious that C_r increases with the rise of initial nitric acid concentration, reaction time and stirring speed, but decreases with the increase of particle size. Oxidation kinetics indicates that the rate of reaction is diffusion controlled. The activation energies were determined to be 10.70 kJ/mol in the 10% HNO₃ and 12.25 kJ/mol in the 25% HNO₃.     

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.     

Release of arsenic from cyanidation tailings

At a gold mine in northern Sweden, gold occurring as inclusions in pyrrhotite and arsenopyrite is leached by cyanidation of the ore. The main sulphide minerals in the ore are pyrrhotite and arsenopyrite. Effluents from the cyanidation process are treated with Fe₂(SO₄)₃ to form Fe-precipitates suitable for the co-precipitation of As. The aim of this study was to perform static and kinetic leaching tests on the ore and tailings to define geochemical processes governing As mobility. Sequential leaching tests suggested that the majority of dissolved As deriving from the sulphide fraction in the ore was incorporated in newly formed Fe-precipitates in the tailings. The mobility of As in the tailings was therefore mainly dependent on the stability of these As-bearing Fe-precipitates. Weathering cell tests (WCT) involving 31 weekly cycles of wetting and air exposure were conducted to assess the stability of the As in the tailings under accelerated weathering conditions. The first stage of the WCT was characterized by a pH ≈ 5 and low As leaching, probably driven by the dissolution of amorphous Fe-As species. In the second stage of the WCT, leaching of Fe, S and As increased and the pH decreased to <3.5. An increase of the leachate’s molar Fe/S-ratio suggested that pyrrhotite oxidation was occurring. The falling pH destabilized As-bearing Fe-precipitates, causing further As release. The total As release during the WCT corresponded to only a small proportion of the tailings’ total As content. The accelerated As-leaching observed towards the end of the WCT could thus indicate that its release could increase progressively over time.     

Pyrrhotite oxidation and its influence on alkaline amine flotation

In the production of ground calcium carbonate (GCC) for the paper industry, pyrrhotite dramatically reduces the GCC brightness and the removal of pyrrhotite through froth flotation is essential. The present study aims to study the effect of pyrrhotite oxidation on flotation recoveries during typical GCC flotation (i.e. alkaline pH, CaCO₃ saturation, amine collector). EDTA extraction and measurements of Eh (redox potential) showed a significant difference in pyrrhotite surface oxidation state when comparing exposure times of 40 and 60 min, the latter being significantly more oxidised. Microflotation results show that when pyrrhotite is exposed to air extending for more than 5 min at pH 8, the recoveries drop significantly. At higher pH, recoveries were generally low at all exposure times tested. Flotation recoveries showed strong correlation with zeta potential measurements. Bench scale flotation experiments on a sulphide bearing marble, confirmed that pyrrhotite oxidation significantly lower the GCC quality at low collector concentrations. By increasing the amine concentration, the flotation performance became independent of pH and exposure time.     

Cyclic voltammetry responses of gold electrodes in thiosulphate electrolyte

The gold thiosulphate system has been studied for some time due its prospective favourable economics over the gold cyanide system when treating certain ore types. Despite the large amount of work which has been done with gold thiosulphate system, various questions (uncertainties) still remain. For example, the causes for slow gold dissolution kinetics in thiosulphate and the gold losses during leaching observed in some applications. In this study, cyclic voltammetry was used to investigate the oxidation and reduction of gold in magnesium thiosulphate electrolyte. The anodic dissolution of gold is discussed in terms of possible reactions and mechanisms involving catalysts such as thallium and thiourea. The abrupt increase in anodic current in the presence of these catalysts indicates that the passivation of gold anodic dissolution is related to surface phenomenon and not to the depletion of thiosulphate ion at the reaction interface. The Tafel-like plots identified the increase of exchange current for gold oxidation and reduction reactions for catalyst assisted systems. The large anodic to cathodic peak separation ranged between 0.25 V and 0.37 V allowed most of the leached gold to diffuse away from the electrode. This indicates the electrochemically quasi-reversible to irreversible character of this system even in the presence of catalysts. The relationship between reduction peak current and concentration of gold thiosulphate in the bulk electrolyte was estimated based on relevant equation for electrochemically irreversible reaction.     

Biohydrometallurgy in Turkish gold mining: First shake flask and bioreactor studies

The first laboratory tests on biooxidation and cyanidation of gold ores in Turkey were carried out using samples of the Copler Gold Mine. Over a 3 year R&D test period, mixed bacterial/archaeal cultures improved biooxidation of the Copler ore. The highest sulphide oxidation of 87.35% over 432 h was achieved in shake flasks in the presence of the mixed culture (MODM: Sulfolobus acidophilus and Sulfolobus thermosulfidooxidans). Bioreactor tests resulted in greater dissolution rates for iron and arsenic than did shake-flask tests, which led to a greater extent of sulphide oxidation within a shorter period of time. The maximum sulphide oxidation in the bioreactor tests was 97.79% after 240 h when the EXTM (Acidianus brierleyi and Sulfolobus metallicus) mixed culture was used. After the biooxidation experiments with solids contents of 10% and 20% (w/v), the gold recovery from the oxidised ore was lower than that achieved in the presence of 5% solids (w/v) because the extent of sulphide oxidation was reduced as the pulp density increased. A strong correlation between the sulphide oxidation and gold recovery was also established. The highest gold recovery of 94.48% was achieved during cyanidation from the biooxidised ore produced from the experiment conducted using the EXTM mixed culture.     

Flotation of pyrrhotite and pyrite in saturated CaCO3 solution using a quaternary amine collector

In the production of ground calcium carbonate (GCC) for the paper industry, any colouring contaminants must be removed during processing (usually flotation) to achieve sufficient GCC brightness. Flotation of GCC feedstock is characterised by the presence of both particulate and dissolved calcite (CaCO₃), an alkaline pH and the use of amine collectors. This paper investigates the possibility of removing pyrrhotite and pyrite, under these conditions. Microflotation results show that the recoveries are highly dependent on pH. Pyrrhotite recovery decreases dramatically when going from pH 8 to pH 10. Pyrite display high recoveries at pH 8 and 9, and lower recoveries at higher pH. Recoveries were dependent on conditioning time. Pyrite floats considerably faster than pyrrhotite at all pH levels investigated, whereas pyrrhotite seems to be slow floating and could benefit from prolonged flotation times. Compared to the saturated CaCO₃ system, flotation recoveries decreased when using deionised water or CaCl₂ solution. The flotation results could not be fully explained by zeta potential. Bench scale flotation experiments on a sulphide bearing marble, confirmed the microflotation results at low collector concentrations (i.e. 200 g/t). By increasing the amine concentration, the flotation performance became independent of pH.     

The selective flotation of arsenopyrite from pyrite

Flotation tests were conducted to compare the arsenopyrite depressing effect of MAA with pH control, potassium permanganate and pre-aeration. The recovery and selectivity were optimised by varying the collector and depressant concentration, pH and the effect of copper sulphate addition. It was found that with MAA, depression of arsenopyrite was possible however, the recovery of pyrite was also adversely affected. To achieve the best separation, the pulp conditions were a pH of 8, 2.14 × 10⁻³ mol/l sodium ethyl xanthate and 250 mg/l of MAA depressant. These conditions resulted in an overall arsenopyrite recovery of 25.5% and a pyrite recovery of 62.1%, a decrease in recovery of 63.0% and 7.5%, respectively.     

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.