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.     

Electro-remediation of copper mine tailings. Comparing copper removal efficiencies for two tailings of different age

This work compares and evaluates the copper removal efficiency when applying electric fields to two mine tailings originating from the same mine but of different age. Eight experiments were carried out – four on tailings deposited more than 20 years ago (old tailings) and four on tailings deposited less than 2 years ago (new tailings). Parameters analyzed were the applied voltage drop, acid concentration during pretreatment, and the use of either passive or ion exchange membranes in the experimental setup.The comparison of the results confirms that there are differences in the electroremediation between the two tailings, even if the pH is similar and a mineralogical analysis showed similarities between the samples with respect to composition. It was found that an electroremediation is more favorable on the old tailings. The results showed that the best experimental conditions for both tailings is a pretreatment with H₂SO₄ 1 M followed by applying 40 V for 7 days, using ion exchange membranes. In this case 16.7% of copper was removed from the anode section for the old tailings, whereas only 11.2% was removed from the new tailings. The current efficiencies with respect to copper for the old and new tailings were 1.7% and 1.6%, respectively.     

Extraction of titanium (IV) from acidic media by tri-n-butyl phosphate in kerosene

The extraction of titanium (IV) from sulfate, and nitrate solutions has been studied using tri-n-butyl phosphate (TBP) in kerosene. Extraction of titanium was affected by acid concentration over the range of 0.5–4 mol L^?1. The titanium distribution coefficient reached a minimum between 1 and 2 mol L^?1 acid for both sulfate and nitrate solutions. Third phase formation was observed in the extraction of titanium from acidic media at all condition tested. At the next stage, the stripping of titanium was studied using H₂SO₄, H₂SO₄ + H₂O₂ and Na₂CO₃. The kinetics of the stripping were very slow for H₂SO₄. The use of complex forming stripping agents (H₂SO₄ + H₂O₂) and Na₂CO₃ significantly improved the kinetics of stripping. About 98% recovery was achieved by extracting titanium from an aqueous nitrate solution using TBP and stripping with sodium carbonate.     

Effect of thiocyanate on BIOX® organisms: Inhibition and adaptation

Stringent environmental legislation and the desire to become zero discharge have motivated mining operations to treat and recycle process water. Cyanidation tailings effluent contains elevated concentrations of cyanide and thiocyanate (SCN⁻), precluding recycling to the BIOX® process without prior treatment to reduce SCN⁻ to below 1 mg/l. The current study investigated the effect of SCN⁻ on individual microbial species. Iron oxidation by Leptospirillum ferriphilum was not affected by SCN⁻ concentrations below 0.5 mg/l, with concentration dependent inhibition observed between 0.75 and 1.25 mg/l and complete inhibition of iron oxidation above 1.25 mg/l. Sulphur oxidation by Acidithiobacillus caldus showed a similar trend, with limited inhibition below 1.25 mg/l and almost complete inhibition above 1.25 mg/l. Repeated sub-culturing at low concentrations induced adaptation, with adapted cultures currently growing at SCN⁻ concentrations of 7 mg/l. The phenomenon of inhibition at low concentration, with subsequent adaptation was repeated in stirred tank reactors, leaching a pyrite/arsenopyrite concentrate.     

Effect of fluidized-bed carrier material on biological ferric sulphate generation

High biomass hold-up and high iron oxidation rates of a biological ferric sulphate generating fluidized-bed reactor (FBR) requires a carrier material with high specific surface area, high porosity and inertness. In this work, the effect of activated carbon (AC), diatomaceous earth (Celite) and Al₂O₃ (Compalox) carrier materials on the ferric sulphate generation in FBRs were studied. Compalox dissolved during the experiments and formed an unfluidizable aggregate, and was therefore rejected. The slow dissolution of Celite resulted in a light, fine-grained, layer on top of the fluidized bed that had to be changed into fresh Celite. AC resisted well the friction caused by fluidization. The iron oxidation in the continuous-flow FBRs became limited by oxygen supply already at loading rates of 2.5 kg Fe²⁺ m⁻³ h⁻¹. Iron oxidation rates of 27.6 and 25.7 kg m⁻³ h⁻¹ were obtained in batch FBR experiments with AC and Celite, respectively.Biomass accumulation of 6.2 × 10¹⁰, 2.4 × 10¹⁰ and 8.0 × 10⁹ cells per g of carrier was detected on Celite, AC and Compalox, respectively. The bacterial community structures on the carrier materials were revealed by Polymerase Chain Reaction and Denaturating Gradient Gel Electrophoresis (PCR-DGGE) followed by partial sequencing of the 16S rRNA gene. Two bacterial strains, Leptospirillum ferriphilum and a strain similar to a strain unofficially named “Ferrimicrobium acidiphilum”, were detected. Examination of the carrier material surfaces with scanning electron microscopy (SEM) coupled with energy dispersive spectroscopy (EDS) revealed that all carrier materials were covered with jarosite precipitates and that the bacteria were mainly retained on the jarosite covered areas. In conclusion, AC was the most promising carrier material for a large-scale biological ferric sulphate generating FBR based on its availability, durability and the achieved high iron oxidation rates.     

An experimental investigation into the oxidation of four pyritic shales from Western Australia

The oxidation of pyritic shale under different experimental conditions was studied using an isothermal batch reactor technique. Four pyritic shale samples collected from different areas of an iron ore mine with different stratigraphic compositions in Western Australia were employed in this study. The influence of shale properties on the oxidation of pyritic shale was studied. It was found that the reaction rate constant k (L kg^?1 h^?1) of the shale oxidation as measured with the present isothermal reactor technique on a per unit pyrite mass basis, was not constant for the different shales and was dependent on their physical properties, where increasing grain size and increasing encapsulation of pyrite grains both result in decreased surface exposure of pyrite to oxidation (per unit mass) which coincides with a decrease in the oxidation rate of pyritic shale (per unit mass). This study demonstrates that pyrite grain size and encapsulation are important parameters for the interpretation and evaluation of acid mine drainage potential associated with individual shales.     

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.     

Beneficiation of colemanite tailings by attrition and flotation

Boron minerals are generally concentrated using attrition methods followed by screening and classification to remove clay minerals in industrial scale. Physical concentration methods are used in Kestelek Boron Mine for the concentration of colemanite. Because of the inefficient process operation, the tailings containing about 20% B₂O₃ are discarded into the tailings pond. In this study, colemanite tailings sample taken from tailings pond was treated using scrubbing + screening followed by flotation to recover the lost boron. As a result of the experimental studies, a concentrate containing 44.5% B₂O₃ was produced with 68.4% B₂O₃ recovery.     

A novel approach for recovery of rare earths and niobium from Bayan Obo tailings

It is difficult to economically recover rare earths (RE) and niobium (Nb) from Bayan Obo tailings by the existing metallurgical processes. In this study, a novel hydrometallurgy process was employed for separating and recovering RE and Nb from Bayan Obo tailings. Firstly, by sulfating roasting at 250 °C and subsequent leaching at 60 °C, the RE and Nb present in the polymetallic minerals can be efficiently extracted into the leach solution. Secondly, after the reduction of Ti⁴⁺ and Fe³⁺ ions (to Ti³⁺ and Fe²⁺ ions) with iron powders followed by hydrolysis at pH 2.01, the Nb can be efficiently precipitated from the leach solution. The impurities present in the precipitated product can then be removed by treating with NH₃⋅H₂O–H₂C₂O₄ system at pH 4.50. Thirdly, the RE can be efficiently precipitated at pH 7.15 from the filtrate of above hydrolysis reaction mixture. Finally, the impurities present in the crude RE can be removed by oxalate co-precipitation method. The yield of RE and Nb in this novel process reaches up to 90% and 78%, respectively. Both the Nb (60.67 wt% Nb₂O₅) and RE products (>88.65 wt% RE_xO_y) have high application value.     

Biological treatment of wastewater produced during recycling of spent lithium primary battery

Wastewater produced during recycling of spent lithium primary battery was biologically treated with Acidithiobacillus ferrooxidans to decrease the pH and metal concentration. Since the wastewater contains high concentrations of Cr, Ni, and Li, the effects of these metals on the bacterial activity in a 9 K medium were also investigated. Samples of the medium with different metal concentrations were treated, and the oxidation ratio of Fe²⁺ ions was measured to examine the activity of bacteria. In the treatment of simulated wastewater, the presence of Cr and Ni ions with concentrations of 8000 g m^?3 and 13,000 g m^?3, respectively, did not inhibit the bacterial activity, whereas the oxidation ratio of Fe²⁺ ions was observed to be low in the medium when Li ion was present with a concentration at 5000 g m^?3. This observation suggested that at this concentration, Li ion suppressed the bacterial activity. In the case of treatment of real wastewater containing Cr, Ni, and Li, the oxidation ratio of Fe²⁺ to Fe³⁺ was observed to be low while the Fe concentration and pH decreased to 21,633 g m^?3 and 1.8, respectively. Thus, the wastewater produced during the recycling of spent lithium primary batteries can be effectively treated biologically for re-circulating in the recycling process.     

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.     

Solvent extraction and separation of palladium(II) and platinum(IV) from hydrochloric acid medium with dibutyl sulfoxide

The solvent extraction and separation performances of Pd(II) and Pt(IV) from hydrochloric acid solutions were investigated using dibutyl sulfoxide (DBSO) diluted in kerosene. Pd(II) was strongly extracted by a lower concentration DBSO in a lower concentration hydrochloric acid solution while the reverse was obtained for Pt(IV) extraction. Based on independent extraction and separation experiments of Pd(II) and Pt(IV), the separation parameters of Pd(II) and Pt(IV), including dibutyl sulfoxide concentration, contact time of aqueous and organic phases, organic/aqueous (O/A) phase ratio and H⁺ concentration of aqueous phase, were studied in detail, and the optimal separation parameters were obtained and summarized as the following: dibutyl sulfoxide concentration 0.6–1.2 mol dm^?3, organic/aqueous (O/A) phase ratio 0.6–1.0, H⁺ concentration of aqueous phase 1.0–1.5 mol dm^?3 and contact time of two phases 5 min. The as-prepared separation parameters were corroborated by the extraction and separation from a synthetic stock solution containing Pd(II), Pt(IV) as well as several common impurities like Fe(II), Cu(II) and Ni(II). The results revealed that Pd(II) could be separated efficiently from Pt(IV) with a high separation coefficient of Pd(II) an Pt(IV) (2.7 × 10⁴) by predominantly controlling dibutyl sulfoxide and hydrochloric acid concentrations. The extraction saturation capacity of Pd(II) was determined from 1.0 mol dm^?3 HCl solution with 3 mol dm^?3 dibutyl sulfoxide and its experimental value exceeded 14 g dm^?3 under the experimental conditions.Stripping of Pd(II) from loaded organic phase was performed using a mixed aqueous solution containing NH₄Cl and ammonia solutes. Pd(II) (99.2%) was stripped using the stripping solution containing 3% (m/v) NH₄Cl and 5 mol dm^?3 ammonia, respectively.     

Bioleaching of phosphorus from rock phosphate containing pyrites by Acidithiobacillus ferrooxidans

Pyrites can be oxidized by the bacterium Acidithiobacillus ferrooxidans (At. f.), producing H₂SO₄ and FeSO₄. Rock phosphate is dissolved by H₂SO₄, forming soluble phosphorus. Fe²⁺ in FeSO₄ is oxidized to Fe³⁺, producing energy to sustain the growth of At. f. The effects of four factors (rock phosphate dosage, pyrite dosage, culture temperature and time) on the fraction of phosphorous leached were investigated. It is suggested that the optimal conditions are as follows: rock phosphate dosage 1 g/L, pyrite dosage 30 g/L, culture temperature 30 °C, culture time 84 h. The fraction of phosphorous leached is up to 11.8%.     

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.     

The effect of particle size on acid mine drainage generation: Kinetic column tests

The rate of acid mine drainage (AMD) generation is directly proportional to the surface area and so to the particle size distribution of acid-forming minerals exposed to oxidation. Materials in various particle sizes are subject to weathering processes at field condition; however, the particle size dependent oxidation rate has not been investigated for understanding entire geochemical behavior at a mining site. Therefore, a comprehensive research program was aimed to investigate the effect of particle size on pH variation and acid mine drainage generation using kinetic column tests, and then to find convenient methodologies for upscaling laboratory-based results to the field condition. For this purpose, ore samples collected from Murgul Damar open-pit mining were grinded in three different particle size distributions that are coarse (minus 22.5 mm), medium (minus 3.35 mm) and fine (minus 0.625 mm) sizes, 34 columns were designed in different dimensions for kinetic column tests. It was found that the cumulative concentration of the many constituents measured from medium particles (minus 3.35 mm) are higher than coarser samples due to decreasing specific surface area with increasing particle size. Similarly, because of decreasing of hydraulic conductivity with increasing the fine content, the cumulative concentration of constituents measured from medium particles (minus 3.35 mm) are also higher than finer particles (minus 0.625 mm). Based on statistical and analytical analyses of the results of kinetic column tests, the time required to initiate acid formation at field condition varied between 489 and 1002 days depending on particle size distribution. In addition, considering the effect of particle size and the results of related statistical analysis, main oxidation (SO₄²⁻) and neutralization (Ca²⁺, Mg²⁺, Mn²⁺ etc.) products were also successfully upscaled to the field condition.     

Formulating and optimising the compressive strength of controlled low-strength materials containing mine tailings by mixture design and response surface methods

Controlled low-strength materials (CLSM), like other cement-based backfill materials, are typically formulated by trial-and-error methods to yield the desired product characteristics. This paper presents the use of mixture design and response surface methods as tools to optimise formulations of CLSM to achieve desirable mechanical integrity with a minimum amount of statistically-sound experiments; while minimising the amount of cement and maximising the amount of by-products used. Statistical combinations of three-component mixtures were formulated to investigate the unconfined compressive strength (UCS) of CLSM comprising: Portland cement, fly ash and mine flotation tailings from a Ni–Cu ore. The data is analysed using the response surface method (using a mixture design of a constrained triangular surface) and ANOVA. Optimum formulations are simulated using a desirability function set at lower (1.0 MPa), target (2.0 MPa) and upper (3.0 MPa) UCS values after 28 days curing. All mix combinations had a constant spread diameter of 229 ± 10 mm, the standard workability for conventional CLSM. Results are compared to conventional CLSM incorporating silica sand in the place of the tailings. A significant quantity of tailings (up to 80 wt% solids) and low quantity of cement (up to 5 wt% solids) produced CLSM with UCS within the 2 MPa target value of excavatability. UCS of CLSM is a function of the individual component proportions, and the mixture design approach can be an important tool to help develop and optimise formulations of cement-based materials consisting of several components.     

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.     

Paste backfill of high-sulphide mill tailings using alkali-activated blast furnace slag: Effect of activator nature,concentration and slag properties

The effect of activator type, concentration and slag composition on the strength and stability properties of paste backfill (CPB) of high-sulphide tailings using alkali-activated slag (AAS) as binder (7 wt.%) were investigated in this study. Acidic and neutral (AS–NS) slags were activated with liquid sodium silicate (LSS) and sodium hydroxide (SH) at 6–10 wt.% concentrations. Ordinary Portland cement (OPC) results were used for comparison. The strength development was found to remarkably improve with increasing the concentration from 6 to 8 wt.%. Further increase in concentration did not enhance the strength. SH was determined to produce higher early-age strength whilst LSS produced higher long-term strengths as an indication of slag selectivity for activators. More extensive gypsum formation was observed at lower concentrations in SEM/EDS studies. An increase in Na₂O concentration raised the activator consumption. High concentrations also led to poorly crystallized C–S–H gel, loose structure and drying shrinkage cracks especially in NS–SH samples. A reduction in total porosity up to 20% was obtained in AAS samples compared to OPC. Amorphous structure, chemical modulus ratio and/or basicity index (BI) values were seen to control the pozzolanic reactivity, and therefore, the alkali-activation and hardening process.     

Kinetics of saprolitic laterite leaching by sulphuric acid at atmospheric pressure

Mineralogical analyses of the saprolitic laterite material have been characterized by X-ray fluorescence (XRF), X-ray diffraction (XRD), Fourier transform infrared (FTIR) spectroscopy, thermal analysis, scanning electron microscope (SEM) and energy dispersive X-ray analysis (EDAX). Results showed that the saprolitic laterite material consists mainly of nickel-substituted lizardite showing the pebble-like morphology and traces of magnetite and phlogopite. Leaching results showed that as much as 84.8% nickel could be leached under the experimental conditions of 10% (v/v) H₂SO₄, 90 °C reaction temperature, leached within 5 min, particle size d₅₀ = 25 μm, stirring at 500 rpm and liquid to solid ratio 3:1. The kinetics of nickel and magnesium leaching from the saprolitic laterite material have been investigated in a mechanically stirred reactor and the activation energies were determined to be 53.9 kJ mol^?1 for nickel and 59.4 kJ mol^?1 for magnesium respectively, which are characteristic for a chemical reaction controlled process. The similarity of the activation energies of nickel and magnesium leaching from the saprolitic laterite material by sulphuric acid means that nickel in lizardite is loosely bound within the octahedral layer and almost all of the nickel could be leached simultaneously with magnesium but without complete decomposition of the silicate structure.     

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.