A novel low pH sulfidogenic bioreactor using activated sludge as carbon source to treat acid mine drainage (AMD) and recovery metal sulfides: Pilot scale study

In this work a pilot scale sulfidogenic bioreactor was used to treat acid mine drainage (AMD) from Zijinshang copper mine. In this process, S²⁻ produced in the Up-flow Anaerobic Sludge Bed (UASB) reactor were recycled in the two precipitation tanks for copper and iron precipitation, activated sludge from local waste water treatment plant was used as the carbon source. The reactor were steady operated in acid condition (with no pH control) for 4 month, AMD with a copper concentration of 100–120 mg/L, iron concentration of 170–200 mg/L, sulfate concentration of 2000–2500 mg/L and pH of 2.34–2.56, were feeding into the reactor under a feed rate of 1 m³/days and HRT of 3 days, copper and iron removal were 60.95%, 97.83% respectively. The precipitant in the precipitation tank containing 15.7% Cu and 22.66% Fe, thus indicating a recovery possibility of copper by pyrometallurgy process. From these results we can conclude that an SRB process would be a viable method of treating Zijinshan AMD.     

The application of wood ash as a reagent in acid mine drainage treatment

The paper deals with a possible utilisation of wood ash as a reagent in treating acid mine drainage (AMD) from opencast mining of brown coal. Wood ash samples were obtained having combusted deciduous and coniferous tree wood in a household furnace. The dominant mineral phases in wood ash are calcite, quartz, lime and periclase. The used AMD is characteristic of high contents of sulphates, iron, manganese, heavy metals and low pH. The AMD treatment process included dosing of wood ash to adjust pH values about 8.3 (a dose of 0.5 g l⁻¹) or calcium hydroxide (a dose of 0.2 g l⁻¹) for comparison. The reaction time was 20 min. Dosing of wood ash in AMD resulted in an increase of pH in solution from 3.5 to 8.3, which caused the removal of metal ions mainly by precipitation, co-precipitation and adsorption. Comparing the application of Ca(OH)₂ in AMD treatment, at an almost identical pH value the concentrations fell in both cases for Fe, Mn, As, Co, Cu, Ni, Zn, Mg, Al and Mo. Applying wood ash the drop was even more distinct in Mn, Zn and Mg. The results of sedimentation tests in an Imhoff cone confirm that the settling capacities of sludge using wood ash are significantly better than when using calcium hydroxide in acid mine drainage treatment.     

Enhancement of Sludge Sedimentation Properties in a Concentrated Acid Mine Drainage Using Nano- and Micro-magnetite Particles

Conventional treatment of AMD involves neutralization with consequent precipitation of metals as hydroxides. In AMD with a high concentration of metals, the settling rate of the sludge/water interface is low. We investigated the use of nano- and micro-magnetite particles to assist the settling and thickening of floc particles. The magnetite was produced from ferrous sulphate crystals (melanterite, Fe₂SO₄·7H₂O) obtained by leaching pyrite from a coal mine. AMD was obtained from the treatment plant at the same mine and the water was neutralized with Ca(OH)₂ at pH 8.7?±?0.1. Laboratory studies were conducted in 1 L test tubes with and without the addition of magnetite particles and a flocculant. Sedimentation curves (interface settling) were generated to evaluate the rate of sedimentation. For the studied effluent, the best option was 4 g L^?1 of magnetite particles and 5 mg L^?1 of high molecular weight anionic polyacrylamide. The magnetite particles were recovered magnetically from the sludge with ≈ 90% efficiency. Thus, the combined use of magnetite and a flocculant increased the sludge settling rate and, consequently, reduced the area needed for settling basins.

     

Comparing the effect of salts and frother (MIBC) on gas dispersion and froth properties

The Raglan concentrator (Xstrata Nickel) does not employ frother. It was considered this might be the result of the high salt content in the process water (ca. 30 000 ppm). Two-phase (solution–air) and three-phase (slurry–air) tests were undertaken in a laboratory column to quantify the effect of inorganic ions present in the water (a range of polyvalent ions). The measurements focused on gas dispersion (bubble size and gas holdup) and froth overflow rate. The results were compared to a typical frother (MIBC) system. The two-phase tests revealed reduced bubble size, increased gas holdup and limited froth formation in salt solutions. The gas holdup correlated with ionic strength. At an ionic strength ca. 0.4 (=0.4 M NaCl) the increase in gas holdup was comparable to ca.10 ppm MIBC. In three-phase tests on a sulphide ore, bubble size and froth overflow rate were again comparable between 0.4 M NaCl and 10 ppm MIBC. The observations help explain why the Raglan plant can operate without frother addition.     

Recovery of boron from borax sludge of boron industry

A two-step process for boron recovery from borax sludge is proposed in the present work. The borax sludge was leached with sulphuric acid solution. Then, for the removal of alkaline species from the leachate, calcium and magnesium were precipitated by adjusting the pH of leachate using 1.5 M NaOH and 1.5 M HCl solutions. The effects of pH, temperature, concentration and time on the precipitation process were investigated. It was determined that the calcium and magnesium concentrations in the leachate were reduced from 121 mg/L to 2.6 mg/L and from 145 mg/L to 4.0 mg/L, respectively, at a pH value of 12, a temperature of 70 °C, an initial boron concentration of 500 mg/L and a precipitation time of 3 h. Under these optimum conditions, it was observed that the boron concentration in the solution decreased very slightly. In this process, the alkaline species were successfully separated from the boron.Finally, borax pentahydrate (Na₂B₄O₇·5H₂O) was produced by the evaporation of the final solution obtained after precipitation process.     

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.     

Extraction of indium from zinc plant residues

The main purpose of this study was to extract indium from the Irankoh zinc plant residue. The Irankoh zinc plant residue contained 145 ppm indium. The optimum conditions for leaching of indium and reduction of ferric ion in reductive leaching were obtained at temperature of 90 °C for a leaching duration of 3 h with sulfuric acid concentration of 100 g/L and the amount of required sodium sulfide for reduction of ferric was 1.5 times of stoichiometric quantity of iron. Then, to prepare concentrated indium solution, indium was selectively precipitated from the leach solution. The pH of leach solution was adjusted to 6 with ammonia solution in 90 °C for selective indium precipitation, and reaction time was considered to be 10 min. Then the resulting precipitation was dissolved using hot sulfuric acid solution, and the solution was subject to solvent extraction and cementation using zinc powder to recover indium.     

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.     

Determination of the concentrations of calcium and magnesium released from fluid inclusions of sphalerite and quartz

This study investigates the morphology of fluid inclusions trapped within a natural sphalerite and its closely intergrown quartz, and determines the total concentrations of calcium (Ca) and magnesium (Mg) released from the inclusions after grinding. The results indicated that numerous fluid inclusions with sizes from few microns to dozens of microns exist in the sphalerite and quartz. The inclusions contain abundant Ca and Mg. However, the concentrations of Ca and Mg in the inclusions of quartz are significantly lower than those of sphalerite. The inclusions were broken during the grinding process, releasing Ca and Mg to the solution; these released concentrations increased with the increase of grinding time. The maximum concentrations of Ca and Mg released from the fluid inclusions were 61.19 ppm and 5.23 ppm for sphalerite, and 3.28 ppm and 0.21 ppm for quartz, respectively. This study provides new understanding for the source of Ca and Mg in flotation pulp.     

TGA kinetic study on the hydrogen reduction of an iron nickel oxide

Reduction of a mixed iron nickel oxide by hydrogen to produce ferronickel alloy was investigated by thermo-gravimetric analysis (TGA). The iron nickel oxide, which contains 50 wt% Fe and 10 wt% Ni mainly in the form of hematite (Fe₂O₃) and nickel ferrite (NiFe₂O₄), is a residue produced from the water leaching of a selectively sulfation roasted nickel calcine. Continuous heating tests with a fixed heating rate as well as isothermal tests were performed. Results indicate that the reduction of the mixed oxide began above 350 °C. The reduction reactions occurred in a non-topochemical mode at low temperatures between 350 °C and 600 °C with its rate controlled by the solid–gas chemical reactions. The reduction rate increased with the increase in temperature from 350 °C up to 1100 °C. Between 600 °C and 1200 °C, the rate controlling step was the diffusion of reducing gas through the pores of the sample bed with an apparent activation energy of 34.1 kJ/mol. Due to the melting of the silicate material at 1200 °C which substantially reduced the sample porosity, the reduction rate decreased. Above 1200 °C, the reduction rate increased again due to the increased diffusion of the reducing gas through the molten sample at higher temperatures.     

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

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

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.     

Biosorptive removal of Cd and Zn from liquid streams with a Rhodococcus opacus strain

The biosorption abilities of Rhodococcus opacus were studied for cadmium and zinc removal for liquid aqueous streams. The influence of pH, initial metal concentration and time removal were evaluated on the biosorption studies, in a batch scale basis. The Cd²⁺ and Zn²⁺ species uptake capacity by R. opacus has been also compared using Langmuir and Freundlich models. At pH 7.0 and 26 °C C, Cd²⁺ removal achieved a value around 60% from an initial concentration of 15 ppm. On the other hand, Zn removal achieved a value around 88% from an initial concentration of 5 ppm. Kinetics studies revealed that the biosorption process followed a pseudo-second order model for the two metal species (Cd²⁺ and Zn²⁺) and the kinetic constants were 3.90 and 3.37 g mg⁻¹ min⁻¹, for an initial concentration of 15 and 5 ppm for cadmium and zinc, respectively. The results showed that the R. opacus is a potential engineering biosorbent for environmental and extractive metallurgy sustainable applications.     

Biosorption of arsenic(V) with Lessonia nigrescens

Conventional treatment methods for arsenic removal from copper smelting wastewaters create sludge that is difficult to handle. Biosorption of arsenic using algae as sorbent is an interesting alternative to the conventional methods.This work shows results from biosorption of arsenic(V) by Lessonia nigrescens at pH = 2.5, 4.5 and 6.5. The adsorption of arsenic could be explained satisfactorily both by the Freundlich and the Langmuir isotherms. Maximum adsorption capacities were estimated to 45.2 mg/g (pH = 2.5), 33.3 mg/g (pH = 4.5), and 28.2 mg/g (pH = 6.5) indicating better adsorption at the lower pH. These values are high in comparison with other arsenic adsorbents reported.The sorption kinetics of arsenic by L. nigrescens could be modelled well by Lagergren’s first-order rate equation. The kinetics were observed to be independent of pH during the first 120 min of adsorption with the Lagergren first-order rate constant of around 1.07 × 10⁻³ min⁻¹.     

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).     

Froth flotation of sphalerite: Collector concentration,gas dispersion and particle size effects

This experimental work on sphalerite flotation investigated the effect on flotation performance of three particle size fractions, namely, coarse (d₈₀ = 100 μm), medium (d₈₀ = 39 μm) and fine (d₈₀ = 15 μm), bubble size distribution, superficial air velocity, and collector dosage. Bubble size distributions were characterized with the image analysis technique. The two-phase (liquid–gas) centrifugal pump and frother addition (MIBC, 5–30 ppm) allowed generating bubble diameters between 150 and 1050 μm, and air holdup ranging from 0.2% and 1.3%. Main results showed that each particle-size distribution required an optimal bubble-size profile, and that sphalerite recovery proceeded from mechanisms involving true flotation (when J_g = 0.04 cm/s and 1.9 × 10⁻⁴ M SIPX). However, cluster-flotation occurs at high collector dosage (when J_g = 0.04 cm/s and d₃₂ between 285 and 1030 μm), and requiring further investigation.     

Effect of gas rate and impeller speed on bubble size in frother-electrolyte solutions

In a flotation cell, bubble size is a function of both coalescence and breakup phenomena. Two phase tests, conducted in a conventional 5.5 L Denver mechanical flotation cell, studied the effect of impeller speed, gas flow rate and frother concentration on bubble size in various electrolyte-frother solutions. The addition of frother to a synthetic sea salt did reduce the measured bubble size (at certain mechanical conditions); whereas the effect of frother addition to NaCl was too small (when compared to measurement errors) to make significant conclusions. This led to more detailed CCC curves (0–50 ppm MIBC) for NaCl, NaCl + MgCl₂, NaCl + CaSO₄, and NaCl + KCl solutions, at constant electrolyte concentrations, to be conducted. They showed an increase in bubble size with the addition of MIBC. This was attributed to the saturation of frother at the air-water interface, reducing local surface tension gradients that help produce smaller bubbles. This occurrence is typically masked in traditional CCC curves due to the dominance of coalescence effects at low frother concentrations.     

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

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

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.