Arsenic,Selenium, and Sulfate Removal using an Ethanol-Enhanced Sulfate-Reducing Bioreactor

A laboratory scale sulfate-reducing bioreactor column study was used to investigate removal of arsenic, selenium, and sulfate in neutral to alkaline simulated mine water. Selenium was effectively removed as insoluble elemental selenium from 200 μg/L to below the maximum contaminant level (50 μg/L). With the addition of ferrous chloride in the influent, arsenic was also effectively removed to near the maximum contaminant level (10 μg/L). Sulfate removal was apparently limited by the presence of undissociated hydrogen sulfide, which was found to limit sulfate-reducing bacterial activity in the range of 20–40 mg/L and will need to be considered for sulfate-reducing bacteria treatment system designs.     

Lime Treatment of Mine Drainage at the Sarcheshmeh Porphyry Copper Mine,Iran

Laboratory and field treatment tests were performed to evaluate the effectiveness of lime treatment for mitigation of environmental effects of acid mine drainage (AMD) at the Sarcheshmeh porphyry copper mine. AMD associated with the rock waste dumps is contaminated with Al (>36,215 μg/L), Cd (>105 μg/L), Co (>522 μg/L), Cu (>53,250 μg/L), Mn (>42,365 μg/L), Ni (>629 μg/L), and Zn (>12,470 μg/L). The concentrations of other metals (Fe, Mo, Pb, and Se) are low or below detection limits (As, Cr, and Sb). Due to the very high Al and Mn content and the low concentration of Fe, a two-stage lime treatment method was chosen for the laboratory tests. In the first stage, the AMD was treated at four pH set points: 7.5, 8.9, 9, and 10. In the second stage, after removing the sludge at pH 9, treatment was continued at pH 10 and 11. The results indicated that a two-stage treatment method was not necessary because elements such as Al, Cu, Co, and Zn were easily treated at pH 7.5, while complete removal of Cd, Mn, and Ni only required a pH of 10. Increasing pH during the treatment process only caused a slight increase in Al. Field treatment tests support the laboratory results. Lime treatment of highly contaminated AMD from dump 11, using simple low density sludge pilot scale equipment, show that contaminant metals are treatable using this method. The mean treatment efficiency for contaminant metals was 99.4% for Al, % for Cd, 99.6% for Co, 99.7% for Cu, 98.5% for Mn, 99.7% for Ni, 99% for U, and 99.5% for Zn. The optimum pH for AMD treatment by lime was in the range of 9–10. The produced sludge in the treatment process was highly enriched in the contaminant metals, especially Cu (>7.34%), Al (>4.76%), Mn (>2.94%), and Zn (>1.25%). A correlation coefficient matrix indicates that the distribution pattern of the contaminant metals between soluble and precipitated phases is consistent with the hydrochemical behavior of the metals during the lime treatment process.     

Influence of Past Mining on the Quality of Surface Waters at Funtana Raminosa (Sardinia)

Cu–Pb deposits at Funtana Raminosa in Central Sardinia were intensively exploited, mostly underground, from 1917 until 1983. Flotation tailings were dumped near the mine plant. A hydrogeochemical survey carried out in 2004 showed that mine drainage collected from several galleries was circumneutral, due to the availability of carbonate minerals that buffer the acidity produced by the oxidation of Fe-bearing sulphides. The mine waters contained higher concentrations of dissolved SO₄, F, Zn, Cd, Pb, Mn, and Mo than was observed in uncontaminated spring and stream waters in the area. Drainage from the oldest flotation tailings showed much lower concentrations of Zn, Cd, and Pb than those generally observed in mine waters. In contrast, drainage from the recent flotation tailings had the highest levels of dissolved SO₄, Zn, and Cd (1,600, 30, and 0.8 mg/L, respectively) when sampled in the dry season; these were two orders of magnitude lower in the rainy season under high flow condition. Pb was ≈ 5 μg/L under different flow conditions. Water in the Rio Saraxinus, a stream that drains the entire mining area, had a relatively low level of contamination (170 μg/L Zn, 7 μg/L Cd, and 0.9 μg/L Pb).     

添加剂对双膜三室锰电积清洁节能工艺效果的影响

采用双膜三室电解工艺,研究了添加剂对电沉积锰效果的影响。分别选用两种添加剂亚硒酸、聚丙烯酰胺 硫脲作为无机含硒添加剂和有机复合添加剂的典型代表,结果表明：无机添加剂亚硒酸最佳浓度为0.2 g/L时,电流效率可以达到74.2﹪,酸回收率为37.3﹪,电耗为6648 kW·h/t;有机复合添加剂聚丙烯酰胺浓度0.01 g/L、硫脲浓度0.015 g/L时,电流效率达到了71.1 ﹪,酸回收率为35.1 ﹪,电耗为6990 kW·h/t,此时表面形貌效果达到最佳,表面光亮致密。     

An Evaluation of Metal Contamination in Surface and Groundwater around a Proposed Uranium Mining Site, Jharkhand, India

The East Singhbhum region is a highly mineralised zone, with extensive mining of copper, uranium, and other minerals. The concentrations of certain metals (Fe, Mn, Zn, Pb, Cu, and Ni) were measured in 10 groundwater locations and eight surface water locations for four seasons during 1 year around a proposed uranium mining area. The ranges of Fe, Mn, Zn, Pb, Cu, and Ni in surface water were 0.08–1.21, 0.02–0.32, 0.02–3.48 mg/L, 0.84–14, 1.25–36, and 1.24–15 μg/L, respectively, while in groundwater, the ranges were 0.06–5.3, 0.01–1.3, 0.02–8.2 mg/L, 1.4–28, 0.78–20, and 1.05–20 μg/L, respectively. Only Fe and Mn were found to exceed India’s drinking water standards. The data have been used to calculate a metal pollution index (MPI). The MPI of both groundwater (28) and surface water (10) is well below the index limit of 100, which suggest that neither is generally contaminated with respect to these metals.     

Iron Removal by a Passive System Treating Alkaline Coal Mine Drainage

The Marchand passive treatment system was constructed in 2006 for a 6,000 L/min discharge from an abandoned underground bituminous coal mine located in western Pennsylvania, USA. The system consists of six serially connected ponds followed by a large constructed wetland. Treatment performance was monitored between December 2006 and 2007. The system inflow was alkaline with pH 6.2, 337 mg/L CaCO₃ alkalinity, 74 mg/L Fe, 1 mg/L Mn, and <1 mg/L Al. The final discharge averaged pH 7.5, 214 mg/L CaCO₃ alkalinity, and 0.8 mg/L Fe. The settling ponds removed 84% of the Fe at an average rate of 26 g Fe m⁻² day⁻¹. The constructed wetland removed residual Fe at a rate of 4 g Fe m⁻² day⁻¹. Analyses of dissolved and particulate Fe fractions indicated that Fe removal was limited in the ponds by the rate of iron oxidation and in the wetland by the rate of particulate iron settling. The treatment effectiveness of the system did not substantially degrade during cold weather or at high flows. The system cost $1.3 million (2006) or $207 (US) per L/min of average flow. Annual maintenance and sampling costs are projected at $10,000 per year. The 25-year present value cost estimate (4% discount rate) is $1.45 million or $0.018 per 1,000 L of treated flow.     

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%.     

Underground Mine Water for Heating and Cooling using Geothermal Heat Pump Systems

Abstract  In many regions of the world, flooded mines are a potentially cost-effective option for heating and cooling using geothermal heat pump systems. For example, a single coal seam in Pennsylvania, West Virginia, and Ohio contains 5.1 x 10¹² L of water. The growing volume of water discharging from this one coal seam totals 380,000 L/min, which could theoretically heat and cool 20,000 homes. Using the water stored in the mines would conservatively extend this option to an order of magnitude more sites. Based on current energy prices, geothermal heat pump systems using mine water could reduce annual costs for heating by 67% and cooling by 50% over conventional methods (natural gas or heating oil and standard air conditioning).     

Synchrotron X-ray photoelectron spectroscopic study of the chalcopyrite leached by moderate thermophiles and mesophiles

SXPS (Synchrotron X-ray Photoelectron Spectroscopy) and NEXAFS (Near Edge X-ray Absorption Fine Structures) have been applied to study the surface chemical species of chalcopyrite leached by a moderate thermophilic consortia, Leptospirillum ferrooxidans and a mesophilic mixed culture of L. ferrooxidans, Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans. A sulfur-rich layer dominated by S_n²⁻ developed with time, which was found to control the rate of bioleaching. Fe L_2,3-edge NEXAFS and Fe 2p spectra indicate the formation of jarosite during bioleaching. Thermophiles significantly enhanced the leaching efficiency, in which 1.34 g/L copper was dissolved in 25 days, while less than 0.3 g/L copper was released in 30 °C bioleaching. This was mostly caused by the increased abiotic reaction rate. The solution copper concentration in presence of L. ferrooxidans was higher than that with mesophilic mixed culture, which suggests the synergistic effect of mixed microorganisms did not play a comparably important role as temperature under the conditions used in this study. Explicit evidence of elemental sulfur was only found in samples leached by L. ferrooxidans by Raman spectroscopy. However, the formation of elemental sulfur does not significantly hinder the leach rate.     

Recovery of zinc and cadmium from industrial waste by leaching/cementation

Metallic zinc production from sulfide zinc ore is comprised by the stages of ore concentration, roasting, leaching, liquor purification, electrolysis and melting. During the leaching stage with sulfuric acid, other metals present in the ore in addition to zinc are also leached. The sulfuric liquor obtained in the leaching step is purified through impurities cementation. This step produces a residue with a high content of zinc, cadmium and copper, in addition to lead, cobalt and nickel. This paper describes the study of selective dissolution of zinc and cadmium present in the residue, followed by the segregation of those metals by cementation. The actual sulfuric solution, depleted from the electrolysis stage of metallic zinc production, was used as leaching agent. Once the leaching process variables were optimized, a liquor containing 141 g/L Zn, 53 g/L Cd, 0.002 g/L Cu, 0.01 g/L Co and 0.003 g/L Ni was obtained from a residue containing 30 wt.% Zn, 26 wt.% Cd, 7 wt.% Cu, 0.35 wt.% Co and 0.32 wt.% Ni. The residue mass reduction exceeded 80 wt.%. Cementation studies investigated the influence of temperature, reaction time, zinc concentration in feeding solution, pH of feeding solution and metallic zinc excess. After that such variables were optimized, more than 99.9% of cadmium present in liquor was recovered in the form of metallic cadmium with 97 wt.% purity. A filtrate (ZnSO₄ solution) containing 150 g/L Zn and 0.005 g/L Cd capable of feeding the electrolysis zinc stage was also obtained.     

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%.     

Antimony Mobilization through Two Contrasting Gold Ore Processing Systems,New Zealand

Antimony, a toxic metalloid similar to arsenic, is present at variable levels in most gold-bearing rocks. Antimony is soluble in the surface environment, so antimony (Sb) mobilization in mine waters is an environmental issue around gold mines. The Reefton gold mine was originally developed in gold-bearing quartz veins; Sb concentrations were low (<100 mg/kg) compared to arsenic (As) concentrations (>1,000 mg/kg), and the mine waters had low dissolved Sb (<0.1 mg/L). A second stage of gold mineralization at Reefton involved brecciation and cataclasis of quartz veins and wall rocks, with addition of stibnite (Sb₂S₃). Processing of this ore has resulted in higher dissolved Sb in mine waters (0.1–1 mg/L), even after water treatment that removes most dissolved As (to 0.01 mg/L) by adsorption to suspended iron oxyhydroxide. Competition between As and Sb for adsorption sites on iron oxyhydroxide particles may have resulted in partial exclusion of the more weakly adsorbed Sb. The high rainfall (2,000 mm/year) at Reefton ensures adequate dilution of mine waters after discharge. The Macraes gold mine has no stibnite, and most Sb is in solid solution in the abundant arsenopyrite (Sb up to 2,000 mg/kg). Pit waters have both Sb and As dissolved up to 0.1 mg/L, partly because of evaporative concentration in a low-rainfall environment. Macraes tailings waters have high As (up to 3 mg/L) but negligible Sb (<0.001 mg/L). Reefton mine gold-bearing concentrate, containing stibnite, is transported 700 km to be processed by autoclave oxidation and cyanidation at the Macraes mine. This introduction of additional Sb to the Macraes site substantially increases the Sb content of the process stream periodically. Tailings from this process have up to 3 wt% Sb, dispersed through As-rich iron oxyhydroxides that are formed in the autoclave. The Sb-rich tailings are strongly diluted (approximately 100:1) by the Macraes tailings, and adsorption of Sb to iron oxyhydroxides in the tailings piles ensures that there has been no increase in the Sb content of the tailings water since the Reefton concentrate has been added at Macraes.     

Covering Pre-existing, Acid-producing Fills with Alkaline Sandstone to Control Acid Mine Drainage

In Boone County, West Virginia, small valley fills composed of overburden from the Number 5 Block coal seam were constructed during previous mining operations before overburden analysis was required to predict acid-producing potential. Using the modified acid–base accounting procedure, we found that these fills contain acid-producing materials, and in fact, they do produce acidic drainage. Black Castle Mining Company began operating in the area during the late 1980s and, in the process of removing coal from deeper in the geologic column, encountered alkaline sandstone, which they used to construct valley fills around and below the pre-existing fills that were discharging acidic drainage. Water quality from existing fills had pH from 3.5 to 4.5, and acidities up to 200 mg/L as CaCO₃. After covering the existing fills with alkaline sandstone, drainage water had a pH of >6.2 and contained net alkalinity. This has eliminated the need for chemical treatment of acid mine drainage from the pre-existing, pre-law valley fills.     

Antimony Mobilisation and Attenuation During Ore Processing at Orogenic Gold Mines,Southern New Zealand

Mine waters at the Reefton orogenic gold mine (active from 2007 to 2016) in southern New Zealand contained dissolved arsenic (As) and antimony (Sb) up to 5 mg/L during production of a sulfide concentrate that included arsenopyrite (FeAsS) and stibnite (Sb₂S₃). Ferric oxyhydroxide adsorption extracted As down to?<?0.1 mg/L but dissolved Sb remained elevated due to adsorption competition with As. The Reefton sulfide concentrate was transported 700 km to the Macraes orogenic gold mine (active from 1990) for processing through a pressure oxidation autoclave at 225 °C. The Macraes ore has low Sb contents, so the temporary introduction of a Sb-rich component produced a short-term Sb signal in the autoclave system and tailings waters. Oxidation of stibnite occurred rapidly in the autoclave, in parallel with the As in the arsenopyrite, producing ferric antimonate (tentatively identified as As-bearing tripuhyite) and ferric arsenate (FeAsO₄). Dissolved Sb in the Macraes tailings waters remained?<?0.1 mg/L throughout the period of Reefton concentrate addition. The formation of tripuhyite in the high-temperature autoclave stabilised Sb in the Macraes tailings, so that dissolved Sb?<?As, in contrast to the low-temperature processes at Reefton where dissolved Sb?>?As.

     

Hydrologic Impacts of Multiple Seam Underground and Surface Mining: A Northern Appalachia Example

An underground mine complex overlain by extensive surface mining in north-central Pennsylvania is drained principally by one discrete discharge point at which the flow rate (median of 2,167 L/min) increased significantly (67%) above background (median of 1,317 L/min) during a 3 year period. The source of this major discharge rate increase and other unusual hydrologic characteristics were investigated. Subsequent to background monitoring, about 440 ha of surface mining and reclamation (85% of the recharge area) occurred on numerous seams overlying the underground mines, which induced greatly increased infiltration rates. A direct correlation was observed between the surface mined area and increased recharge to the underlying deep mines. Atypically, in-mine storage does not exist to any substantial degree in the basal Lower Kittanning underground mine from which the main discharge emanates. The overlying Middle Kittanning mine is the main storage unit for mine water. The Middle Kittanning mine behaves like a perched aquifer system because of the moderate vertical hydraulic conductivity (a median rate of 1.0 × 10⁻⁷ m/s) of the thin (mean of 11.7 m) clay-rich shale and siltstone interburden and local structural features. During periods of low recharge, pool levels decline to a point where most of the mine water flowing downward from the Middle Kittanning mine to the underlying Lower Kittanning mine is diffuse in nature. The discharge rate is consistently in a narrow range of 1,745–2,381 L/min about a median of 2,040 L/min. When surface infiltration rates are high, the mine pool levels rise, and a portion of the recharge from the Middle Kittanning mine to the lower seam mine is apparently more channelized, flowing through the backfill over the buried highwalls and into the underlying Lower Kittanning mine. During these periods, the flow ranges more broadly from 5,725 to over 11,356 L/min, about a median of 8,328 L/min.     

Hydrologic Characteristics of a 35-Year-Old Underground Mine Pool

The hydrology of a 14,672 acre (5,940 ha) coal mine complex in Cambria County, Pennsylvania, USA, was characterized. This flooded mine complex was evaluated to determine the potential of using the mine water for downstream agricultural purposes in an adjoining watershed. The hydrologic characteristics of the mine complex dictate the amounts and rates of mine water discharge that are available. The original coal extraction rate was known to be 63%, but post-mining subsidence has reduced the effective porosity to a mean of 11%. Thus, the mine stores considerably less mine water than was anticipated, a priori. The mine receives vertical recharge averaging 0.27 gallons (gal) per minute per acre (24.6 L/s per ha), which is equivalent to 11.6% of the mean precipitation. The recharge rate fluctuates about the mean by ±22%. The low storage capacity combined with the moderately low recharge rates allow the large mine complex to be rapidly drawn down when the pumping rate is raised from 4.68 to 9.36 × 10⁶ gal (17.7–35.4 × 10⁶ L/day). Conversely, the mine refills rapidly, up to 0.8 ft (0.24 m) or spatially 33 acres (13.4 ha) per day, once the pumping rate is reduced back to 4.68 × 10⁶ gal/day (17.7 × 10⁶ L/day), which is well below the total recharge rate. In addition to vertical recharge, 6.3–40.4% of the inflow into the mine pool complex occurs from coal barrier seepage from an adjacent flooded mine. The seepage rates are relatively constant and are estimated to be insensitive to changes in head up to 50 ft (15.2 m).     

Influence of silver ions on bioleaching of cobalt from spent lithium batteries

The influence of silver ions (Ag⁺) on the bioleaching of cobalt from spent lithium batteries using Acidithiobacillus ferrooxidans (A. ferrooxidans) bacteria was investigated. The best effect was observed at Ag⁺ concentration of 0.02 g/L, and the amount of leached cobalt was 98.4% in a period of 7 days, whereas in the absence of Ag⁺, the amount of leached cobalt was only 43.1%. Based on the results of SEM, XRD and EDX investigation, it is deduced that Ag⁺ first reacted with cobalt to form AgCoO₂ as an intermediate. A mechanism that is based on catalytic interactions is proposed to explain the enhanced leaching of Co using A. ferrooxidans.     

CIGS废料综合回收利用工艺研究

本文针对CIGS废料采用“硫酸化蒸硒-水浸-铜萃取-铟萃取-中和沉镓-碱化造液”工艺梯次回收硒、铜、铟、镓四种元素,并产出粗硒、阴极铜、精铟和金属镓产品。在优化后试验条件下,蒸硒过程硒挥发率为99.5%,萃取过程铜、铟的萃取率分别为98.4%和99.9%,中和沉镓-碱化造液过程镓浸出率为99.5%。     

微波辅助碱分解低品位黑(白)钨精矿

考察了微波辐射加热条件下微波辐射功率、NaOH用量及质量浓度、精矿粒度等对低品位黑(白)钨精矿中WO3浸出率的影响。结果表明:在常压、微波功率为320W、NaOH用量为理论量的2倍、NaOH质量浓度为500g/L、精矿粒度为-0.074mm占80%的条件下,含WO330%左右的黑、白钨混合精矿经40min左右的分解,WO3浸出率可达到96%以上。     

Leaching behaviour of natural and heat-treated brannerite-containing uranium ores in sulphate solutions with iron(III)

Uranium leaching tests were conducted on two naturally occurring, highly metamict brannerite ores from the Crockers Well and Roxby Downs deposits, South Australia. The ores were leached over a range of temperatures and Fe^(III) and H₂SO₄ concentrations. As well, samples of the ores were calcined at 1200 °C in air to investigate the effect of thermally induced recrystallisation on uranium dissolution. For the unheated samples, a maximum of ∼80% U dissolution was obtained using an Fe^(III) concentration of 12 g/L, an acid concentration of 150 g/L H₂SO₄ and a temperature of 95 °C. The heat treated samples performed poorly under identical conditions, with maximum uranium dissolution of <10% recorded. High uranium dissolution from natural brannerite can be achieved providing; (i) acid strength, oxidant strength and temperatures are maintained at elevated levels (compared to those traditionally used for uraninite leaching), and, (ii) the brannerite has not undergone any significant recrystallisation (e.g. through metamorphism).