Search Documents

By S. Kursunoglu

The aim of this study was to investigate the effectiveness of Ascorbic (C6H8O6) Acid (AA), Activated Carbon (AC) and Guar Meal (C6H12O6) (GM) as a reductant for the simultaneous complete dissolution of zinc (Zn) and manganese (Mn) from a spent and mixed Zinc-Carbon (Zn-C) and alkaline battery powders in sulphuric acid (H2SO4) leaching. AA was tested previously as a reductant by other researchers while AC and GM were tested as a reductant for the first time in battery recycling. The effects of H2SO4 concentration, AA/AC/GM dosages, reaction temperature and leaching time on the Zn and Mn dissolutions were investigated according to 2k full factorial experimental design and Central Composite Design (CCD) technique; then, a simple optimization study was carried out for comparing the reductants. The optimum reductive acid leaching conditions for AA were determined at 3 hours leaching time, 70°C leaching temperature, 0.5M H2SO4 concentration, 13 g/L AA dosage, 1/20 g/mL solid/liquid (S/L) ratio and 200 rpm stirring speed. Under these conditions, the dissolution efficiencies were 99.9% for Zn and 99.3% for Mn. The best reductive acid leaching conditions for AC were determined at 3h leaching time, 80oC leaching temperature, 1M H2SO4 concentration, 3 g/L AC dosage and 400 rpm mixing speed. Under above conditions, Zn and Mn dissolutions were not complete. At 2M H2SO4 concentration, 10 g GM dosage, 700C leaching temperature, 200 rpm stirring speed, 1/20 g/mL solid/liquid (S/L) ratio and 3 hours leaching time, full dissolution of both Zn and Mn was achieved. As a result, three different reductants were evaluated in this study for complete simultaneous Zn and Mn dissolutions. This paper also compared present test results with previous spent battery recycling studies for Zn and Mn dissolutions. Full valuable metals dissolutions from spent batteries were achieved at low H2SO4 concentrations with AA and high concentrations with GM. Selective precipitations of Zn and Mn from acid solutions were carried out using NaOH and KOH. Using 3M NaOH as precipitating agent at room temperature and at pH 8; 95.4% of Zn and at pH 10; 93.7% of Mn were precipitated from the leach solution. Under the same conditions, using 3 M KOH at pH 8; 91.6% of Zn and at pH 10, 96.4% of Mn were precipitated. Based on the experimental results obtained, an appropriate flow sheet was proposed to recover both Zn and Mn.

Aug 1, 2013

By Hiwa Salimi, Foley. Stephen R, Loghman Moradi

Common water-miscible organic solvents including DMSO, DMF, methanol, ethanol, acetone and THF in combination with HCl and low concentrations of H2O2 were investigated for dissolution of gold. THF showed the best result demonstrating that gold powder could be completely dissolved in 1.5 hours in 1.0 mol/L HCl and 0.2 mol/L H2O2. No complete gold dissolution was observed for the rest of solvents even after 24 hours. The rate of gold dissolution was measured by dissolving gold wire in THF/HCl/H2O2 mixture. The rate of gold dissolution in optimised condition was 75 g m-2 h-1 at room temperature and significantly increased to 590 g m-2 h-1 by raising the temperature up to 60 oC. Although a THF/HCl/H2O2 mixture showed a slower dissolution rate in comparison with aqua regia, however the applied conditions are considerably milder which is highly promising for gold leaching in an industrial scale.

Jan 1, 2015

By A. Y. Senchenko, S. G. Rybkin, A. V. Aksenov

In gravity processing of gold-containing ore the target produced concentrate which is high gold grade primary concentrate contains a significant amount of cassiterite (SnO2). The plant attempts aiming at concentrate processing using a standard pyrometallurgical technology did not provide any positive results. The present paper is dedicated to development and testing of improved technology for pyrometallurgical processing of rich gold primary concentrate with high cassiterite content. According to material composition study primary concentrates produced at the plant contained 1.5–2.0% Au, 0.1–0.3% Ag, 55–65% SnO2, 20–25% Fe2O3, and 10–15% Fe3O4. The developed technology included alkaline nodulizing of primary concentrate mixed with sodium hydroxide and carbonate at 350–650°? with cake production. In blend nodulizing stannum and iron passed into sodium stannate (Na2SnO3) and sodium ferrate (NaFeO2) respectively. The blend of cake and fluxing agents was smelted at 1,150–1,200°? resulting in dore gold and slag production. Sodium stannate with 360°? melting point is well dissolved in slag system Na2OSiO2- B2O3-CaO. During plant tests 18.5 kg of high gold grade primary concentrate were processed, 1350 g of dore gold and 75 kg of slag were produced. Dore gold contained 16.5% Au, 0.95% Ag, 78.0% Sn, 3.1% Fe. Slag contained 12.4 g/t Au, 6.7 g/t Ag, 9.8% Sn, and 5.6% Fe. Gold recovery to target alloy was equal to 99.58 %, 87.8% of Sn passed into slag. This new technology was accepted for application at the plant.

Jan 1, 2016

By C. W. Hancher, B. D. Patton, W. W. Pitt

"There are a number of nitrate-containing wastewater sources, as concentrated as 30 wt % NO3 and as large as 2000 m3/day, in the nuclear fuel cycle. The biological reduction of nitrate in waste water to gaseous nitrogen, accompanied by the oxidation of a nutrient carbon source to gaseous carbon dioxide, is an ecologically sound and cost-effective method of nitrate waste disposal. These nitrate-containing wastewater sources can be successfully biologically denitrified to meet discharge standards in the range of 10 to 20 g N(NO3-)/m3 by the use of a fluidized bed bioreactor. The denitrification bacteria are a mixed culture derived from garden soil; the major strain is Pseudomonas. In the fluidized-bed bioreactor, the bacteria are allowed to attach to 0.25- to 0.50-mm-diam. coal fluidization particles, which are then fluidized by the upward flow of influent wastewater. Maintaining the bacteria-to-coal weight ratio at approximately 1:10 results in a bioreactor bacteria loading of greater than 20,000 g/m3,This paper describes the results of a bio denitrification R&D program based on the use of fluidized bioreactors capable of operating at nitrate levels of up to 7000 g/m3 and achieving denitrification rates as high as 80 g N(N03-) per day per litre of empty bioreactor volume. IntroductionThe discharge of wastewater containing nitrate at concentrations greater than stipulated by the Environmental Protection Agency (EPA) creates an environmental hazard and constitutes a civil and/ or criminal legal liability.(I.2) At present, we have three options: (1) recycle and reuse of nitrate streams the total economic and technical balance should be Closely examined; (2) use of an alternative acid, which may pose an entirely different set of problems ; or (3) disposal of nonreusable nitrate sources by an environmentally acceptable procedure."

Jan 1, 1980

By T. Kalliomäki, A. T. Aji, M. Lundström, J. Aromaa

The settling properties of anode slime in copper electrorefining have crucial effect on copper cathode quality. Anode slime settling is affected by electrolyte composition and temperature, electrolyte circulation, and the sizes and density of the slime particles and agglomerates. Typically, the industrial copper electrolytes contain 40–60 g/dm3 Cu2+, 130–220 g/dm3 H2SO4, 0.04–25 g/dm3 Ni2+, 0.005–30 g/dm3 As, and have operating temperature of 55–70°C. In optimal electrorefining process, the settling rate of anode slime is sufficiently high to avoid the risk of getting the slime entrapped in copper cathodes. In this research the settling properties of copper electrorefining bottom slimes, mainly consisting of Cu, Ni, Ag, Se, Pb, As, Bi, Sb and Ba compounds, were studied by settling rate tests. The tests were conducted in water and synthetic electrolytes with known viscosities and densities. In addition, the effect of agglomeration was taken into account – it was observed that breaking the agglomerates which existed in the as-received slimes decreased the particle sizes on average 56–73%, which decreased the settling rate. The results of this study provide detailed information of anode slime settling and thus clarify the conditions causing cathode contamination risks in copper electrorefining. INTRODUCTION Settling rates of Cu-ER anode slimes affect the purity of cathode copper and faster settling rate decreases the contamination risk (Davenport et al., 2002; Shi & Ye, 2013). The settling rates are affected by the properties of anode slime particles (Wills & Napier-Munn, 2011; Aaltonen, 2014; Kiviluoma, Aaltonen, Aromaa, Lundström & Forsén, 2016) as well as viscosity and density of the electrolyte (Davenport, King, Schlesinger, Biswas, & Robinson, 2002; Shi & Ye, 2013). Thus, it is crucial to find out how the viscosity and density affect the settling rates. Small spherical particles with diameter below approximately 50 µm obey Stokes’ law (Equation 1) (Wills & Napier-Munn, 2011). v = gd 2 (ρS – ρF) (18η) -1 (1) where v is terminal velocity (m/s), g acceleration due to gravity (m/s2), d diameter of particle (m), ρS density of particle (kg/m3), ρF density of fluid (kg/m3), and η dynamic viscosity (Pa∙s). The approximate effects of kinematic viscosity (ν = η / ρS), density (ρS) and particle sizes on settling rates defined for small anode slime particles based on Stokes’ law are presented in Figure 1. However, Stokes’ law does not take into account other than spherical shapes, and it is only applicable for small particles. Consequently, the real settling rates of the particles are lower than these (Wills & Napier-Munn, 2011). Moreover, some of the anode slime particles are larger than 50 µm (Aaltonen, 2014).

Jan 1, 2019

"Batch flotation testwork demonstrated that a low-grade chromite ore from Bird River, Manitoba, was best floated at pH 2 using Armac C (coco amine acetate) collector and sulphuric acid as pH modifier. A grind of 100% minus 150 um was necessary for liberation, and flotation selectivity was greatly enhanced by prior desliming. A concentrate analyzing 24.2% Cr203 with a recovery of 90.2% was obtained, after three cleaning stages, from ore grading 6.05% Cr203. Total rougher collector concentration was 500 g/t and cleaner collector concentration 380 g/t. Similar grade and recovery were obtained during the pitot-plant evalualion, when a concentrate analy zing 22.3% Cr203 with a recovery of 93% (based on deslimed feed) was obtained for a total collector addition of 630 g/t.Selective flotation in a region of positive surf ace charge implies anionic activation, and a possible flotation mechanism is suggested.IntroductionIn an earlier study(1) of the beneficiation of chromite ore from Bird River, Manitoba, several methods of beneficiation were explored. These included ore sorting, heavy-media separation, and gravity concentration with flotation using Armac C (coco amine acetate) collector at pH 2, and sulphuric acid as pH modifier. Because flotation was not fully evaluated, a laboratory study was undertaken to determine the optimum parameters of chromite flotation. Optimization of parameters was established first on a laboratory scale followed by confirmation a t the pilot -plant level. Flotation of chromite using Armac C and sulphuric acid as a pH modifier has been reported as being selective for low-grade chromite ores(2)."

Jan 1, 1990

By Yanping Xiao, Ton van Sandwijk, Gus Van Weert, Larry McHugh, Thomas Zabev

"All attempts to remove significant amounts of arsenic from enargite via neutral or reductive pyro routes have required operating temperatures between 600 and 700 º C, too high to propose an equipment development program. Operating at reduced pressures, or using hydrogen or elemental sulphur as a sweeping gas did not help to lower these temperatures. Admixing iron powder gave excellent results, but still required 700 º C. By fortuity, in recent experiments to produce tennantite from an enargite mineral sample, it was observed that enargite was completely converted to bornite, chalcopyrite with minor residual tennantite within an hour at 400-450 ºC. This is believed to be due to the presence of pyrite in the mineral sample treated. All earlier decomposition work appears to have been done on predominantly enargite (i.e. low pyrite) samples, or at high operating temperatures, ignoring possible reactions at lower temperature. The decomposition of enargite into iron substituted tennantite plus some bornite under evolution of liquid elemental sulphur, followed by further decomposition of this tennantite into bornite and volatile arsenic sulphides, appears experimentally feasible at these low temperatures and mimics described ore genesis pathways at > 250 ºC in saline solutions. We suggest two pathways to describe the observed experimental results:20 Cu3AsS4 + 12 FeS2 ???5 Cu10Fe2As4S13 + 2 Cu5FeS4 + 31 S(l)31 S(l) ???31 S(g) (evaporation)5 Cu10Fe2As4S13 ???10 Cu5FeS4 + 10 AsS(g) + 5 As2S3(g)"

Jan 1, 2014

By V. Barranco, K. Yliniemi, A. Hubin, B. P. Wilson, D. Janas, M. Lundström, P. Hannula

Typical copper electrorefining electrolyte contains tens of grams of copper, some nickel and a wide variety of impurities. Some of these impurities comprise trace amounts of dissolved precious metals like Ag, Pd and Pt that currently are not fully exploited. This study presents an electrochemical method to recover precious metals via a combined electrodeposition-redox replacement (EDRR) process, where energy is used only to deposit copper during the ED-step, while the RR-step relies on the spontaneous replacement of copper by the precious metals (i.e. cementation). In order to conduct the recovery successfully, process parameters such as deposition potential (EDE), deposition time (tDE), redox replacement time (tRR) and the number of cycles (n) was adjusted. Investigations performed in synthetic electrolyte (40 g/L Cu, Ag = 10 ppm) proved the possibility to recover a high purity silver. Short deposition times (tDE) were found to result in small Cu particles that provided a suitably high surface area for silver cementation. In addition, experiments performed in real industrial electrolytes containing 44 g/L Cu and 54 ppb Ag (Boliden Harjavalta, Pori), demonstrated that EDRR could attain Ag recoveries of between 46–67% in an energy efficient, additive-free manner. INTRODUCTION The raw materials used in the state-of-art primary and secondary copper smelters contain mainly base metals, but also notable amounts of Ag, Pt and other precious metals. These valuable metals distribute mostly into the copper matte during the pyrometallurgical treatment (Tirronen et al., 2017) and end up into the cast anodes. During the electrorefining stage, metals more noble than copper such as gold and platinum, are not dissolved but enrich into the anode slimes. In contrast, silver may dissolve slightly and thus, minor chloride addition is used to retain silver in the slime as sparingly soluble AgCl. Moreover, other metals less noble than copper may also dissolve, such as nickel. Therefore the electrolytes used in electrorefining typically contain substantial amounts of copper, nickel and even arsenic as well as minor metals such as antimony and bismuth. In the Boliden electrorefinery these were reported to be in the magnitude of 51.2– 60.4 g/L of Cu, 10.9–15.8 g/L Ni, 7.2–15.9 As, 0.2–0.4 Sb and 0.1–0.4 Bi with an associated sulfuric acid concentration of 139–155 g/L (Lehtiniemi et al., 2018). Copper is the most noble of the dominating base metals in the electrolyte and is thus deposited on the cathode as 99.99% purity LME-A grade. Anode slimes are collected and subjected to precious metals treatment e.g. via pressure leaching—selenium recovery— Doré smelting—Ag refining—Au/Pt/Pd leaching and reduction process (Ludvigsson & Larsson, 2003). The Pori refinery, operating in Finland since 1941, not only produces 133 000 t Cu cathode and 2300 t nickel sulphate annually but also 66 t silver, 4.6 t gold and palladium and platinum (in year 2017).

Jan 1, 2019

By K. Gie Tan, Frank Wright

"Misty Mountain Gold Limited (MGL - TSE VSE, MGLCF - NASDAQ) of Vancouver, owns the Harmony Gold Project located in the Queen Charlotte Islands of British Columbia. The deposit contains a mineable reserve of 2.3 million ounces of gold, at a 1.2 g/t cut-off and an average grade of 2.1 1 g/t Au.Discovered in 1970 and frequently investigated by previous owners, full-scale production has not yet occurred because of the complex metallurgy of the deposit. Recent process studies that were completed in six months at Process Research Associates Ltd. (PRA) of Vancouver resulted in significant improvements to the project viability. The work involved detailed mineral processing studies aimed at achieving a marketable concentrate. An option for bio-oxidation and cyanidation to allow on-site dor6 production was also investigated.The main findings at PRA followed the front-end recovery of finely divided free gold (Au) by gravity concentration. Approximately 20% of Au reported to less than 0.2% of the mass, resulting in a grade exceeding 500g/t Au. The resulting gravity tailing provided a uniform feed for flotation evaluation. The optimization of this step allowed for a reasonable primary grind size of 80% passing (&) 50pm. Efficient slime’s rejection was made possible by flotation at a reduced pulp density, and pre-oxidized sulphides were reactivated with Na,S. The regrinding of the bulk flotation concentrate followed by up to five stages of cleaning resulted in a combined gravity + flotation product of 50g/t Au with 82% recovery in less than 4% of the feed mass.Further studies indicated that the flotation concentrate was amenable also to tank bio-oxidation. At 65% sulphide oxidation the gold recovery improved from less than 70% to 90% by cyanide extraction using CIL (carbon in leach). This compares to less than 45% gold recovery from whole-ore cyanidation in several previous studies. Heap leaching may also be considered for lower grade material. PRA’s methodical approach and consistent results dispelled the uncertainty about predicting the material behavior for the Harmony Property."

Jan 1, 2000

By Z. Yang, M. J. Collins, N. E. Tuffrey, R. J. Kalanchey, I. M. Masters

The high elevation of China Western Mining Company (WMC)' s Ganhe refinery, and the high iron content of concentrates produced by the WMC mines, were key factors in the selection of pressure leaching by WMC for their new zinc refinery. This paper describes the results of pressure leaching test work conducted by Sherritt in Fort Saskatchewan with a sample of WMC concentrate from Qinghai Province in China. Results from batch pressure leach tests and a four-day pilot plant campaign are reported. The pilot plant campaign included continuous operation of a two-stage pressure leach circuit, as well as neutralization and iron removal steps, resulting in the production of iron-free zinc sulphate solution. Greater than 98% zinc extraction was achieved in pressure leaching, while producing solution containing about 9 g/L H2SO4 and 2 g/L Fe. Following neutralization and iron removal, the solution contained about 170 g/L Zn and less than 10 mg/L Fe. Overall recovery of sulphur in the zinc concentrate feed to elemental sulphur byproduct was about 85%, while 75 to 80% of the indium in the feed was recoverable from the neutralization residue. The test results highlighted the technical and economic advantages offered by the Sherritt Zinc Pressure Leach Process and provided the required data for the engineering design.

Jan 1, 2008

By P. E. Weidmark

"PART I - Published PapersBallivy, G., and Dupuis, M., ""Laboratory testing for the design of grouted rock anchors""; Proc 6th Pan Am Conf Soil Mech Found Eng; Lima; 1:277-291; 1979.Ballivy, G., Ladanyi, B., and Gill, D., ""The effects of moisture content on the strength of fissured rocks""; Proc 1// 1ntnl Congr 1.A.E.G.; Madrid; //:2:7-16; 1978.Baumgartner, P., and Stimpson, B., ""Development of a tiltable base friction frame for kinematic studies of caving at various depths""; Int J Rock Mech Min Sci Geomech Abstr; 16:4:265-267; 1979.Bawden, W.F., and Roegiers, J.-c., ""Discussion and interpretation of the mathematics governing radial flow in a single horizontal rock fracture""; ISBN 07727-7002-6; Dept Civil Engng publication 79-11; Univ Toronto: 134p;1979.Chatterji, P.K., Smith, L.B., and Insley, A.E., ""Construction of Saline Creek Tunnel in Athabasca oil sand""; Can Geotech J; 16:1:90-107: 1979.Coates, D.F., and Gyenge, M., ""A case history of pit slope design""; Proc 4th ISRM Congr; Montreux, Switzerland; Sept 1979; 1:Theme IV:591-596; 1979.Crawford, A.M., and Curran, J.H., ""A dynamic direct shear machine for determining the time-dependent behaviour of rock joints"": Dept Civil Engng publication 79-08; Univ Toronto; 1979.Crawford, A.M., and Curran, J.H., ""A displacement discontinuity approach to modelling the creep behavior of rock and its discontinuities""; Dept Civil Engng publication 79-09; Univ Toronto; 1979.Crawford, A.M., and Curran, J.H., ""A higher-order displacement discontinuity element""; Dept Civil Engng publication 79-10; Univ Toronto; 1979.Cruden, D.M., ""A computer-based system for rock mass description""; Proc 1// Intnl Congr I.A.E.G.; Madrid; //:2:155-159; 1978.Cruikshank, D.J., and Curran, J.H., ""A.B.LE. approach to modelling of fluid flow in fractured rock masses""; Proc 7th Cdn Congr Appl Mech; May 1979; Sherbrooke, Que; 887-888; 1979."

Jan 1, 1980

By F. Parada, R. Parra

"The neutralization process of SO2 from coal burning plants is a well-defined technology using limestone to form CaSO4 according to the following reaction: CaCO3 + SO2(g) + 1/2O2(g) = CaSO4 This paper presents the results of a study carried out to validate the option of capturing SO2 using gases from a non-ferrous smelter; SO2 content was up to 12% using a three-stage fluidized bed reactor. The difference in SO2 content affects equilibrium conditions and kinetics, rendering a direct extrapolation of the high efficiency achieved for low SO2 content impossible. The R&D program consisted of a laboratory study using a semi-pilot operation of a one-stage fluidized bed reactor and the construction and operation of a three-step reactor prototype. Results show that the reaction proceeds according to the unreacted core model that imposes a compromise between CaCO3 conversion and SO2 capture. For high SO2 capture, the calcine from the reactor contained CaSO4 and CaO. A complete capture of SO2 from gases with 8 vol.% can be achieved with a mean reaction time of two hours at 800°C."

Jan 1, 2007

By J. R. Parga T., P. J. Valenzuela G., L. S. Quiroz C., P. Guerrero G., J. L. Valenzuela G.

"Pressure oxidation is a hydrometallurgical pre-treatment process used for refractory auriferous sulfide ores. In the work reported here, a refractory pyrite ore sample from the Mulatos mine, Sonora, Mexico was characterized. The sample was subjected to atmospheric cyanidation, and simultaneous pressure oxidation and cyanidation under different conditions of temperature and pressure. The mineral characterization showed that pyrite is the main species in the sample. The gold extraction tests were carried out in an 1-litre Parr pressure vessel at a temperature range of 90-120 °C, a pressure range of 87.5–150 psi (using oxygen), and a cyanide concentration range of 0.8-1.2 %wt. The gold extraction results showed that diffusion through product layer is the rate controlling step.INTRODUCTIONIn hydrometallurgical precious metals extraction, the most commonly used process is sodium cyanide leaching, as shown in the next reactions involving gold and silver (Fleming, 1992; Marsden & House, 2006):4Au(s) + 8CN- (aq) + 2H2O(l) + O2(g) ? 4Au(CN)2 -(aq) + 4OH- (aq) (1)4Ag(s) + 8CN- (aq) + 2H2O(l) + O2(g) ? 4Ag(CN)2 (aq) + 4OH- (aq) (2)“Refractory” auriferous ores or concentrates often contain gold and silver locked inside other sulfide phases such as pyrite. This renders cyanidation ineffective, so oxidative pre-treatments are used in order to convert the sulfide minerals into porous oxide minerals and make cyanidation work more efficiently. Reaction (3) shows the oxidation of pyrite in alkaline media and the formation of the porous oxide (Koslides & Ciminelli, 1992):2FeS2 (s) + 15/2 O2 (g) + 8H- (aq) ? Fe2O3 (s) + 4SO4 2- (aq) +4H2O(l) (3)A process of simultaneous chemical pressure oxidation and cyanide leaching for a flotation concentrate produced in the mine of Bacís, Durango, in Mexico, gave good results (Parga et al., 2007). In the study presented here, it was intended to use this treatment for a refractory pyrite mineral from the Mulatos mine in Sahuaripa, Sonora, Mexico, modifying variables such as pressure, temperature and cyanide concentration to find the best conditions for oxidation and leaching."

Jan 1, 2012

By J. C. Gathje

Newmont&apos;s Phoenix Project is located near Battle Mountain, Nevada and is scheduled to produce copper concentrates containing high precious metal values. High temperature pressure oxidation (PDX) of the concentrates was considered as a treatment option during process development. The metals content in Phoenix concentrates range from: 15-25% for Cu, 1-10% for Zn, 0-2% for Pb and ?27% Fe. The precious metals content of the Phoenix concentrates can range up to ?300 g/t Au and >3,000 g/t Ag. As a consequence it was necessary that any process for consideration should have the capability of recovering both the copper and precious metal values. Typically, gold can easily be leached from the autoclave discharge solids, but silver is chemically bound with the jarosite formed and cannot be directly leached. Laboratory investigations led to the development of a patented PDX process that minimizes the formation of jarosite so that the silver can be leached easily along with the gold. The process was confirmed with extensive pilot plant testing. Excellent solution extraction of copper and zinc was achieved, and both iron and lead were precipitated in the PDX residue. The process produces clean PDX discharge liquors with very low iron concentrations. As an example, conventional PDX techniques resulted in pregnant liquors containing ?5 g/L Fe. However, the Newmont process, which controls the iron and sulfate species, achieved high iron precipitation in the PDX solids and pregnant liquors containing <0.5 g/L Fe. High gold and silver recoveries were achieved by direct cyanidation of the PDX solids. This process has the added benefit of producing PDX solids that demonstrate superior settling and filtration characteristics.

Jan 1, 2006

By F. Barriga-Ordonez, J. A. Finch, L. Calzado, S. R. (Ram) Rao

Sludge, formed by lime treatment of acidic drainage and stored in tailing ponds of the Sudbury region, contains 5-7% Ni, mostly in a mix of hydroxides with Fe (~ 10%), Mg (~ 10%) and Cu (-0.5%), and Ca (-3.5%) as sulphate. Investigations were conducted to recover nickel by a 2-stage process. In the first stage, the sludge was leached with sulphuric acid at pH 3.2. About 90% of the nickel was extracted, producing a solution, which contained mainly nickel (2 g/L) and magnesium (3.7 g/L). In the second stage, nickel was separated by treating with ozone which recovered 99% of the nickel in the Ni(III) oxidation state as -85% pure nickel oxide-hydroxide (NiOOH) precipitate. The oxide settled and filtered readily. A potential use of the product in battery making is suggested. The leach residue contained 0.5% Ni making it less toxic for disposal. Some physical properties of nickel oxy-hydroxide and process parameters including ozone consumption will be discussed.

Jan 1, 2005

By B. D. Pandey, D. Bagchi, Premchand

During the separation and recovery of metals from the ammoniacal leach liquor of Indian Ocean nodules by solvent extraction-electrowinning (SX-EW) in close-loop operation, the copper electrolyte containing ~40g/L Cu, 20 g/L Ni and 180 g/L sulphuric acid is bled-off to control the impurities. The recovery of copper and nickel as powders has been examined from the copper bleed solution following partial decopperisation, crystallization of mixed sulphate of the two metals and aqueous hydrogen reduction. From the dissolved sulphate solution containing 17.68 g/L Cu, 18.83 g/L Ni and 0.03 g/L Fe, the copper powder was precipitated at initial hydrogen pressure of 26 bar in an autoclave in 90 min at 140 °C. The residual copper from the spent solution was removed as sulphide and nickel powder was then recovered (99%) in two stages from a solution of pH 4.5 at 40 bar pressure and 190 °C. However, 98% nickel recovery in one stage was achieved when hydrogen reduction was carried out from an ammoniacal solution in the pH range 9-11. Copper and nickel powders were produced from the bleed solution with acceptable purity and specifications suitable for P/M applications.

Jan 1, 2004

By S. Koppalkar

ABSTRACT: The development and evolution of the gravity recoverable gold (GrG) test procedure at McGill university is reviewed in this paper. Since the late 1980s, the research group led by the late Professor André Robert Laplante at McGill university has undertaken fundamental studies and applied research of the Knelson concentrator to develop a sustainable method of estimating GrG. Historically, a concept of “free gold” was used, and it was de?ned as the gold that could be separated and recovered from gold ore by amalgamation. Due to health and safety hazards, the practical use of amalgamation has declined. around 1994, a methodology using a 3-in Knelson concentrator to measure gravity recoverable gold was established at McGill university. as a result of more than 10 years of improvement and optimization, this technique has become a standard method of evaluating GrG for gold ores, gold plant streams and platinum group metals. By 2006, about 200 GrG tests were carried out at McGill universityon ores from all over the world. recently, a modi?ed version of the GrG test (a simpli?ed test) was developed to provide a quick and rough estimate of the GrG in a sample; this simpli?ed test has theadvantage of reduced mass, processing time and cost. to recognize and honour the contributions madeby the late Professor Laplante towards the establishment of a standard test procedure for assessing gravity recovery potential, the authors propose that the GrG test be named after Laplante.

Jan 1, 2010

By David E. Pearson

"Western Canadian coking coals are located in the Rocky Mountains arid foothills of British Columbia and Alberta. They are typical of non-marine coals, characterized by total sulphur contents of less than one per cent and with ash contents of 10-30 weight per cent. This mineral matter is composed principally of kaolinite and quartz.The petrographic compositions of most of the coking coals which constitute the 2.065 billion tonnes of British Columbia&apos;s measured reserves, and the compositions of the metallurgical coal currently exported from Alberta, show them to be rich in the inertinite macerals, macrinite, fusinite and semifusinite. When the mean maximum reflectance of vitrinite in oil/Romax) is used as the coal rank parameter, this high inertinite content is reflected in the lower volatile-matter yields, but the relatively inert character is also displayed by lower free swelling indices (PSI), lower maximum dilatations, lower maximum fluidities and lower coke strengths than other coals of the same rank, but richer in the reactive macerals, vitrinite, exinite and semifusinite. Despite these reduced values in the so-called caking tests, inertinite-rich coals produce strong coke.A system of classification of coking coals is introduced which is based on rank, as indicated by Ro max, and either petrographic composition, maximum dilatation, maximum fluidity or, to a lesser extent, PSI and volatile-matter yield. Six principal coal groups can be distinguished among coking coals using this versatile classification. Inertinite-rich coals are assigned to the Keystone Group (GI), Balmer Group (G3) or Moura Group (G4), each group being named for a coal typical of that population ofcoals. Reactive-rich coals are assigned to the Pittston Group (G2), Kellerman Group (G5) or Big Ben Group (G6). A multiple-regression analysis of coal quality versus price shows how the value of coking coal in the Japanese market appears to be related to reflectance and the free swelling index, and confirms the economic viability of the classification.British Columbia&apos;s measured reserves of coking coal contain representatives of four of these groups; Balmer Group (73%), Moura Group (/6%), Pittston Group (3.4%) and Kellerman Group (7.6%).This analysis of the quality of Western Canadian coking coal indicates that despite the need for large proportions of inertinite-rich coking coal in blends produced in the Japanese steel industry, there is nevertheless a demonstrated overabundance of Balmer Group and Moura Group coals available to the international coking coal market."

Jan 1, 1980

By P. West-Sells, B. Wassink, N. Fisher

Gold leaching from a commercial cyanide-amenable gold ore was studied at 25°C using sodium bisulfide-hydrogen sulfide solutions containing elemental sulfur. Gold extraction from the ore using 0.55 M NaSH, 4.1 atm (60 PSI) H2S, 1 g/L S8 and 26% solids by weight was about as good as the best extraction obtained by cyanidation. Maximum extraction was achieved in roughly 48 hours. Sulfur dissolves in the bisulfide solutions to form a mixture of polysulfides. Simple speciation calculations indicate that gold solubility as bisulfide complexes should be roughly ten times higher than in the absence of added zerovalent sulfur. The principal bisulfide complex is believed to be [Au(HS)2]-. Gold dissolution is favoured by increasing HS- concentration and by increasing H25 pressure. Polysulfide complexes of gold are also likely to be formed, and may contribute substantially to gold leaching. Two experiments using pure gold were also performed. Up to 100 mg/L gold was dissolved in 140 hours using 0.55 M NaSH, 1 g/L added elemental sulfur and 6.2 atm H2S.

Jan 1, 2005

By Vincent Jourdain

Located in the James Bay Territory of Quebec (figure 1), the Clearwater property is jointly owned by SOQUEM Inc. and Eastmain Resources Inc. Exploration on the property since the early 70&apos;s has resulted in the discovery of several gold showings. Exploration work by the joint venture since 1995 has led to the definition of the Eau Claire Zone containing an indicated resource of 540 000 tonnes grading 11.1 g/t Au (8.7 g/t Au cut at 34.28 g/t Au). Hosted by the archean Eastmain Greenstone Belt, the property geology is dominated by mafic volcanics containing thin horizons of felsic volcanics and/or sedimentary rocks (figure 2). The sequence is metamorphosed to the amphibolite facies and contains two small toalitic intrusions. Regionally, the structural trend is oriented E-W but in the area of the Eau Claire Zone a second trend oriented NE-SW is also present. The first exploration work on the property was conducted by SEREM and SDBJ from 1973 to 1976. It consisted of an airborne electromagnetic survey and a ground follow-up program on several grids. This program, comprising geological mapping, geophysics and drilling, led to the discovery of the Natel showing where a drill intersection returned 8.89 g/t Au over 2.65 m. From 1978 to 1985 the area was withdrawn from staking because of the hydroelectric projects of the James Bay. In 1985, Westmin Resources Limited and Eastmain Resources began an exploration program based on a helicopterborne electromagnetic survey. Once again, a ground follow-up program was conducted on several grids. This program included a soil geochemical survey that identified a strong (up to 10 g/t) gold anomaly. Drilling of the anomaly led to the discovery of the Eau Claire Zone where resources were estimated at 270 000 tonnes grading 6.1 g/t Au at the end of 1990. In 1995, SOQUEM Inc. took an option to gain a 50% interest on the property and obtained its interest in 1998. Since 1995, two principal mineralization zones were identified on the property. The "Rosemary Zone" consists of veinlets and disseminations of pyrite and chalcopyrite within a magnetite-rich amphibolite. The mineralized zone was identified by stripping within an area of 1000 by 500 metres. Drilling indicates copper grades up to 0.32 % Cu over 49.2 m. with locally grades of 0.75 % Cu over 5.9 m. The mineralised zone is associated with QFP dykes and corresponds to a strong magnetic anomaly and a strong copper and molybdenum soil anomaly. The "Eau Claire Zone" consists principally of a set of gold bearing quartz-tourmaline veins oriented E-W and dipping at 45-50??to the south (figure 3). More than ten different veins are recognised in the zone and up to eight veins were observed in a single trench (figure 4). The host rock between the veins shows a degree of deformation similar to the regional volcanosedimentary rocks. Each vein is hosted by a unit of actinolite-tourmaline-biotite schist up to one meter thick. The veins and their hosting schist clearly cross-cut the schistosity of the hosting lithologies and the contacts between amphibolites and felsic to intermediate rocks. Usually, the veins are less than a meter thick and are composed of quartz, tourmaline and calcite. The proportion of quartz and tourmaline varies from 0 to 80% within a single vein. The veins commonly show massive, banded or brecciated texture. Sulphides, mainly pyrite, pyrrhotite and chalcopyrite, generally form less than 5% of the vein material.

May 1, 2001