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By A. J. Rigby

Anode vessels are often relined with refractory every one to two years and a typical reline time can take between 16-20 days. Modem refractory designs using dense, corrosion resistant brick applied directly to the vessel shell in the high wear areas around and above the tuyeres and at the slagline can yield a lining life of up to 4-5 years before a complete reline is necessary. Installation methods using brick delivery on roller conveyors through an endwall access port and complete anchoring of the brick to allow for lining self-support can take days off the installation time. It is possible, therefore, to reduce the refractory relining period by 50% and increase lining life by 100% compared to typical anode repair schemes. Cu2007

Jan 1, 2007

By P. Wanjara, M. Xiao, M. Jahazi, C. Poon

Using a robotic 4kW Nd:YAG-laser system, Inconel 718 (1 mm sheet) was welded with and without filler wire to investigate the effect of processing parameters, including power input and welding speed, on the weld geometry, microstructure and mechanical properties. By means of both nondestructive testing and microscopic examination, the weld quality was studied to determine the nature of the welding defects (porosity, cracking and bead shape) over the processing range of 3-3.5 kW for the power input and 4.5-6 m-min-1 for the welding speed. Application of the design of experiment methodology permitted the examination of the relationship between the process parameters and mechanical properties, from which the optimized laser welding conditions could be determined for joining Inconel 718. In particular, the research results indicate that Nd:YAG-laser welding technology can achieve defects free, mechanically sound welds with narrow fusion and heat effected zones.

Jan 1, 2006

Train on Different Equipment - * Immersive?s interchangeable cabin controls technology - ?clip on & off? * Change Simulator from one machine to another in minutes * No technical expertise required to swap * Real, functioning, controls and displays are used to aid rapid learning Immersive has produced a range of Conversion Kits to suit most equipment. Modular design - Easy to convert from one machine model to another with a Conversion Kit. Caterpillar Models Trucks - 769D, 773D, 777D Trucks - 785B, 789B, 793B Trucks - 785C, 789C, 793C, 797 Excavator/Shovel - 5130, 5230 Front end loader ? 992, 994 Dozer ? D10, D11 Komatsu Models Trucks - HD325, 785-5, 985-5 Trucks - 730E, 830E, 930E Dozers - D375A-3, D475A-3 Shovel - PC8000 P&H Shovels - 4100A, 4100XPB, 4100BOSS Hitachi Shovels - EX2500, EX5500 O&K Shovel - RH200 Bell Articulated Truck - B40C

May 1, 2003

By M. Mohammadi Zahrani

The morphological modification of primary and eutectic Mg2Si phases by bismuth (Bi) addition in Mg-5wt.% Si in-situ composite was investigated. It was found that with addition of 0.5wt.% Bi, the morphology of primary Mg2Si precipitates was altered from coarse dendritic to polygonal shape, and their mean size was significantly decreased from more than 20 µm to about 8 µm. Moreover, the rod-like eutectic Mg2Si in Bi modified Mg-Si alloy solidified much finer than that of the base alloy. Such improved microstructure of the modified alloy resulted in 10% tensile strength and 20% elongation improvement. The isotropic growth manner of primary Mg2Si particles in modified alloy was attributed to the poisoning of growth in preferred directions by dissolving Bi atoms in Mg2Si lattice. The reduction in size of eutectic Mg2Si was found to be the result of Mg3Bi2 compound formation around eutectic Mg2Si rods.

Oct 1, 2011

By T. Watanabe, M. Nakagawa, K. Kimapong

Ultrasonic wave was irradiated on metal surface in water to introduce compressive residual stress. The ultrasonic wave that was oscillated with 19.5 kHz ultrasonic transducer and amplified with a step type horn was irradiated. Slight plastic deformation occurred at the surface region of pure metal such as Cu and compressive residual stress was introduced. It seemed that the plastically deformed depth from the metal surface corresponded approximately to the depth of introduced residual stress. The higher compressive residual stress could be introduced into the surface of Cu-Zn brass and SUS304 stainless steel because work-hardening was simultaneously induced to those alloys. Furthermore, since martensitic transformation was induced by the plastic deformation in the case of SUS304, the compressive residual stress could be introduced more deeply into the surface region. The residual stress of SUS304 reached to the maximum value of 574 MPa by irradiating for 5min.

Jan 1, 2004

By A. Fyson, M. P. Smith, Margarete Kalin

"Drums containing 0.07 m3 of pyritic waste rock from a zinc/copper mine, with and without Natural Phosphate Rock (NPR) additions (40 kg/t waste rock) were set up outdoors. The effluent chemistry has been monitored over a period of 709 days. In the presence of NPR, the cumulative acidity discharged from 7 out of 8 drums was 80 % less than that discharged from drums without NPR. Mixing NPR throughout the drums was more effective in reducing acidity than placement of the NPR in a layer in the drum. NPR treatment was more effective for 'fresh' waste rock than for material which had weathered for more than 4 years. It is concluded that application of NPR to pyritic waste rock could greatly reduce acidity generated.IntroductionWaste rock dumps are significant sources of acid, generated by the oxidation of sulphides. Oxygen and water supply are oxidation rate-controlling factors, and oxygen supply is the reaction-limiting component. If secondary minerals can be formed on the surface of the rocks, pyrite oxidation and therefore acid generation may be reduced or inhibited. Neutralizing agents such as limestone do not prevent the oxidation of pyrite but treat the resultant acid mine drainage (AMD) or acid rock drainage (ARD).Meek et al. (1991) determined in pilot scale experiments that apatite mixed with tailings was more cost effective than liming or use of PVC liners to reduce oxygen and water penetration into the tailings. Phosphate is not acting as a neutralising agent but it creates a chemical condition on pyritic surfaces which prevents reactions of the oxidising agents. Precipitation of principally iron phosphates physically isolates the pyrite surfaces from further oxidation (Ziemkiewicz 1990; Evangelou 1994). The chemistry of the phosphate effects has been described by Spotts and Dollhopf (1990). Secondary mineral formation was evident in the leach pile at Gibraltar when leaching was impaired (Scott, 1991) indicating a link between oxidation and the mineral-formation process. This paper reports on effluent characteristics when waste rock is generating acid in the presence of NPR."

Jan 1, 1995

By Lin Li, Ahmad Ghahremaninezhad

"In this paper, the electrochemical behavior of ferric and ferrous couple on the pyrite surface was studied by application of the chronoamperometry technique in an acidic solution at room temperatures and atmospheric pressure. The effects of the pyrite origin, ferric and ferrous ions concentration on the electron transfer kinetics of Fe3+/Fe2+ reaction were probed. Tafel behavior of all the test conditions were developed. The results showed that the ferric-ferrous reaction on pyrite electrodes with different origins have a common mechanism, despite different reaction kinetics. The exchange current densities of the redox reaction have the same magnitude of about 10-5 A·cm-2, which is far lower than the value reported for the gold and platinum electrodes, but larger than that of the chalcopyrite electrode. The cathodic transfer coefficients are between 0.4 and 0.5, while the anodic transfer coefficients appear unusual with the lowest value being 0.23. It was also demonstrated that an increase of one order of magnitude of iron concentration has no clear effect on the exchange current density and cathodic transfer coefficients of Fe3+/Fe2+ redox reaction.INTRODUCTIONPyrite, FeS2, is mostly considered as a gangue mineral which frequently coexists with valuable minerals and metals, such as gold, copper and nickel. Its separation from valuable minerals can be a costly process (Chandra & Gerson, 2010). In refractory gold ores, the fine gold particles are locked inside the pyrite structure. Oxidative leaching of pyrite is a hydrometallurgical method to destroy the structure of pyrite and expose the fine gold particles to the cyanide solution in the downstream cyanidation circuit. Pyrite oxidative leaching can be achieved by oxidation of pyrite via oxygen and ferric ion(McKibben & Barnes, 1986). In fact, pyrite oxidation by oxygen and ferric is a significant source of acid mine drainage(AMD), which adversely causes serious environmental problems(V. P. Evangelou, 1995).The pyrite oxidation mechanism is accomplished in three consecutive steps: (i) pyrite is oxidized by oxygen and ferrous ion formation (Eq. 1); (ii) ferrous oxidation by oxygen and ferric ion formation (Eq. 2); (iii) pyrite oxidation by means of ferric ions (Eq. 3). Hence, both oxygen and ferric participate in the pyrite oxidation process (Moses & Herman, 1991):"

Jan 1, 2016

By M. J. Baker

Energy and Minerals Australia Ltd are undertaking a study of the recovery of uranium from the Mulga Rock Deposits in Western Australia. Approximately 80-90% of uranium identified to date is present as tabular layer up to 5 m thick within a lignite seam 30 m to 50 m below surface. Other metals are also present. Metallurgical studies at ANSTO Minerals laboratories in Sydney have focused on identifying a suitable uranium leaching process and assessing the possible downstream impacts of the organic material in the ore. The extraction of other metals has also been examined. This paper provides the results of testwork performed at ANSTO to date.

Aug 1, 2010

By Richard C. Williams, Herman Pieterse, Ken Thomas, K. S. Fraser, Ron H. Colquhoun

"The Banick Goldstrike Mines Inc. (Goldstrike) operation is located in Eureka County in north-central Nevada on the Carlin geological trend. Gold in the sulphide mill feed, occurs primarily as submicroscopic inclusions or possibly solid solution in a pyrite and marcasite matrix. Sulphide mill feed is pretreated in a pressure oxidation plant allowing gold recoveries of 90 per cent and greater to be achieved. Anticipated 1994 gold production for the sulphide gold facility is approximately 1,500,000 ounces.In January 1989 American Banick Resources Corporation (Banick) announced the US$365 million Betze Development Plan for the Goldstrike property. The Betze plan was based on a three (3) staged expansion of the original 4,500 STPD oxide gold plant. This expansion would involve addition of ore and gold processing capacity and the total conversion of all ore handling facilities to treating Betze sulphide ores.Phase 1 of the Betze Development Plan involved the construction of a pressure oxidation circuit, including a single autoclave and supporting ancillary services. The pressure oxidation circuit was commissioned in February 1990. Currently a throughput of 2,000 STPD is being achieved.Phase II of the Betze Development Plan, commissioned throughout April and May 1992, included the expansion of grinding, leaching, and gold recovery facilities enabling the treatment of combined oxide, and sulphide ore. Pressure oxidation capacity was increased by the addition of two (2) autoclave circuits each treating 2,600 STPD of sulphide ore and other required ancillary services. By June 1992, the Phase II gold facility had achieved a combined throughput of 15,000 STPD of oxide and sulphide ore.Phase Ill, the final stage of the Betze Development Plan, was successfully commissioned in February and March 1993. This final stage of development increased the design throughput of the pressure oxidation plant to 15,000 STPD by the addition of three (3) autoclave circuits with combined capacity of 7,800 STPD."

Jan 1, 1994

By A. L. Rutledge, R. S. Nichols

"Waste dump development at the Fording River Operation s in British Columbia began in 1971. In 1981, several unique as well as unplanned situations provided an opportunity to study water flows through waste dumps.The key to the drainage is naturally segregated rock sizes .created by end dumping from the crest of a pile. The largest size material accumulates at the toe. Flow rates along the base of the studied dumps varied from 2,800 to 13,500 m3/day or 0.032 to 0. 156 m3/s (1.2-5.5 cfs}. Sediment concentrations in the water ranged from 1 to 1,800 mg/I.This paper describes water flows through dumps of varied lithology and maturity. The results suggest that neither long-term water flow nor contained sediments in small streams are jeopardized when crossed by end-dumped fills.IntroductionThe Fording River Operations are located in southeastern British Columbia approximately 115 km north of the United State s and 10 km west of Alberta, within the medial range of the Rocky Mountains. All production is carried out at elevations ranging between 1,600 and 2,200 m above sea level.The property is within a continental climatic zone, with temperature extremes of - 40°C in January to +35°C in July. The average annual precipitation of 335 mm of rainfall and 390 cm of snowfall feeds the numerous seasonal, small drainage courses along the Fording River valley.Since start-up in 1971, 149 million bank cubic met res of waste from truck-shovel and dragline mining have been placed in numerous dumps or ""spoils"". The dumps were developed to the point that several unique situations could be studied. During the peak run-off period of May-June 1981, water flow studies were performed on dumps of varied lithology and maturity. This paper describes the results of the study."

Jan 1, 1982

By L. A. Nazarova, M. Vandamme, L. A. Nazarov, R. I. Rodin

"The paper describes a new method of interpreting the “canister test” data for the quantitative estimation of gas-kinetic parameters of coal bed based on the developed model of gas emission from coal sample placed in a sealed container. The model accounts for grain-size composition of the sample, initial gas content S, as well as coefficients of diffusion D and mass transfer ?. The inverse coefficient problem is formulated to find S, D and ??from the data of pressure measurements in the sealed container. It is shown that the inverse problem is uniquely solvable if the pressure is dropped during the experiment. The method has been tested using the data from an underground coal mine in the Kuznetsk Basin.Based on the poroelastic model of a coal bed and Khristianovich’s concept about origination and evolution of the filtration zone F in the neighborhood of a degassing borehole the authors considered the corresponding direct problem on rock mass geomechanical state. It was offered interpretation method for data on pressure measurements in a plugged degassing borehole allows evaluation of the diffusion coefficient, the horizontal component of natural stress field and parameters of F. The related inverse problem of the mixed type is formulated, and the cost function is constructed. The analysis using synthetic input data has shown that the cost function has a number of local minimums arranged on the same straight line, one of the minimums being the inverse problem solution. INTRODUCTIONPre-mine degassing is an integral part of coal mining. It allows mitigation of induced accidents in mines (coal and gas outbursts, rock bursts) (Guide, 1990; Malyshev & Airuni, 1999; Seidle, 2011) and also furnishes an alternative and ecology-friendly energy source—methane (Rice, 1997; Nambo, 2001). Choosing an optimized pattern of degassing boreholes, as well as estimating degassing time and amount of gas recovered requires data on gas-kinetics indices of coal beds out of which the key parameter is a gas content S. According to the latest guidelines for coal mines in Russia (Ruban et al, 2012), mining is allowed if S in a coal bed section is under 7 kg/m3."

Jan 1, 2015

By A. R. Byers

"Porcupine Peninsular mine is at the extreme southwestern end of North peninsula in Night Hawk lake, Cody township, Ontario. It lies approximately 10 miles east of the gold-producing district of Porcupine.Gold in quartz veins was discovered on Gold -island in Night Hawk lake in 1906, the same year in which the now famous Porcupine gold camp was discovered farther west. Most of the claims now held by the Company were originally staked by Charles Auer in 1907, In 1910 when the level of the lake was exceptionally low, gold was discovered in a quartz vein near the shore and close to the present mine shaft.The mine is developed by a main shaft to a depth of 440 feet, with levels at 80, 180, 300, and 425 feet vertical depth and by a winze to 630 feet. A mill of 200 tons daily capacity was built during 1923, Production commenced on January 4, 1924, and continued without interruption to May 31; 1926, when operations were suspended due to a decrease in the grade of ore. Total production amounted to 27,416 ounces of gold from 99,708 tons of ore, an average grade of 0.275 oz. a ton. Diamond drilling and re-sampling of the mine during 1934 and 1935 indicated approximately 1,480,000 tons, having an average grade of 0.123 oz. gold a ton, above the 625-foot level."

Jan 1, 1954

By Bill Charbonneau

Lumen, lm The measurement unit of light emitted by a light source 60W incandescent840 lumens 32W T8 linear fluorescent3200 lumens 54W T5HO linear fluorescent5000 lumens 400W metal halide24,000 lumens

Oct 1, 2010

By F. Parada Luna, F. Parada Torres, J. Salvo, A. Reghezza

In the pyrometallurgical processes of copper concentrates, arsenic is preferentially concentrated in the dusts carried by the gases as a result of their volatility and subsequent condensation in the gas cleaning train. In this work results obtained by alkaline leaching of the dusts from Ministro Hales roasting plant, belonging to Codelco, are presented. For the Ministro Hales plant the alkaline leaching would be an alternative to the recirculation of dusts to the roasting furnace. The experimental study considered NaOH and NaOH plus NaHS leaching of two samples of dusts, with 4.24 and 2.55% arsenic respectively. The results show that it is feasible to extract 90% of the arsenic and 50% of the antimony at room temperature using only NaOH, and when the temperature increases to 95ºC, the extractions of As and Sb increase. Copper remains in the solid residue with less than 1% dissolution, and the NaOH net consumption was approximately 140 kg/t of dust. In leaching tests with NaOH + NaHS, two types of methods were tested, baking plus leaching with water and direct leaching with a NaOH+NaHS solution. With the second method arsenic extraction increased to 95% and copper dissolution was less than 0.01% at 85ºC, NaOH 100 g/L, NaHS 90 g/L, 1.5 h and L/S ratio of 5/1 (L/kg). The bismuth remains stable in the solid in all cases. INTRODUCTION The presence of arsenic in copper concentrates has increased with time and the environmental regulations are becoming increasingly more stringent. In concentrate treatment pyrometallurgical plants, most of the arsenic is volatilized condensing in the gas train and is collected along with the entrained solids to form the flue dusts. Under these conditions, in smelters of copper concentrates with high arsenic content, the dusts are being treated in hydrometallurgical plants to extract and stabilize the arsenic, which also means that most of the copper contained in the dust is co-extracted with the arsenic. This affects the recovery of copper in the smelter and an alternative such as the selective alkaline leaching of arsenic, would allow the copper from flue dust to be recovered.

Jan 1, 2019

By Tony B. Szwilski

"The efficiency of operation and life of an advancing longwall coalface depends, to a great extent. on the face-end roof strata conditions and the stability of the gateroads. The factors that influence the rate of closure of gateroads are examined, in particular the gateside pack and the coal rib (crush) pillars. The performance of mechanized packing systems is briefly discussed. Gateroad deformation surveys were carried out at two coal mines to determine the influence of the rib pillar width on gateroad stability; also, the influence of the gateside pack on the immediate roof and floor strata is analyzed based on strata bed displacement measurements from instruments installed in the gateroad and pack-hole of a longwall advancing face operating in a thick seam (7.55 m). Concluding, an evaluation of the fac e-end systems examined is given and recommendations made f or the most suitable rib pillar width. face-end geometry and roof strata reinforcement.IntroductionKevword s: Coal mining, Pillars, Rib pillars, Ground control, Longwall mining, Gateroads , Packing systems, Roof control. The face-end region of a longwall advancing face is the center of mining activity, being the turn-around point for the coal cutting machine. and the place of gateroad formation and movement of special face-end supports. The design of face-end layouts can be aided by knowledge of strata control parameters, such as the effective loading of the immediate roof strata , convergence of the roof and floor, magnitude and distribution of fractures in the roof strata around the face-end , optimum placement of supports, coal seam thickness and strength characteristics, geometry of layout and interaction from neighbouring workings. Generally, the gateroad is formed by one of three face-end methods: (a) advanced headings, (b) in-line system (gateroad formed at the face) or (c) conventional ripping with or without stable elimination (also halfheading). The quality of gateroad is influenced principally by the face-end system employed, standard of gateside pack and width of pillar. A good standard of gateroad is generally achieved by employing the half-heading face-end system, the gateside pack and coal rib pillar having comparative support/ yield (stiffness factor) characteristics. However, the halfheading is often not favoured due to face operation restrictions."

Jan 1, 1981

By J. Lu, M. Mohammadi, C. Pelletier, B. Wassink

Chemical reduction of selenate by chromous ions is proposed as a novel method to remove selenate from sulfate-bearing solutions Chromous ions were generated by electrochemical reduction of chromic ions using an electrochemical cell. The investigation was conducted in sulfate medium and under various concentrations of chromous ions (0.0062 to 0.05M), pH 1 to 2.2 and temperature 10 to 60°C. Selenate ions were rapidly reduced by chromous ions to hydrogen selenide. The overall reaction at a chromous/selenate molar ratio of 8 or higher can be expressed by the following reaction. At an initial chromous/selenate molar ratio below 8, the products are hydrogen selenide and elemental selenium. At a lower chromous/selenate molar ratio and pH, more elemental selenium and less hydrogen selenide are produced. The reduction kinetics is controlled by both chromous ion and hydrogen ions concentrations. The reaction rate is first order with respect to selenate, hydrogen ion and chromous ion concentrations. It was found that the selenate reduction rate with chromous is independent of the concentration of sulfate present in the solution. The general rate equation was proposed and the rate constant was calculated as k= 0.38 M-2s-1. An activation energy of 58.7 kJ/mol was obtained in the temperature range from 10 to 50°C. Nearly 100% of selenate can be reduced to H2Se from solution containing 17 ppm selenate in less than 30 minutes at pH 2.2 and 60°C. Various strategies have been tested to remove the produced H2Se from solution. INTRODUCTION Selenium in nature is mostly associated with metal sulfide minerals such as copper, iron, zinc, and lead sulfides (Zingaro & Cooper, 1974). The extraction of metals from ores has the potential for selenium contamination of ground and surface water which can pose an environmental and health concern. Therefore, selenium contamination in the receiving environment is a key issue for many industries (specifically copper industry) and the treatment of wastewater contaminated with selenium is a growing challenge. The effective removal of selenium from contaminated water is very important to support sustainability in the copper industry and the other industries with selenium pollution issues.

Jan 1, 2019

In the last 30 years numerous mining plants have been developed in North America but only a handful of these have required freezing the area where the shaft was sunk. The most recent mine shaft to be sunk using this technique is located on an island off the coast of Louisiana. The project required the input and skills of numerous people in several companies, including the owner, engineers and contractors. The Cote Blanche installation is one of several mines in the area which extract salt from a dome which extends under the Gulf of Mexico. The geology of this area is, however, markedly different from others in the region because there is approximately 460 ft of water logged soil above the salt dome. Because it is not possible to sink a shaft in ground that is this soft, the overburden where the shaft is located must be frozen before excavation can take place. The Cote Blanche mine is a producing facility which was built in 1958 by the Carey Salt Company. The original facility included a 14 ft diameter production shaft and a 7 ft-8 in diameter service/ventilation/escape shaft. This system was installed by The Cementation Company. In 1996 the owner decided to improve both the mine ventilation and production hoisting system by installing a third shaft. This shaft would be designed to extract 650 tons per hour from a depth of 1600 ft and provide 500,000 cfm of ventilation to the mine at a maxium design pressure of 4 in WG. These parameters resulted in a requirement for a 16 ft finished inside diameter concrete lined shaft and a headframe designed as a pressure bulkhead. Because the mine would operate from one level only, the most efficient hoisting system for the application was determined to be a ground mounted koepe running on rope guides. With these criteria fixed Cote Blanche embarked on a program which would see construction start in June 1998 with the first salt produced in February 2001 at a cost of $US18M. The major players in the design and construction were G.L. Tiley & Associates Ltd., BLM Bharti Engineering, Freezewall Inc., Phillips Mining, Geotechnical and Grouting Inc., Don Hart Consulting and Dynatec International.

May 1, 2002

By M. K. Walter

"The underground operations of Kidd Creek Mines uses an average of 450 U.S. gpm water during regular operations, and discharges an average of 800 U.S. gpm through consecutive staged pumping lifts of 1200f t.Water is used as a flushing medium for the drilling operations and as a dust suppressant for drawpoint mucking. The water retains an average of 5% suspended solids by weight which is removed underground before being pumped to the surface. This paper outlines the treatment of the mine water for the removal of suspended solids which are chemically treated to settle out as a sludge product of 60% solids by weight. This process reduces the sludge handling operation by 50%. The sludge is systematically disposed into the ore passes.IntroductionThe Kidd Creek minesite, located 27 km north of Timmins, Ontario, produces 4.5 million tonne s of copper, lead, zinc and silver ores annually from its No. I and No.2 underground mining operations. Both mines use a sublevel blasthole stoping method of extraction.Water from the underground operation passes through a series of surface settling ponds, where it is treated to remove heavy metal ions from solution. A blend of the pond effluent with fresh water from Kidd Creek, supply the underground with water at an average rate of 450 U.S. gpm. Water flows from the production zones through coarse sumps and into a main pumping station. There are five main pumping stations underground, three in No. I Mine and two in the No.2 Mine.The No. 2 Mine pumping stations are located on the 4800 level and the 3600 level. All water below the 3600 level to the shaft bottom is controlled from the 4800 pumping station. The water from the 4800 pumping section is combined with water from the remaining levels of the No.2 Mine at the 3600 pumping station. Water from the 3600 pumping station is pumped to the 2400 pumping station in No. I Mine.The No. I Mine pumping stations are located on 2800, 2400, and 1200 levels. All water below the 2400 level to the shaft bottom is controlled through the 2800 pumping station. This water is pumped to the 2400 pumping station. The 2400 pumping station also controls the water below 1600level to 2400 level."

Jan 1, 1988

"An atmospheric process for the synthesis of scorodite(DMSP®)has beendeveloped,in which ferrous ions were oxidized by oxygen gas in the presence of pentavalent arsenic ions. Under optimum conditions, the synthesized scorodite (FeAsO4?2H2O) was well crystallizedand the crystals were faceted; the particle sizes were as large as 15 to 20 µm. In thisstudy, iron (II) sulfate heptahydrate (FeSO4·7H2O), arsenic (V) acid (60mass%H3AsO4) solution, and distilled water were used to prepare the original experimental solution (700 ml), in which the contents of Fe(II) and As(V) were 55.9 g/L and 50.0 g/L, respectively. The effects of various synthesis conditions, such as the type of oxidizing gas (oxygen or air), gas injectionrate, temperature, and the initial pH of the solution, on the size and shape of the scorodite particles wereinvestigated. From the experimental results, the changes in the shape of the scorodite particles wereclarified for each synthesis condition.INTRODUCTIONDOWA has developed a processfor synthesizing well-crystallized scorodite (FeAsO4•2H2O). Theprocess is named andtrademarked as DMSP® (Fujita et al., 2008; Fujita et al., 2008; Fujita et al., 2009; Fujita et al., 2009; Fujita et al., 2009, Kubo et al., 2010). In this process, oxygen gas is injectedinto a sulfuric acid solution containing large amountsof As(V) and Fe(II). Facetedscorodite crystals as large as 15–20 µm weregeneratedthrough the pre-formation of a gel-like precursor. However, it is known that the synthesized scorodite particles become smaller and more aggregatedwhen the synthesis conditionsdeviate from the optimumconditions.In this study, the effects of various synthesis conditions, such as the type of oxidizing gas (oxygen or air), gas flowrate, temperature, and initial pH of the solution, on the size and shape of the scorodite particles were investigated. The experimental resultsclarified the changes in the shape of the scorodite particles for each synthesis condition."

Jan 1, 2016

By D. E. M. Plaumann, B. S. Fee, R. D. Hansen, R. R. Klimpel

"This paper briefly summaries the status of a recent program aimed at identifying new chemistry for use as sulfide mineral frothers and collectors. The background motivation and planning behind this work is given along with the general families of chemical structures synthesied and studied in detail followed by typical flotation application results in various laboratory and plant scale tests. Flotation testing to date has been very encouraging with often improved recoveries of certain minerals and improved pyrite selectivity being observed over commercially available reagents. In addition, in some cases, the dosage and lime requirements of these new collectors are significantly less than that associated with standard sulfide mineral collectors. INTRODUCTIONFroth flotation is one of the most widely used and economic means of concentrating metal sulfide ores. There are numerous reference books that have been published on the chemical aspects of flotation [e.g., ref. 1-7] which are based on studying the influence on selected minerals of chemical structures that have been used commercially for many years. In fact, almost 98% of the volume of flotation collector and frother reagents used at the industrial level today were known and manufactured in some form 25 years ago [8]. There are many reasons for this lack of significant new flotation reagent development [9] including the general utility and relative cheapness of existing collectors such as xanthates and existing frothers such as the alcohols and glycol ethers. Another reason is the inherent complexity of the flotation process itself coupled with highly variable and difficult to standardize feed materials. This environment makes reliable reagent comparisons difficult and seriously limits the use of theoretically based chemical structure studies to predict new or optimal reagents. In addition, the higher cost of specialty flotation reagents demands that their performance attributes be significantly greater than that achieved with existing all-purpose commercial reagents - something often difficult to prove in practice."

Jan 1, 1988