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By C. F. Harris

Following the initial development of the zinc-lead blast furnace, efforts have been directed both at increasing the scope of the process and increasing the cash return. The paper reviews the process developments which enable the process to smelt efficiently a range of material which encompasses high-grade zinc sulphide concentrates. mixed concentrates, Waelz oxide, zinc ashes and other materials blended to give a furnace feed ranging from 25 to 65% zinc and from 1 to 25% lead, with a copper-lead ratio exceeding 1 to 8. The paper also reviews the process developments which have enabled the capacity of a furnace and condenser system of standard size to be boosted from 30,000 to 80,000 tonnes/annum. The benefits achieved are used as a justification for continuation of the development of the zinc-lead blast furnace. Predictions are made of the benefits expected from further developments.

Jan 1, 1979

By Y. G. Yaroshenko, V. Y. Baldin, Y. M. Gordon

The manufacture of sulphuric acid from the off-gas of pyrometallurgical processes in the production of heavy metals is one of the common methods of effective recycling of sulfuric anhydrite. The new Impact-Inertial Devices with the rectangular contact channels (IIRCh) and the internal circulation of liquid allow the injection of an absorber into the phase's interactive cell by utilizing the gas stream energy and by this way increasing the sulphuric acid yield and reducing the power consumption. These devices have the following advantages: simplicity of design, smaller size, low capital expenses, absence of irrigating system, high operational reliability, simultaneous cleaning of solid, gaseous and liquid impurities and also cooling of gases etc. The IIRCh efficiency is almost 18 % higher and they require 1.3-2 times lower energy when compared with Venturi Absorbers. Also the elevation of spraying nozzles or injectors is 1.5 times lower in the IIRCh case.

Jan 1, 2004

By H. Ito, T. Takasu

Effects of a mechnochemical treatment to oxide Ni ores have been investigated to enhance the extractive ratio of Ni from them in acid leaching process. High intensity grinding by a planetary boll mill showed strong effect to enhance the extractive ratio of Ni, for example , 10% of extractive ratio of Ni in sulfuric acid leaching(0.05mo1/1,at room temperature) increased to more than 90% with 60min. of planetary boll milling. Effects of additives(sulfur and some oxidants) on the mechnochemical treatment have also studied. Sulfur showed almost no effect to enhance the total Ni extractive ratio but increases the selectivity of Ni extraction. More than a half of Ni could be extracted by water when Ni ore was grinded at addition of Na2S207 in the milling treatment. Some effects of mechnocemical treatments to Ni ore are summarized in this paper.

Jan 1, 2005

By Fathi Habashi

"On August 3, 1888, German Patent No. 43977 entitled ""A Process for the Production of Aluminum Hydroxide"" was issued*. The discovery which led to the patent was made by the Austrian chemist Karl Josef Bayer (1847-1904) (Fig. 1) who was at that time in Russia, and the process became known as the Bayer Process in his honour(1)(2)- The process imediately achieved industrial success, displacing the pyrometallurgical process that had been used until that time to produce alumina. The Bayer Process involved the pressure leaching of bauxite with NaOH solution to obtain sodium aluminate solution from which aluminum hydroxide was precipitated by seeding. Today bauxite is not only the principal source of aluminum but also of gallium, both metals being effectively recovered by the Bayer Process. The purpose of this paper is to outline the historical background of this discovery and its importance to hydrometallurgy."

Jan 1, 1999

By W. Curlook, L. M. Pidgeon

THE PURPOSE of this investigation was to study part of the ternary system Co-Fe-S. Jensen ( 1) established the solubility limits of sulphur in pyrrhotite. The compound with 61.7 per cent Fe, corresponding to a formula Fe12S13, was found to melt congruently at l,192°C. Hulsmann and Weibke(2) determined the Co-S phase diagram from O to 35.2 per cent S. The solubility limits of sulphur in the com-pound Co(l-x)S were not determined. The initial object of the present in-vestigation was to determine the composition at which the compound Coci-xiS melted congruently. Once this was ascertained, it was decided to investigate the ternary system Co-Fe-S from the region of 100 per cent Co(1-x) S to the region of 100 per cent Fe12S13.

Jan 1, 1953

By J. Laganiere

A considerable part of the sulphuric acid requirements of the Aluminum Company of Canada's Arvida smelter comes from the burning of 180-200 tons per day of zinc '"'sulphide concentrates in a 22-ft.-diam-eter fluidized bed roaster. The plant and its operation were fully de-scribed in a publication that appeared about ten year ago (1). A simplified flowsheet of the roasting and gas-cleaning sections of the plant is given in Figure 1. The wet zinc sulphide concentrates are un-loaded into a pug mixer, by means of a Pay loader, and mixed with water in a mixing tank to make a slurry of 80 per cent solids. The slurry is transferred to a .feed tank and then introduced into the roaster at a point four feet above the constriction plate, which is the height of the settled bed of zinc calcines. The roasting temperature is con-

Jan 1, 1963

By R. J. A. Fricker

IT is the purpose of this essay to give a reasonably complete description of shaft sinking under a rock pentice in a small Northern Ontario gold property. Hallnor is in many ways an exceptional mine, but the sinking operations here described are but typical examples of the excellent practice for which Canadian mines and mining engineers are justly famous the world over. The worth of the subject lies, therefore, nor in its originality, but in its soundness; not in the spectacular, but in the economical and safe. Attention is particularly drawn to the excellent safety record and to the exceptionally low overall sinking cost. INTRODUCTION Prior to April, 1939, the Hallnor No. 1 shaft had been sunk vertically to a depth of 997 feet below the surface. It is a three-compartment shaft, 19 ft. 10 in. by 7ft. 7 in. outside the timbers (see Figure 1).

Jan 1, 1941

By Randy McMeekin, Chris Twigge-Molecey, J. Peter Blake

The business objectives of a metallurgical project determine targets and constraints, such as the range of products, production capacities, in-service dates, and capital and operating cost parameters. Permitting is usually left to the project team, as are decisions regarding site location, technology, and construction approach. Success requires an orderly methodology, represented in this book by front end loading (FEL) phases: FEL1 examines several alternative approaches; FEL2 ranks and eliminates less economically or technically attractive options; and FEL3 develops the preferred option to the point of major permit completion, with defined scope, schedule, and cost to submit for approval to proceed with implementation in FEL4. Working through FEL1, FEL2, and FEL3, the project definition improves, estimates become more accurate, and project risks are reduced. The project design basis at the end of FEL3 is the control document for project implementation in FEL4.

Jan 1, 2015

By Z. -E. Nakagawa, O. Yamamoto

After the formation of a concentration gradient of SiC within C-C composite, either zircon or mullite thin films were coated on the surface by sol-gel process. Oxidation resistance tests of as-prepared composites with a couple consisting of a ceramic thin film and a SiC-gradient were then carried out in air at the temperature range from 600 to 1600°C. By the coating of zircon with a thickness of 1.5 µm, a pronounced oxidation resistance was observed; the mass loss by oxidation was almost negligible at 1000°C and one hundredth of original composite at 1400°C. In the case of the coating of mullite with a thickness of 1.0 pm, the marked improvement in oxidation resistance over the whole temperature range examined, from 600 to 1600°C, was found with the combination of a SiC-gradient and a mullite thin film

Jan 1, 2004

By Pingchao Ke, Zhihong Liu

"To overcome the problem that scoroditeis not stable in neutral and alkaline solutions, we proposed and investigated a synthesis method of scorodite coated in-situ with hydrated ferric oxide in the Fe(II)-As(V)-H2O system at the initial Fe(III)/As(V) molar ratios larger than 1.0.The synthesis was carried out at 90?,pH 1.5 under atmospheric pressure. During the synthesis, oxygen gas was conducted and bubbled into the system to oxidize Fe(II) ions into Fe(III)ions, and Na2CO3aqueous solution as neutralizing agent was added controllably into the system for keeping the pH constant at 1.5. In this study, three samples of scorodite were synthesized at the controlled initial Fe(II)/As(V) molar ratios of 1.0, 2.0 and 3.0, respectively. The characterizations of the samples by HRTEM, TEM/EDX, and chemical composition indicated that the particle surfaces of scorodite synthesized at the initial Fe(II)/As(V) molar ratio of 2.0 and 3.0 were coated in situ by a film of hydrated ferric oxide with the thickness of several nanometers, while the sample synthesized at the initial Fe(II)/As(V) molar ratio of 1.0 was uncoated scorodite. The TCLP and long-term leaching for the toxicity characteristic of the samples showed that scorodite coated in situ with hydrated ferric oxide was stable in a wide pH range from 3.0 to 10.0, with the arsenic concentrations in the leaching solutions less than 1.0mg/L. Nevertheless, uncoated scorodite was unstable under the conditions from weak acidity to alkalinity, with the arsenic concentrations in the leaching solutions larger than 5.0 mg/L.INTRODUCTIONArsenic pollution is a challenge for environment and health and has been an important issue for scientists (Hopkin, 1989).Arsenicis commonly associated with ores or concentrates of nonferrous metals, mainly as independent arsenic bearing minerals such as arsenopyrite (FeAsS), tennantite(Cu12As4S13), enargite (Cu3AsS4), cobaltite (CoAsS), and gersdorffite (NiAsS), realgar (As2S2), orpiment(As2S3), etc.(Matschullat, 2000; Mandal & Suzuki, 2002; Long & Peng, 2012). In the nonferrous metallurgical processes, arsenic is enriched mainly in such arsenic bearing materials as various flue dusts, residues, and anode slimes, which results in the accumulation of arsenic in the system owing to the circulation of those arsenic-bearing materials back to the system and the pollution to the environment (Harris, et al, 2003).In addition, the utilization of arsenic is limited and has been decreasing owing to its toxicity and carcinogenicity. Therefore, arsenic has to be removed from those arsenic-bearing materials and discarded safely (Dutrizac & Jambor, 1988; Monhemius & Swash, 1999; Drahota & Fillippi, 2009).In this case, how to immobilize arsenic has become a difficult problem which must be solved. Among the methods of immobilizing arsenic which have been proposed and investigated, immobilizing arsenic into a stable mineral to fix it, has been realized as a promising method and becomes a research focus(Dutrizac et al, 1987; Dutrizac & Jambor, 1988; Demopoulos et al, 1995; Fujita et al, 2008a; 2008b).Scorodite has been known as one of the best minerals for immobilizing arsenic, owing to its high stability under weak acidic condition, the arsenic content as high as 32wt%,and good crystallinity benefiting the separation of solid to liquid(Daivid et al, 1999; Zhu & Merkel, 2001). Scorodite can be synthesized from arsenic containing solutions under both hydrothermal and atmospheric conditions, and a lot of work has been done on the synthesis of scorodite and its leaching stability characterization (Filippou & Demopoulos, 1999; Gomez et al, 2011).Two plants for the immobilization of arsenic as scorodite have been putin to operation in Chile and Japan, respectively (Filippou & Demopoulos, 1999; Amir et al, 2016)."

Jan 1, 2017

By V. I. Lakshmanan

Process Research ORTECH Inc. (PRO), with its in-depth knowledge on cobalt behaviour in nickel matte smelters and its novel mixed chloride hydrometallurgical processing, is developing a hybrid process to enhance cobalt recovery. PRO’s experimental data is compared with plant information on cobalt behaviour in smelters. In final stages of iron rejection in the converters, the thermodynamics of the matte-slag system dictates the undesirable transfer of the cobalt to the final converter slags. This results in much lower cobalt recovery in the product Copper/Nickel matte (<60%) in present copper and nickel smelters. This information is used to transfer cobalt into desired quantity of converter slag. The enriched cobalt slag is subjected to PRO mixed chloride process to recover nickel, and cobalt as high value salts of importance in lithium ion batteries. From a typical converter slag containing 3.5% Ni, 1.5% Co, and 1.0% Cu, PRO leach process extracts 98 % Ni, and 98 % Co, and 99% Cu. Application of this process to enhance cobalt recovery in typical Copper/Nickel smelters is presented. In addition, this will also enhance the Copper and Nickel recovery. Both these aspects are of interest to the Copper, Nickel, and cobalt producers. INTRODUCTION In pyrometallurgical extraction of copper and nickel from sulphide ores cobalt often occurs with copper and nickel. The requirement of cobalt is increasing due to its need in lithium ion cells used in electric vehicles. Therefore its efficient and economical extraction in a form suitable for use in lithium batteries is of great interest at present. In the pyrometallurgical processing of copper and nickel sulphide concentrates the concentrate is smelted to produce a furnace matte, which is then sent for converting to remove the iron to produce iron free (~1% Fe) matte for further processing to copper or nickel matte. The copper/nickel matte is processed for metals recovery by hydrometallurgical or vapo-metallurgical or oxidation – reduction routes to make market products. As shown in Figures 1 and 2 the slags are produced in the smelting step and in the converting steps. The metal losses in smelting are controlled by the oxygen pressure in that step and slag production which is controlled by concentrate grade. The slag produced in the converting step which is carried out at a higher oxygen pressure contains higher level of metals is often returned to the smelting step to increase value metals recovery.

Jan 1, 2019

By M. Kolahdoozan, H. Noori

"Molybdenite flotation at the Sarcheshmeh Copper Complex has been studied as a function of pulp potential controlled using two types of sodium sulfide (mineral and chemical origin). Air and either plant or laboratory nitrogen was used as a flotation carrier. Potential measurements were taken in the rougher cells as it was proved to be vital for molybdenite flotation. Flotation test with reducing potentials between -525 and -575 mv gave the highest molybdenum recovery (93%) and the lowest copper recovery of between 4 to 15% can be achieved. These potentials did not change considerably by adding increasing amounts of sodium sulfide. At more oxidizing potentials, i.e. more than -450 mv, the recovery of copper increases while the recovery of Molybdenite drops.INTRODUCTIONThe average copper ore feed grade to the Sarcheshmeh flotation circuit is typically set at about 1.2% Cu. to be delivered to processing unit is set at 1.2%. There is no control of molybdenum (Mo) which usually varies between 0.01 and 0.04%. During concentration, a Cu-Mo concentrate is produced by one stage of roughing, one stage of scavenging and two stages of cleaning. Final concentrate contains 0.5% of Mo and 32% of Cu while the particle size is 90% under 42 microns (325 mesh). Because of the high grade of Mo in this concentrate and economic interests, a molybdenite flotation circuit plant was built in 1984. This plant was designed to treat 1432 tons of Cu-Mo concentrate (on dry basis) and consists of one stage of roughing followed by seven stages of cleaning. After dewatering the concentrate contains 53.7% Mo which is dried and shipped overseas.The majority of molybdenite production worldwide comes as byproduct of copper porphyry ores. The Mo grade typically varies between 0.01 and 0.06% and accompanies chalcopyrite and chalcocite. Due to the hydrophobic nature of molybdenite, it can be floated along with copper sulfides. Thus the overall concentrate of Cu and Mo contains 0.2-1% of molybdenum as well as 20-42% of copper. To separate the Mo, Cu and Fe minerals, depressants are used. The most important of which are sodium sulfide, sodium hydrosulfide, Anomal D, sodium cyanide, ferropottasium and ferro-cyanide. The first three regeants have been extensively used for successfully treating porphyry ores [1]. In the molybdenum plant of the Sarcheshmeh Copper Complex only sodium sulfide has been used to date."

Jan 1, 2006

By W. G. Jewitt

THE Box mine, situated in Saskatchewan on the north shore of lake Athabasca, is unique in two respects. It is the first and, so far, the only gold mine in Saskatchewan, and it is the only gold mine in Canada to employ a large-scale forced caving system as the main mining method. Discovered in 1934 by Tom Box, an Edmonton prospector, its subsequent history, while breaking no records for speed, has been one of steady and more or less normal development. Camp and plant are situated close to the shore of Beaverlodge bay, sorne 50 miles east of the Saskatchewan-Alberta boundary and 35 miles south of the north boundary of the Province. The railhead at Waterways, Alberta, is about 220 miles to the southwest. Prince Albert lies some 450 miles to the south. The surrounding country presents a rather rugged variation of the usual Precambrian topography, with hills rising five to six hundred feet above lake Athabasca. The numerous lakes and islands, together with the wide expanse of lake Athabasca, compose an attractive scene, at least in the not over-long &apos;open water&apos; season. The winters are long but not extremely cold, judged by prairie standards. The small lakes break-up around May 20th and the &apos;freeze-up&apos; stops navigation early in October. Though apparently remote, the Box mine is fortunate in having reason-ably cheap transportation from Edmonton. The 300-mile railway from Edmonton to Waterways is the most expensive section, the rate for 1st class freight being $1.70 per hundred pounds and for 5th class freight $0.77 per hundred. Freight from Waterways to the mine, approximately 140 miles down the Athabasca river and 90 miles along lake Athabasca, is delivered by barges and tugs for 50 cents per hundred pounds. All freight is delivered during the &apos;open water&apos; shipping season, which lasts from about .Tune 7th to October 7th. In normal years, barges are loaded to 3 to 3 1/2feet draught up to September 1st. After that date, and earlier in dry years, shallow water at Waterways reduces the barge loads. Barges of 100 to 200 tons capacity handle all mining freight, including pieces up to 15 tons in weight, without undue difficulty.

Jan 1, 1940

By R. C. Gilmore, J. McMack, G. E. Bogin, J. F. Brune, C. DeRosa

"Researchers at the Colorado School of Mines are designing and building a 1:40 scaled physical model of a fully operational longwall coal mine. The purpose of the model is to minimize the occurrence of methane explosion hazards by testing various ventilation strategies and determining optimal methane sensor placement throughout the longwall face. Fluid scaling factors were analyzed using Computational Fluid Dynamics software Ansys FLUENT™ at both full and 1/40th scales. Researchers were able to identify key scaling factors and subsequently developed suitable flow parameters that would maintain full scale accuracy of flow patterns in the 1/40th model. KEYWORDS Coal, Longwall, Methane, Explosion, Scale Model, Ventilation, Fluid Dynamics, Richardson, Gob, Porous INTRODUCTION Mine ventilation systems are a key component for ensuring a safe working environment by introducing fresh air from the surface and diluting explosive gases and hazardous contaminants. This holds true for longwall coal mines where the emission of flammable methane gas from the longwall face as well as the gob can create explosive mixtures of methane and air. Therefore, an effective ventilation strategy can help mitigate the potential for accidents within a mine. Developing an effective mine ventilation system (MVS) requires a detailed understanding of the physics of gas transport and how this relates to specific geometries and obstructions in the face. Researching this phenomenon requires the development of computational fluid dynamics models capable of simulating fluid flow and methane species transport that can be validated through physical experimentation. Unique site conditions determine ventilation layouts for every coal mine. Researchers have designed a 1/40th scaled physical model of a longwall coal face and adjacent gob that includes movable shield supports and a movable shearer with rotating cutter drums that can be adjusted to a variety of ventilation conditions. The proposed work will address the need for an improved Mine-wide Atmospheric Monitoring System (AMS), and provide the ability to link existing MVS models with AMS in real time. By accurately representing the coal face, the gob area and the mining equipment, gate roads and bleeder entries, the scaled model accurately represents all relevant structures, geometry, and dynamic conditions. A Plexiglas cover allows full optical access for sensor equipment as well as visual inspection of the fluid flow and species transport."

Jan 1, 2019

By A. A. Lavrinenko, N. I. Gluhova, L. M. Sarkisova, D. V. Makarov

"The investigated sulfide Cu-Ni ores are characterized by low PGE contents and emulgated impregnations of pentlandite and palladium minerals in the main ore-forming mineral pyrrhotite, which hinders their selective separation, requiring a collective flotation of sulfides to concentrate, and hydrometallurgical processing. As revealed by studies of floatability of Pt-Cu-Ni ore in a mixture of butyl xanthate and chelating phosphorus containing reagents, the most effective additive is Aerophine 3416 (a phosphinates mixture) manufactured by Cytec. Using collective flotation with a butyl xanthate – Aerophine 3416 composite in the ratio of 1:3 has afforded to enhance the Pt, Pd, and Ni recovery by 13.6 %, 9.2 %, and 9.4 %, respectively. The optimal ratios of Aerophine 3416 to butyl collector for the collective flotation are the following: 1:1 for Pt and Pd and 3:1 for Ni. The effect of these reagents on the electrode potential and surface hydrophobicity of pyrrhotite, pentlandite, and platinum black has been investigated. The extraction UV spectrophotometry experiments demonstrate that DIBDTPI exists on the pyrrhotite surface mainly in the form of disulfide in weak acid medium and in the form of compounds with Fe on the mineral surface in minute amounts. There appears some disulfide and few amount of (DIBDTPI)2Ni on the surface after interaction of DIBDTPI and pentlanditeINTRODUCTIONXanthates are widely used in flotation of copper-nickel ores containing platinum group metals (PGM) as sulfides collectors which can concentrate the most part of PGM connected with the sulfides. Though xanthates are hardly efficient for flotation of free PGM, the latter are floated more slowly compared to the sulfides of base metals. The reasons for poor floatation of free (mineral) forms of PGM are under-researched, however the key accepted reasons are: fine disseminations of free PGM, insufficient interaction with xanthate and negative influence of blue copperas and sulfur-containing ions which are sulfides oxidation products (Penberthy et al. 1999; Shackleton et al., 2007). That’s why an additional collector, normally dialkyldithiophosphates, is used during flotation of PGM-containing copper-nickel ores. A reagent DP-4 is used for recovery of PGM at Norilsk mining and metallurgical plant. The reagent represents a slightly-soluble organic calcium sulfonate which is given to flotation pulp in the form of water emulsion of 5% - solution in automotive diesel oil."

Jan 1, 2016

By D. Warkentin, T. Yeomans, R. Tavakoli, A. Nacu

"One of the most successful commercial methods of extracting gold from ores involves solubilizing gold as a gold cyanide complex, followed by adsorption of the complex onto activated carbon. The gold cyanide is subsequently eluted from the carbon to produce a concentrated gold solution suitable for electrowinning to recover the precious metals. Stripped carbon is treated in a kiln under a flowing steam atmosphere to reactivate its surface, allowing it to be reused. The aim of this study was to investigate the gold elution using a mixture of sodium hydroxide and isopropyl alcohol in both batch and continuous tests, with the eluates being electrowon on carbon foam cathodes. The effects of various factors, including alcohol content and sodium cyanide addition were studied. The addition of alcohol considerably improved gold elution kinetics, while the presence of sodium cyanide did not have a significant impact on either elution efficiency or kinetics. In all cases, efficient electrolysis was shown to be an important factor in maintaining effective carbon circuit operation.INTRODUCTION Gold dissolved as Au(CN)2- complex during cyanidation is recovered either by cementation with zinc powder, also known as the Merrill-Crowe process, or by adsorption into activated carbon through Carbon-in-Pulp (CIP) or Carbon-in-Leach (CIL) processes. Activated carbon loaded with gold is then eluted and gold is subsequently recovered by electrochemical methods. Conventional elution processes such as atmospheric and pressurized Zadra processes, involve stripping with a cyanide (0.1–0.2% NaCN) and sodium hydroxide solution (1–2% NaOH) at high temperature (90–140°C) over long elution times of 8 to 72 hours. In the Anglo American Research Laboratories (ARL) elution process, the acid-washed loaded carbon is pretreated with a strong sodium cyanide and sodium hydroxide solution (1–5% NaCN and 1–2% NaOH), followed by elution with water at 110–130°C. Solvent-assisted elution processes have also been reported, such methods use varying concentrations of alcohols (methanol, ethanol and isopropanol), ketones, or glycols to enhance atmospheric elutions at temperatures up to 95°C (Muir et al., 1985; Ubaldini et al., 2006; Espiell et al., 2008; Soleimani & Kaghazchi, 2008). The present work investigates the efficiency of isopropanol in the gold elution process, as well as the methodology of evaluating gold stripping processes in the lab setting"

Jan 1, 2017

By J. Ambrose, R. W. Zimmerman

"Shales are highly anisotropic in their mechanical behavior. The strength ( ??1) of anisotropic shales depends not only on the magnitude of the other two principal stresses ( ??2, ??3), but also on the bedding plane orientations (?) relative to the principal stresses. In this study, the role of the intermediate stress ( ??2) on the strength of shale is evaluated using new data from triaxial compression and extension tests. To understand the failure behavior of shales, we have carried out triaxial compression tests ( ??1 > ??2 = ??3) on two organic rich mudstones, at different confining stresses and loading angles. We first fit the triaxial compression data with Jaeger’s plane-of-weakness model, and find that the model is able to represent the strength behavior of anisotropic shales reasonably well. We then carried out five triaxial extension tests ( ??1 = ??2 > ??3) for a high strength-anisotropy shale, and two extension tests for a low strength-anisotropy shale. Results from the compression and extension tests for failure along the weak plane shows that the role of ??2 is not significant for weak plane failure. However, for the low strength-anisotropy shale, the failure strength of the intact rock (i.e., at ??= 0o and 90o) for compression and extension tests were significantly different. This indicates that for intact rock failure, the role of ??2 is important and cannot be ignored.INTRODUCTIONFor many years, various researchers have studied the roles of ??2 for failure of isotropic rocks. (Mogi, 1967 and 2007; Colmenares and Zoback, 2002; Al-Ajmi and Zimmerman, 2005 and 2006; Haimson, 2009; Ma and Haimson, 2012). The findings from these studies show that the effect of ??2 on the strength of isotropic rocks is generally significant, and cannot be ignored.Since the role of ??2 in the failure of isotropic rocks was recognized, other researchers attempted to extend some of these isotropic failure criteria to anisotropic rocks, but with limited success (e.g., Tiwari & Rao, 2007; Singh et al., 1998; Zhang and Zhu, 2007; Pei, 2008). Smith and Cheatham (1980), on the other hand, conducted true-triaxial experiments on the Green River Shale and used a J2-I1 type relationship to predict the strength behavior, but without considering the effect of the three-dimensional orientation of the weak plane on strength. Therefore, unlike for the case of isotropic rocks, additional studies need to be done to determine the role of 2 on the failure of anisotropic rocks."

Jan 1, 2015

By A. Scheltema Beduin, H. Oterdoom, A. Haaks, C. Kempe, A. Liedtke

"For the Kazchrome DC Ferrochrome plant, SMS-group specifically designed an electrode column that would minimize downtime. Special features include the possibility for on load electrode jointing, on load slipping, and back-slipping. Electrode consumption is strongly related to furnace load and impedance, and is very low at high loads, which makes the graphite electrode an alternative to consider even in AC applications. The electrode could also be applied in other DC applications, as for example TiO2 slag, the novel Conroast process, or (twin) DC FeNi furnaces, and could possibly be installed during bigger shut downs. This paper will describe some of the design considerations and operational performance data on this novel column in the FeCr industry that could also be of interest for the FeNi or PGM industry if these would continue to develop towards DC technology.INTRODUCTION Every DC furnace has a cathode and an anode. The anode applied is commonly known as the Concast design, though the original design was by ABB who handed it over to Concast, subsidiary of the SMS-group. Some well-known DC furnaces outside the steel industry where the Concast bottom has been installed are listed in Table 1 as well as a few other key design features.Sidewall cooling and some other data have been left out for reasons of space, and being a lesser relevant design feature to this paper. HES stands for Hollow Electrode System, in which feed is charged through an electrode. Table 1 is built up chronologically according to first start-up date. Below comments are given to briefly explain some of the columns in Table 1: •,“Arm” in the “Cathode” column means that the design of the cathode holding and regulating mechanism consists of an electrode arm as depicted in Figure 1. •,Charging is relevant as some operations started with HES charging, whereas recently most new operations have started without HES but only charging chutes, of which at least some are close to the electrode and possibly others installed closer to the sidewall. •,“Dome” in the “Roof” column means that the design of the roof involves a center piece that can be exchanged if worn. •,The optimal roof design still seems to be a topic of debate, where no design is without problems."

Jan 1, 2017

By The Hollinger Mill Staff

MR. E. L. Longmore (l), in presenting this paper at the Annual General Meeting, suggested that it was neither desirable nor necessary to read any but the briefest portion of the paper. He wished to emphasize that the scale of the milling operations at the Hollinger not only provided several units in continuous operation to ensure that the test work was being carried on under normal conditions, but also provided a large and competent staff to supervise such an extensive experiment and to criticize intelligently the recorded results. He wished to acknowledge the enthusiastic co-operation of the group of assistants whose excellent work was responsible for the confidence with which the conclusions could be presented. The following data were given as a supplement to the published paper. Following conclusion of the test work described in the paper, the size of the balls fed to the mill was changed from 3-in. diameter to 2Yz-in. Within a short time, a gradual rise in the tonnage rate took place and, as no other condition had been changed, the increased tonnage was credited to the changed character of the ball load. At present, the average daily tonnage for the mill with low discharge is 1,507 tons, an increase of 77 tons over the 1,430 tons ground when feeding a 3-in. ball. A similar increase in tonnage should be credited to the high-discharge test, if the 2;Y2-in. ball had been used at that time. This would raise the 1,117 tons ground at that time to a new figure of 1,177 tons. From a study made to determine those factors influencing the rate of pulp flow through the low-discharge mill, it became evident that the ball screen used on the high-discharge mill must have been restricting the flow through the mill, although this was not recognized at the time. Accord-ingly, it was decided to repeat the high-discharge run. Additional length was added to the 21-in. ring grizzly originally used, and the length of the reverse spiral was increased. The end of the grizzly was left open so that, if necessary, excess pulp might flow out. A coarse mesh 3 ft. by 3 ft. flat screen was provided to catch stray balls.

Jan 1, 1937

By A. McLean, X. Men

"With high quality sulphide ores diminishing, nickel production from low-grade oxide ores becomes increasingly important and a major concern is the high sulphur level in the resultant ferro-alloy. So, desulphurization of the ferronickel becomes an important consideration. Red mud is a waste product of alumina industry and causes adverse effects on environment. In the present study, experiments were conducted to determine feasibility of using red mud as a major ingredient of refining fluxes for ferronickel desulphurization. In view of their influence on slag-metal separation, contact angle measurements between ferronickel alloy and molten slag made from red mud were also determined. INTRODUCTIONNickel is a very important metal that finds a variety of applications including consumer products, building, automotive and electronics. There are two kinds of ores containing nickel, namely oxide and sulphide ores. Anthony and Flett [1] reported that about 55% of the world’s nickel production originates from sulphide ores while the balance of new nickel is derived from oxide ores. However, oxide ores count for about 70% of the worldwide reserves. Apparently, oxide ores will play a significant role in the production of nickel in the future. Most of nickel produced using oxide ores is from pyrometallurgical processes namely ferronickel smelting or matte smelting. So, we can draw such a conclusion that using nickel oxide ores to produce ferronickel will become more and more important .Before used to make stainless steel or other materials, the ferronickel need pretreatment to remove impurities especially sulphur and phosphorus that will cause negative influence on the properties of steel product. Red mud is a by-product during alumina production. Usually, about 1 to 1.5 tons of red mud will be produced when we get 1 ton of alumina. About 90 million tones of red mud are produced every year in the world [2].Red mud is deemed as a toxic industrial waste which needs treatment before discharge because of alkalinity and chemical and mineralogical properties. Until now, the main means to deal with red mud are to store it in impoundments or dump it in the sea depending on local environment. Nowadays, how to recycle red mud has become a pretty hot topic around the world. The last 40 years have witnessed a lot of patent applications concerning the ways of red mud utilization. However, a little amount of red mud is consumed until now due to the limitations of recycling red mud."

Jan 1, 2012