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By Rodney A. Cooper

The Mary River iron ore deposits were discovered in the 1960?s. The high quality of the deposits was immediately evident, and drilling programs were successful in defining a resource adequate for supporting a feasibility study. Through a combination of ownership, market and political forces, the deposits were not developed. Having missed the window of opportunity, the project remained dormant for over three decades. Baffinland Iron Mines Corporation, with its 100% ownership of the Mary River deposits, began trading publicly in 2004. Modern exploration and metallurgical test work began in 2004 and has continued in 2005 and 2006, resulting in total expenditures during this period of almost C$60 million. In May 2006, an engineering team lead by Aker Kvaerner E&C, a division of Aker Kvaerner Canada Inc., completed a scoping study based on the first of five identified deposits. Highlights of that study include direct shipping of high quality lump ore to Europe over a 34 year mine life at a production rate of 10 million tonnes per annum, based on an indicated resource of 309 million tonnes at 66.1% Fe and an inferred resource of 28 million tonnes at 65.9% Fe. Capital costs were estimated at C$1.5 billion. Using iron ore prices approximately 36% below that of the 2006 market, a pre-tax internal rate of return of 15% was estimated assuming equity financing. Payback on initial capital was 5.9 years. The pre-tax life-of-mine cash flow was estimated to be C$6.3 billion. As a result of this study, Baffinland initiated a Definitive Feasibility Study, under the management of Aker Kvaerner, which is scheduled for completion in December 2007. This paper touches on the geology of the deposits, the current resource and the outlook for expansion. Project permitting, construction planning, logistical challenges and ocean shipping considerations are discussed, providing a current perspective on the project scale, scope and timeline for commercial production. In addition, an overview of the Company?s bulk sampling program will be provided.

May 1, 2007

By Gordon A. Irons, H. Y. Hwang

An impinging gas jet on a water surface was studied to understand momentum transfer in oxygen steelmaking. Depth, width and oscillation frequency of the jet cavity were measured and the splashes interpreted with Kelvin-Helmholtz instability and combined Rayleigh-Taylor and Kelvin-Helmholtz instability theory. A model to calculate separate gas and liquid velocities, applying the exact momentum boundary condition, is proposed.

Jan 1, 2004

By Thomas Mütze, Frank van der Meer, Gunnar Kretschmar

"The air classification of moist bulk materials is addressed in the literature insufficiently. Therefore a feasibility study investigated the influence of feed moisture on the separation efficiency expecting a reduction with higher moisture content in the feed. Two iron ore samples were characterized in terms of the nature of bonding of moisture as well as its effect on the agglomerate strength. The classification tests themselves were conducted with two dynamic laboratory air classifiers at cut sizes from 45 to 110 µm and a moisture of the feed material between 0 to 10 wt.-%. The quality of classification was assessed by split factor, cut size, separation efficiency, moisture of both products, as well as the yield of the fines as product fraction in a corresponding processing plant. An initial approach was found for a technological implementation: A sufficient drying of a moist feed can be realized simultaneously to the pneumatic transport inside the classifiers. In the present case the drying effect and dispersion intensity was sufficient to provide good separation results up to 3 wt.-% feed moisture. An additional drying step was necessary beyond 3 wt.-%.INTRODUCTIONSpecific information about the influence of moisture in air classifiers is very rare. Mostly only dry products are processed and/or drying prior to classification is recommended. Apart from the specific influence of the feed moisture a lot of literature can be found regarding air classification itself. Altogether air classification can be divided into several sub-process (Leschonski, 1977):- dosage of the feed material,- dispersing the solid feed material into a gas stream,- feeding the dispersed material into the classification zone by pneumatic conveying,- separation of coarse and fine particles, and- separation of fine particles from gas.For wet feed particles also drying plays an important role. Interesting for the present investigation are the humidity of the feed and its influence on the single sub-processes."

Jan 1, 2016

By E. Lee Cameron

The old mine workings on four barite-fluorite veins of the Lake Ainslie district, Nova Scotia, were opened recently and channel sampled. The samples were then forwarded to the Ore Dressing Laboratories, Bureau of Mines, Ottawa, for chemical analyses. Following the analyses, the sample rejects were combined to form bulk samples, representing ore from each vein, for small-scale batch test work. Tonnage samples from the veins, for pilot mill testing, were submitted when requested by the Bureau. The results of the analyses on the channel samples showed that there are two grades of ore in the di strict, represented, for test purposes, by the Trout River vein and Johnson vein ores. Small-scale and pilot-plant test work was undertaken on both these ores. The Trout River ore is amenable co concentration but, because of the high-barite-co-fluorite-co-calcite ratio, a fluorite concentrate was not obtainable from the Johnson vein ore as received. However, Johnson ore blended in a 1 :1 ratio with Trout River ore gave a mill feed that was readily concentrated. The test work on these ores showed that the following concentrates are obtainable : (a) A barite concentrate, with a recovery of 85 to 95 per cent of the available barite, analysing 93 co 96 per cent BaS04, 4 co 6 per cent CaF,, and 0.5 co 2 per cent CaCO3. (b) A fluorite concentrate, with a recovery of 80 co 85 per cent of the available fluorite, analysing 94 to 97 per cent CaF, , 1 to 3 per cent BaSO., and 1 to 3 per cent CaCO3

Jan 1, 1945

By Mike Dry, Mani Manimaran, Mykolas Gladkovas, Mariam Melashvili, Chris Fleming

Development of a new hydrometallurgical flowsheet requires quantitative description and optimisation of all chemical processes in each step in the process, as well as mass and energy balances across a fully integrated circuit incorporating all the unit operations. This paper covers steps leading to early modelling of a gold processing plant based on a bromine/bromide lixiviant. Research conducted at SGS Lakefield in collaboration with Albemarle provided the data that were used in the development of this model. A previous investigation in this series described the leaching behaviour of various mineral types in Stabilized Bromine reagent. This paper is concerned with the gold recovery operations associated with a conceptual bromine leaching process. The recovery of gold by different types of resin and activated carbon was studied, along with various methods of stripping the gold from the extractant. The circuit is closed by electrolysing the spent leach liquor to oxidize some of the contained bromide back to bromine that would allow recycling the regenerated leach solution back to the leach step in the process.

Jan 1, 2015

By K. Huang, L. Pan

"Darcy’s law has been used extensively to model cake filtration. It is a macroscopic model, in which particle size dominates the kinetics and cake moisture. In the present work, a capillary flow model has been developed from first principles to study the effects of microscopic parameters such as capillary radius, boundary layer slip, disjoining pressure, and surface tension that may be important for dewatering ultrafine particles. Laboratory-scale pressure filtration tests show that dewatering kinetics is strongly dependent on capillary radius, which in turn is influenced by surface hydrophobicity, surface charge, and length scales of particles. The model parameters determined from laboratory-scale filtration tests show that use of a hydrophobizing reagent can greatly improve fine particle filtration due to increased slip and decreased capillary pressure. The kinetics of dewatering kaolin clay is fastest at a pH where the layer-structured particles coagulate with a face-to-edge orientation. The capillary flow model developed in the present work may be useful for designing filters and optimizing operating conditions.INTRODUCTIONIn 2014, Canada produced 2.2 million barrels per day of crude oil from the oil sands resources in Alberta, which is expected to grow to 3.2–4.0 million barrels per day by 2030 (CAPP, 2013). One of the challenges to meeting these projections is the tailings management. At present, the industry uses ?77 km2 of the lands to construct tailings ponds. Significant amounts of the tailings disposed to the ponds are silts and clays that are slow to settle and consolidate, creating problems for reclamation. It is, therefore, the objective of the present work to develop a filtration model that may be useful for developing advanced dewatering technologies that can address the tailings management issues.Darcy’s equation is the governing law for filtration. It was derived as an empirical model by H. Darcy in 1856 while he was studying fluid flow in sand filters (Hall, 1956). Since then, many investigators derived it from first principles. It relates superficial fluid velocity to the pressure drop across a porous medium of known permeability. In filtration, the latter can be related empirically to particle size using the Kozney equation (Yoon, 2005)."

Jan 1, 2016

By D Maschemeyer

The cobalt production of the Chemical Metallurgical Division of Sherritt Gordon Mines Limited was supplemented, from May, 1962, to June, 1964, by processing 3,500 tons of a calcined cobalt-nickel concentrate. This paper describes the reduction roast -aqueous sulphuric acid leach process and the operating practices employed in recovering cobalt, nickel and copper from this oxidized by-product of a lead-copper operation. Incorporated in this paper is a description of how this process was integrated with the existing ammonia leach -hydrogen reduction process for the recovery of nickel and with the soluble cobaltic ammine process for the recovery of cobalt. Introduction T HE Chemical Metallurgical Division of Sherritt Gordon Mines Limited at Fort Saskatchewan, Alberta, has recently completed ten years of operation utilizing the ammonia leach process for the recovery of nickel, cobalt and copper from sulphide concentrates, matte and other materials amenable to this process. A continuing development program, along with the expansion of certain facilities, has enabled the refinery to attain a production rate of 31, 000,-000 pounds of nickel per year by hydrogen reduction. This represents an increase of approximately 85 per cent over the original design capacity of the plant.

Jan 1, 1965

By A. E. Flynn

Introduction The purpose of this paper is to record the results of researches on anhydrite that have been undertaken at the Nova Scotia Technical College during the last two or three years. The work was carried out in the Mining Department of the College, and several investigators were engaged on the work from time to time. Messrs. R. E. Hanson and K. M. Ed devoted their mining theses to the investigation, and Messrs. J. C. Hall and G. W. Baker were engaged on the work during the summers of 1928 and 1929, respectively. The writer was responsible for the initiation and direction of these researches, which were made possible by the financial assistance of the Department of Mines of the Province of Nova Scotia. Nova Scotia is the leading Province in Canada for gypsum production, its output being about a million tons per year, or over 80 percent of the production for the whole of Canada. The commanding position of Nova Scotia in this regard is due to the abundant supply of gypsum and its proximity to areas of intense utilization in the eastern United States. The resources of gypsum are contained in a total area of 627 square miles of gypsiferous rocks, so that an enormous reserve of this mineral is available. Present operations are confined to the highest grade and most easily worked deposits close to water transportation.

Jan 1, 1930

By M. Saghaeian, R. Hejazi, Sheikhzadeh. V., A. Abazarpoor

"The morphology of iron ore concentrate particles is known to influence several properties of iron ore pellets, including drop number, compression strength, and reduction properties. This study aimed to determine why grinding methods and particle morphology influence green pellet properties. Particle morphology can be directly related to the grinding method: ball mill (BM) or high pressure grinding roll (HPGR). In this research, two iron ore concentrates, namely Gol-e-gohar and Ardakan, were ground by a BM and HPGR to produce a suitable Blaine value for pelletizing (nearly 1,900 cm2/g). Morphological features of particles were subsequently investigated by sieving and cyclone sizing for particle size distribution, the Brunauer-Emmett-Teller (BET) and Blaine methods for specific surface area (SSA) and scanning electron microscopy (SEM) for the evaluation of shape descriptors. It was concluded that HPGR resulted in generating a greater proportion of fine particles compared to ball milling. The BET results showed that the SSA of HPGR pellet feed was higher than that from the BM due to a greater proportion of cracks, indents and bulges and the creation of new fracture surfaces. SEM images processed by ImageJ software indicated that particles ground by the BM tended to be more rounded and less elongated than HPGR particles because of the breakage model of each method. To show the effect of morphology on the green pellet quality, green pelletizing was performed in a pilot pelletizing disc. The HPGR pellet feed had a higher drop number in comparison to the BM pellet feed, due to the angular and rough shape of the particles.INTRODUCTIONThe demand for fine or ultra-fine particles as feed material for ceramics, composite materials, and mineral processing is on the rise (Kotake, Kuboki, Kiya, & Kanda, 2011). The final concentrate of iron ore processing plants (filter cake) has a size range roughly between 0 and 300 µm, depending on the liberation of Fe containing particles (Meyer, 1980). In general, to produce a suitable pellet feed of appropriate quality, Blaine numbers of 1800–2000 cm2/g should be achieved (Van der meer, 2015)."

Jan 1, 2016

By Davis. B. R.

There are a number of hydrogen storage systems proposed for automotive applications. However, all of the current suggestions fall short of the DOE targets for storage density and volumetric density. The DOE has suggested that only through radically new ideas will a practical solution to hydrogen storage for automotive applications be realized. This paper provides some background in chemical hydrogen storage and discusses roadblocks to the implementation of some proposed technologies.

Jan 1, 2005

By S. A. G. Poppen, J. Konopka

"Petro-Canada's project investigative and experimental underground work in the Galt coal seam at Kipp, Alberta included the construction of an emergency egress and ventilation raise from surface.The Galt coal seam is 206 m below the surface and covered by 8 m of shale, sandstone and clay beds of the Oldman Formation, 128 m of shale, bentonite beds and sandstone of the Bearpaw Formation, and 70 m of glacial till and basal gravel. Free water was encountered from sand lenses in the glacial till and the basal gravel.Excavation for the ventilation raise was a nominal 2 m diameter and it was lined with a welded, continuous, steel liner of 1.2 m 1.D.Excavation in the glacial till was carried out with a surface bucket drill rig combined with hand mining. Ground support in the glacial till consisted of liner plates and backfill grout. The Bearpaw and Oldman formations were excavated by employing standard raise boring techniques. The individual liner sections of the continuous steel liner were assembled on surface and the entire liner column was lowered into the raise by employing hydraulic lift climbers. The liner column was back-filled with concrete.IntroductionThe Lethbridge Coal Field comprises an area of approximately 1 165 square km in the vicinity of Lethbridge in southern Alberta. Pacific Petroleums Ltd. proposed in a Preliminary Disclosure in January 1979 ""to open an experimental mine prior to development of a producing underground mine with an initial rate of 540,000 tonnes per year"". A 2 900-hectare group of leases in the Kipp-Coalhurst area was selected as the primary target for the Kipp experimental mine. The scope of work for the experimental mine included the construction of an access shaft with production capabilities, the construction of a ventilation and emergency exit raise, drive-ages in coal to yield a bulk sample, gathering of geotechnical information and the testing of experimental mining methods."

Jan 1, 1983

By Michael Samis

The real option valuation method is often presented as an alternative to the conventional discounted cash flow (DCF) approach because it is able to recognize additional project value due to the presence of management flexibility. However, these two valuation methods can be separated on a more fundamental level by their differences in risk discounting. Real option valuation applies the risk-adjustment to the source of uncertainty in the cash flow while the DCF method adjusts for risk at the aggregate level of net cash flow. This seemingly small difference is the reason why the real option method is able to differentiate between projects according to each project?s unique risk characteristics while the conventional DCF approach cannot. This paper provides an overview of the real options and DCF valuation frameworks and discusses the differences in risk discounting that exist between the two methods. Using grade-school mathematics, this paper clearly demonstrates how, with real options, a unique project risk discount can be calculated which is directly linked to the project?s unique risk profile. It also highlights why the DCF method fails in this regard and shows why a call to ?increase the Risk-Adjusted Discount Rate? is an incomplete solution at best. Finally, a heap-leach project and satellite reserve development project are valued with both techniques and the difference in investment conclusions is explained in terms of the risk-discounting concepts discussed here.

May 1, 2003

By R. M. de Carvalho

Empirical and phenomenological models of SAG and AG mills have dominated the scene for the last few decades, being used nearly every time a mill is scale-up or an attempt is made to improve its performance. However, in spite of their success, these models fail to describe the details of both the grinding environment and the breakage mechanisms. The result is a generally good track record, with a few ? but very important ? instances when these models fall short in their predictions, which can be very costly. Further, nearly all of these models require detailed calibration to industrial and pilot plant data, being costly in their application, although still with limited predictive capabilities. Recognizing these challenges, a new model framework has been proposed which overcomes the limitations of currently used models by coupling the microscale population balance model, the discrete element model and functions that describe internal classification at the mill grate. This model framework has already been applied to describe breakage in ball mills, with showing very promising results. However, in the case of SAG and AG mills a much greater level of complexity is involved in the model, given the interaction between the mechanical environment that is formed totally or, in part, by the ore itself, and the breakage response in the mill. The paper describes the model framework, demonstrating how it can address appropriately effects such as mill speed, mill filling and size distribution, on the response from the mill. Some of the advantages of the model are related to its intrinsic multi-component and dynamic nature, which allows simulating appropriately mixtures of ores of different competences in a transient manner.

Jan 1, 2012

By Thomas J. Ferree

"IntroductionThe objective of this chapter is to provide important criteria relating to the operation and maintenance of spiral concentrators and Reichert cones, so that optimum performance can be achieved and maintained in a mill circuit. Little space will be devoted to the theory of the forces of separation or the technical details of the devices, as that in(ormation is available from other sources. The discussions in this chapter will be presented in a manner that should be interesting and useful to anyone who may be exposed to these methods of gravity separation. It is most important for the reader to understand that each application of the equipment should be considered on its own merit, as no two suites of minerals respond exactly alike in mineral processing circuits. The information herein should be useful as a guide toward increasing the metallurgical efficiency of the processes, irrespective of the type of minerals being concentrated.It is highly recommended that certain tools be made available to the personnel responsible for metallurgical performance of the spirals and cones; these include a set of standard type laboratory screens, a medium quality petrographic microscope, and a vanning plaque or gold pan. These devices, when used in conjunction with a basic understanding of the mill circuit, provide immediate knowledge on the characteristics of any sample in the mill circuit. The information gained by visual evaluation of sized fractions of products can solve many of the basic problems of mineral beneficiation, and at the same time increase the interest of the non-technical personnel associated with the plant operation. There are many problems associated with obtaining satisfactory performance from a mill operation. Many of the problems have to be solved using a technical approach. Some of the problems, however, can be solved with practical knowledge and common sense. It is hoped this chapter can provide a guide to both techniques."

Jan 1, 1989

By L. A. Kaufman

"AbstractThe Bureau of Mines, Ottawa, in conjunction with Eldorado Mining and Refining (1944), Limited, has developed an automatic concentrator for uranium ores.A brief review of radioactivity is given which shows the present limitations for obtaining •good results by electronic concentration. The ores after being mined, or at any other stage of a concentrating process, can be analysed for the uranium dissemination and from these results the possibility of an economical concentration can be estimated.The electronic and mechanical parts of the equipment are described. Difficulties in the operation's, and means by which they have •been and are •being overcome, are discussed. The capacity of the units, the results obtained, the man-power required, and other pertinent data are presented. Simple control analysis on the percent U3O8 has been developed.Research is continuing along lines to develop a more completely automatic unit using a scintillation detector, and applying a few new ideas and some devices that have been developed in colour sorting.IntroductionIt is well known how important the Geiger counter is in the field of exploration for uranium. It was in late 1946 that Dr. C. Lapointe, of the Bureau of Mines, Ottawa, first started work on an automatic picking unit to concentrate uranium ores. He used the principle that if a number of pieces of unknown ore, some of which contained uranium, were individually brought up close to a Geiger-Müller tube, then the pieces containing uranium could •be easily picked out due to their activity. Since then a good deal of work has been done by the Bureau of Mines in conjunction with Eldorado Mining and • Refining (1944), Limited, to develop an electronic concentrator."

Jan 1, 1950

By J. Kalynchuk, R. Poulin, R. Pakalnis, D. Peterson

"A market survey of all underground mines in Canada (Fig. 1) was completed to determine the ex1ent to which cable bolts are being used as ground support. One hundred and nine questionnaires were distributed with 65% of them being returned completed. As a result, the quantity, types and locations of cable bolt installations have been determined. In addition, the grout pumps used, time required and cost of installations have been addressed. The popularity of other support types used in underground mines is summarized.IntroductionRock reinforcement consists of installing support within the rock mass (internal) or within the excavation (external). Within the past twenty-five years several new types of ground support have been introduced within the mining industry. Cable bolts were first introduced to the Canadian mining industry in the 19605 (Gramoli, 1975).Cable bolts have gained increased acceptance as a method of ground support during the past decade. The cable bolt is comprised of one central wire surrounded by six slightly smaller outside wires and typically having an over-all diameter of 15.2 nun to 15.9 nun (0.6 in. to 0.625 in.). The cable wires are made from high-strength steel with an ultimate strength of 26 tonnes (58 000 Ibs) (Goris, 1990). Current cable bolting practice consists of cement grouting one or more steel cables within a drill hole in order to reinforce the surrounding rock mass. Cable bolts are a form of internal support and are generally used in open stoping and cut-and-fill mining methods (Fig. 2).They can be either tensioned (active support) or untensioned (passive support) and may be of any length, though for underground mining applications, they are generally less than 20 m (65.5 ft) and are untensioned. A major advantage compared to previously used ground support is that the cables are flexible and can be coiled to approximately 1 m (3.3 ft) diameter coils allowing relatively long support to be installed in areas with limited head room (1.8 m).With the advent of bulk mining methods which require comparatively large stopes, it has necessitated the use of relatively long support tendons to stabilize the large spans being created. As a result, the use of cable bolts for ground support in mines has increased significantly within the past decade."

Jan 1, 1995

By John Ramsay, Frederick A. Edwards, J. Eric Belford, Clarence V. Amyotte

"DEVELOPMENT of the underground mine is carried out from a main ramp driven from surface to the 800 and 1200 levels and from a shaft sunk to 3050 ft. Sub-levels at approximately 100-ft vertical intervals, starting 300 ft above the 800 level, were collared from the ramp as it advanced and a considerable amount of development work was pursued. The lateral and ramp work provided an opportunity to train drift crews, test equipment and methods, assess ground conditions and advance the initial stope development. The development of an exhaust ventilation raise system closely followed the extension of the level development to the south end of the orebody (Fig. 26) . On completion of the shaft, temporary lip pockets were installed at the shaft stations on the 800, 12.00, 1600, 2000 and 2400 levels. The permanent loading pocket had been installed on the 2860 level as part of the shaft-sinking contract and the level had been connected to the 2800-level station by a small muck transfer raise. Development of the 2800-level area was given top priority at this stage. This included a 1000-ft cross-cut to the crusher location, the crusher excavation, storage bins, conveyor ways and a ramp to connect the 2800 level to the loading pocket, the rope changing station and the shaft sump. The development and installation of the rock handling system between the 2400 level and the shaft bottom was contracted to Macisaac Mining and Tunnelling Company, of Sudbury, Ontario."

Jan 1, 1974

By B. C. Tripathy, S. K. Padhy, R. K. Paramguru, Pinak Patnaik

"Cobalt is regarded as technology enabling metal among the less abundant metals found in the earth crust. Electrodeposition of the metal from aqueous sulphate solutions is the preferred technology for its commercial production. Certain additives are added in the electrolytic bath during electrodeposition for the production of uniform, smooth and bright deposits of cobalt metal with higher current efficiencies and lower energy consumptions. A detailed investigation has been carried out to determine the effect of additive Sodium octyl sulphate (SOS) on the electrodeposition of cobalt. The results indicated that the additive apart from increasing the current efficiency enhanced the quality of cobalt electrodeposits. A maximum current efficiency of 98.4% was achieved at 10 mg/L SOS. The presence of SOS in the electrolytic bath did not result in any noticeable change in the crystal orientation of the cobalt electrodeposits except at higher concentrations (50 g/L) of the additive. Scanning electron micrographs indicate that smooth, compact and uniform deposits of cobalt were obtained at 10 mg/L of the additive beyond which there is detoriation in the quality of deposits. Polarisation of the cathode was seen in the presence of SOS. The effect of SOS on the electro kinetic parameters: Tafel slope (b), transfer coefficient (a) and exchange current density (I0) has also been investigated.INTRODUCTIONCobalt is recognised to be a strategic metal among the various base metals, owing to its numerous applications in various industries. Cobalt as a metal finds its use in the preparation of rechargeable batteries, as an alloying component in the manufacture of space craft and jet engines, as a drying agent for paint varnishes, ground coats for porcelain enamels and in the manufacture of the catalysts for the petroleum and chemical industries (Davis, 2000). Although there is a huge demand of the metal the supply of the metal is limited due to its scarcity in the earth crust. The metal comprises of only 0.02% of the earth’s crust for which the metal is recovered from various ores of nickel and copper as a byproduct (Hawkins, 1998). The present day techniques for cobalt production includes mostly hydrometallurgical methods which are preferred over the earlier used pryometallurgcial methods carried out for the metal production. Almost 70% of the world’s cobalt produced currently is from hydrometallurgical techniques in which electrodeposition of the metal from aqueous solutions is the final step for metal production (Verney, 1988). The most preferred route for electrodeposition of cobalt metal involves a sulphate medium (Das and Subbaiah, 1984; Lenthall and Bryson, 1997; Sharma et al., 2005)."

Jan 1, 2016

By H. C. M. Gordon

IN every co al mine the hazard of fire in the underground workings is always present. This is a natural condition, because the mineral wrought and the supports normally used are themselves combustible and because underground fires may originate in many ways. The main causes of such fires are: (1) Open lights igniting gas, wood, or the coal itself. (2) Explosives igniting gas or coal. (3) Spontaneous combustion. (4) Electric sparks or short circuits igniting gas, wood, or coal. (5) Frictional heat igniting wood, coal, or oily material. (6) Explosions of gas or coal dust igniting wood or coal. The relative importance of these causes, and others less frequently met with, varies from district to district and depends to a great extent on the equipment and supplies used, on the training and discipline maintained among the workmen, and on the liability of the coal to heat spontaneously. It follows, therefore, that open lights should not be used and that, wherever there is the least possibility of methane being encountered, only locked safety lamps of an approved type should be allowed underground. It follows, also, that only 'permitted' explosives should be used and that the it use should be confined to conditions under which they are ?permitted?. Similarly, electric equipment should be of an approved type and all such stationary equipment should be housed in well-ventilated fireproof rooms. Such rooms, as well as entries to stables and oil houses, should be provided with doors which can be easily and tightly closed and which should be constructed of fireproof material.

Jan 1, 1937

By Gordon E. Forward

Sustainable development in the mining and metallurgical industrial sectors can be implemented by two distinct groups: 1. The managers and operators of existing facilities ? working with industrial processes and systems that have been operating for many years, in some cases several decades 2. Those who are engineering and building new facilities or upgrades of existing plants. The focus of this paper is to look at approaches and tools available to bring sustainable development thinking and action to these two distinct areas. The paper will discuss: 1. The concept of working with existing facilities and processes to minimize waste and maximize product from a regional raw material/resource consumption perspective using By-product Synergy. The BPS concept has been carried out in Texas (TXI-Chaparral?s STAR Example), in Calgary, Alberta with Suncor assistance, in Montreal, Quebec with Noranda as project leader and currently in Ontario with Dofasco as industry champion. Successful outcomes are presented. 2. A new approach to designing greenfield facilities or upgrades to existing plants using a sustainable development review process. Undertaking the review early in every project can institutionalize the concept of discussing and implementing waste elimination, design for environment, eco-efficiency, risk minimization and community awareness and involvement in the future capital stock. The review process has been developed as an excellent tool for measuring and reporting on a corporation?s sustainable development progress. Examples of how this approach has been adopted in the mining and metallurgical industry, and the resultant outcomes are given.

May 1, 2002