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By Todd Kennedy, Kurt Tischler

"This paper presents the field experience of Siemens with regard to Gearless Drives in the mining industry … We consider the elapsed production hours of Siemens Gearless Drives to be extensive enough to provide a reliable estimation of availability calculated on the basis of data representing repair time, downtime, and operating time. The users of Siemens Gearless Drives confirm an availability of more than 98% after exclusion of some major issues. The paper addresses all of these major issues of Siemens Gearless Drives, describes successfully applied corrective measures, explains the influences on the design of the drives, and lists the downtimes associated with each of these issues. INTRODUCTIONThe first Siemens Gearless Drive in mining began operation in 1980 and was installed on an iron ore grinding mill. As of September 2011, 40 Siemens Gearless Drives will have accumulated 136,000 machine days of operating time. According to pure rules of statistical analysis, this accumulated operating time may be too low to be used for a sample of field data to provide precise statistical values of availability, however we consider it large enough to provide indicative values of worthwhile consideration. This paper deduces the availability statistics of Siemens Gearless Drives in the mining industry from information received from end users and factors in all major issues that Siemens has experienced with our Gearless Mill drives. This paper does not consider Gearless Drives in other industries such as the cement industry. The cement plants, especially in the northern hemisphere, interrupt production for several months in the winter. This lengthy downtime is used to maintain the complete plant including the mills, their motors and their drives. The mining industry has more stringent and higher production requirements. In the mining industry, production must run 24 hours per day, 365 days a year."

Jan 1, 2012

?Bank Lending to the Mining Sector affected by 2 broad factors since Mid 2008: 1. Economic Downturn (?the Real Economy?) for mining companies directly reflected in lower metal prices (current and forecasted). In the case of gold companies, minimal effect. 2. Financial Crisis (?the Credit Crunch?) as it has affected the lending activities of those banks active in the mining sector.

May 1, 2009

By Y. Masuda, Y. Nakamura, K. Haraguchi, A. Taniguchi, K. Arita

Availability of the elements originating from steelmaking slag for various marine phytoplankters including purely cultured strains and natural populations was tested in land and shipboard laboratories. The results obtained demonstrate that, although their solubility was not uniform among different slags, dissolved iron, silicate and phosphate from the slag stimulate phytoplankton growth. No negative effect of the slag addition on their growth was observed at lower dose, e.g. less than 200 mg/l. Since most coastal waters are the hypernitrogenous environment due to anthropogenic load of ammonium, the slag can be an ideal source of silicate and phosphate, which restore the hyper-nitrogenous condition into truly eutrophic condition with balanced Redfield ratio. In so-called HNLC (high-nitrate low-chlorophyll) oceanic areas such as the Southern Ocean and the subarctic North Pacific, where iron is almost only limiting factor of the phytoplankton growth, the slag could also be a source of iron and enhance CO2 fixation by phytoplankton photosynthesis. (Key words: carbon deposition in deep-sea, marine environmental remediation, phytoplankton, slag as a source of iron, phosphate and silicate.)

Jan 1, 2004

"The measurement of error in assays collected as part of a mineral exploration program or mining operation historically has been undertaken in a variety of ways. Different parameters have been used to describe the magnitude of relative error, and each of these parameters is related to the standard measure of relative error, the coefficient of variation. Calculation of the coefficient of variation can be undertaken in a variety of ways; however, only one produces unbiased estimates of measurement error: the root mean square coefficient of variation calculated from the individual coefficients of variation.Thompson and Howarth’s error analysis approach has also been used to describe measurement error. However, because this approach utilizes a regression line to describe error, it provides a substantially different measure of error than the root mean square coefficient of variation. Furthermore, because regression is used, Thompson and Howarth’s results should only be used for estimating error in individual samples, and not for describing the average error in a data set. As a result, Thompson and Howarth’s results should not be used to determine the magnitudes of component errors introduced during geochemical sampling, preparation, and analysis.Finally, the standard error on the coefficient of variation is derived, and it is shown that very poor estimates of relative error are obtained from duplicate data. As a result, geoscientists seeking to determine the average relative error in a data set should use a very large number of duplicate samples to make this estimate, particularly if the average relative error is large."

Jan 1, 2007

By S. Beskri, N. Geoffroy

When metallurgical plants are designed, a lot of consideration is almost always given to material selection and possible damage mechanisms. Materials engineers are generally involved and will strive to select optimal alloys or refractories for the design conditions. Once in operation however, things often change. Equipment is overstressed to increase production, corrosive leaks are not always promptly addressed, proper welding procedures are disregarded, etc. This paper describes various cases from the mining and metals industry where serious mishaps, associated consequences, and costs could have been avoided if simple and proactive steps had been taken. It also presents a strategy, based on safety reviews, to help predict and prevent damage to critical pieces of equipment.

Jan 1, 2015

By K. Tada

A surface of a magnesium slab (hereinafter, referred to as Mg slab) or magnesium alloy slab (hereinafter, referred to as Mg alloy slab) becomes black with the use of conventional manufacturing methods. The methods also involve a danger that an explosion of the molten magnesium occurs through contact with cooling water. Thus, it has been impossible to produce a slab having a smooth surface. In the present method, a molten magnesium alloy is supplied into a mold, which is sheltered from the surroundings, from a pool in tundish, which is shielded from the surroundings. AZ31B slab formed in the mold is drawn horizontally out to the outside from an outlet of the mold by revolving of the pinch rolls. A cooling water is jetted at an acute angle onto the slab on the outer side of the mold outlet. Further, an inert gas is sprayed at the inner side of the mold outlet in order to seal in a gap between a free surface of the molten magnesium alloy and a sheltered board surrounding an upper surface of the mold. As a result, oxygen and droplets of the cooling water are prevented from running back to the mold side through the gap. AZ31B slab has a smooth surface condition and high quality without a gas porosity, a segregation or the like. Further, the slab has a size of a width of 100 to 120 mm, a thickness of 5 to 20 mm and a span of 10 m or more.

Jan 1, 2006

By W. J. Witte, M. K. Witte

"INTRODUCTION A pilot heap leach project was conducted on the B. P. Selco Hope Brook ore located in southwestern Newfoundland. The pilot heap leach was undertaken to assess and confirm the following:1) To confirm that heap leach technology now currently used in the United States could be applicable to Canadian sites.2) To confirm gold recovery by the Heap Leach technology3) To assess and study potential heap leach environmental concerns in cold and wet climate.PROJECT LOCATIONThe Hope Brook Project is located near the southern coast of the island of Newfoundland, approximately 100 kilometers southeast from Stephenville, approximately 25 kilometers west of Burgeo, Newfoundland. Access to the property is limited to helicopter, float plane or by sea through Couteau Bay. The property was covered with muskeg and very few trees. Surface relief was moderate with many poorly drained boggy sections confined to small areas. Many small unnamed lakes were located on the property which provided water to small brooks and streams flowing to Couteau Bay.Several locations were reviewed for the installation of the pilot heap leach facility. Important parameters used to select a suitable location included sufficient land area, close proximity to but not on the ore body, reasonable protection from wind, sufficient supply of fresh water, close to a suitable location for a portable crushing plant, reasonable access to aggregate material for preparation of the test pad and ponds and suitable contours to minimize site preparation."

Jan 1, 1987

By Philippe Caudal

This paper presents a case study in which a slope failure was observed on the west wall of the LAB Chrysotile mine in the vicinity of the main ramp in January 2010. Airborne LiDAR results were used to evaluate slope geometry before and after the failure occurred. Back analysis of the failure was performed to assess rock mass properties at the slope scale using various numerical techniques: limit equilibrium, finite elements and fracture networks. The various approaches correlated well with field observations.

Aug 1, 2013

By Jerry Ran

Stope back caving in a mine using open stope mining interrupts the planned mining sequence and causes oversize muck and dilution with adverse economic implications. As many underground mines are mining or have plans to mine at depth, stope back caving could have significant impact on the mine sustainability. Therefore understanding the caving mechanism is important in preventing future occurrence and improving the mine economic performance. Falconbridge Ltd Kidd Mine is an underground mine using open stope method producing more than 2 million tonnes per year. Its current lowest mining level is approximately 2070m below surface and the lower portion of the orebody is being developed. In years of 2002 and 2003 the back of three stopes of approximately 1900m and 2040m below surface caved and caused substantial mining difficulties at the mine. As these caved stopes are subject to stress concentration, seismic activities and near major geological structures, the caving mechanism is much more complex than that of stopes near surface. In addition to the recorded caving history by site observations, the mine carried out cavity surveys and test drilling to define the caved profiles. The mine seismic monitoring database provided information at these stopes. Stress modeling work conducted by an external consultant provided the stress patterns at the caved stopes. Based on the field collected data and stress and seismic information, this paper presents details of the stope caving process and cavity surveyed profiles, reviews the predictability of the stability graph approach for these cases, discusses the caving mechanism and proposes preventive measures to stabilize the stope back.

May 1, 2004

By R. J. Grant, D. W. Scott, M. A. Cooper, E. R. Shannon

"The Brunswick No.12 site concentrator processes 10,500 tonnes per day of a complex lead, zinc, copper and silver ore producing four concentrates by differential flotation. Research over the years has led to increased understanding about the mineralogical and flotation characteristics of the ore. A pilot plant study conducted at Lakefield Research, primarily designed to investigate the technical feasibility of semi-autogenous grinding, provided an opportunity to test recent theories about grind size,, reagent schedule and flowsheet configuration.A fundamental approach was taken with each of the unit processes involved in grinding and flotation. Basic operating and metallurgical criteria were established in each area and a step-by-step implementation program initiated in the plant.Significant improvements in metallurgical and financial performance have been realized through the application of basic grinding and flotation principles using conventional milling equipment and without major capital expenditure. Current operating practice is detailed and metallurgical advancements cited.INTRODUCTIONBrunswick Mining and Smelting operates one of the world's largest lead and zinc mines near Bathurst, New Brunswick. Brunswick is a publicly traded company with majority ownership held by Noranda Minerals (63%). In addition to the No. 12 underground mine, the company operates a lead smelter and phosphate fertilizer plant in Belledune, N.B. and maintains a 25% interest in a joint venture with Noranda Minerals' Heath Steele mine near Newcastle, N.B. Through the four operations, a total of 2000 people are directly employed with fixed assets totalling Cdn$187 million (BM&S, 1991)."

Jan 1, 1993

By C. R. Lattanzi, M. J. Davie, L. D. Smith, J. R. Goode

"The feasibility study determines the technical and economic viability of a project at the level of accuracy required by the owner to make an investment decision. The economic analysis usually includes calculation of the discounted cash flow. Key components of the cash flow include the project capital cost, the operating costs, construction schedule, production rate, product prices and financial details.This paper examines the feasibility study for a typical mining project. The paper discusses acceptable methods for the determination of the geological and mineable are reserves. The prerequisites for the development of suitable designs, and the calculation of capital and operating costs for the mine, processing plant, and infrastructure are reviewed. Important features of the cash flow analysis and its interpretation are presented.IntroductionThe commercial development of a mineral deposit invariably requires the investment of capital in preparing the deposit for production. It has always been necessary, therefore, to demonstrate to the investor of capital that production of a saleable commodity is both technically feasible and economically viable. The instrument typically used to make these determinations is the feasibility study .Traditionally, mining companies have been rather conservative in their approach to financial management. Until about thirty years ago, the capital required to develop a new mine was most frequently provided by internal corporate cash now and virtually all of the necessary technical and economic studies were performed 'in house'. Under these circumstances, there was little standardization as to what constituted an acceptable feasibility study, and even less as to what constituted an acceptable measure of economic viability."

Jan 1, 1991

By A. Vien, B. C. Flintoff, R. P. Edwards

"A major thrust in industrial automation over the past few years has been the exploration and incorporation of higher level (supervisory and optimizing) control strategies aimed at more fully exploiting the existing control system infrastructure. There is little doubt that this is worthwhile pursuit, however, one must not lose sight of the maintenance and development requirements of the basic control system. Statistics on loop performance indicate that this too is an area of considerable potential. In this paper the authors look at some of the less elegant, but arguably equally important issues in mineral processing plant control. The focus is on controller structures, loop tuning, sampling, filtering, and some of the other principal components of the regulatory control foundations. In many cases significant improvements in control performance can be achieved by getting back to the basics.IntroductionEvery modem concentrator has a state-of-the-art computer control system in place. Several of these are contemplating installing, or have installed, advanced control systems, such as expert systems, but the foundation of their control system remains the PI controller. Although the authors are not aware of any survey or published information for the mineral processing industry, it is reasonable to assume that it would be similar to the pulp and paper industry. ""The average [pulp] mill has some 2000 control loops of which 97% use PI control. This means that PID, dead time compensation, adaptive controllers, specialized algorithms, ... , constitute only 3% of the installed base."" (Bialkowski [1992]).Bialkowski [1992] also stated that ""If you have been keeping score, only 20% of the loops surveyed actually decrease variability in automatic over manual mode of operation, in the short term."" He attributed the problems to: poor tuning (30% of all loops), control valve problems (30% of all loops) and process problems (20% of all loops). There appear to be a serious need to go back to basics and review the existing control system."

Jan 1, 1998

By J. Arzúa, L. R. Alejano, P. Rodríguez

"After six years of successful excavation with a support pattern, some roof failures have happened in a room-and-pillar underground mine excavated in a stratified rock mass. In this paper the authors give explanation to the failure mechanisms and redesign the support in order to solve the recent problems of roof collapse.The mine magnesite bed is approximately 14 m thick, and runs in the E-W direction with an average dip of 18º. Due to mine constraints and the inexpensiveness of the ore, the excavation is performed in the mineral itself, avoiding the need for mining dumps and waste excavation. Rooms and drifts are oriented in the mineral bed to form an angle of 19º to the direction of the strata, reducing the roadway gradients to less than 10%. Rooms are 11 m wide and pillars are 7 m wide. They are designed to leave a 1 meter thick magnesite bed in the roof of the rooms to act as a supporting beam. More often than not there is a layer of very poor quality marly shale up to 1 m thick resting on this magnesite beam. Overlying the poor quality marly shale there is a better quality, self-supporting sandy shale.First, a characterization of the involved rock masses (magnesite, marly shale and sandy shale) is performed, collecting the geotechnical data from previous studies and performing new in situ geotechnical surveys to check the soundness of the previous data. A distinction is made between different quality zones inside the mine by means of the RMR classification, associated to the occurrence of faults.Back-analysis of some of the failed rooms and data from exploration boreholes were used to determine the failure mechanisms (associated to Voussoir type failure or occurrence of sub-horizontal joints crossing the magnesite beam). Based on an understanding of these mechanisms, the method used to design the support and to analyse the stability of the different possible thicknesses of the magnesite beams of the rooms and drifts of the mine were improved.Numerical models of the most relevant conditions found in the mine were performed in order to better understand the behaviour of the roofs and to simulate the supporting solution obtained in the previous analyses.The most relevant conclusions with regard to the main elements causing the instability of the roof include the appearance of sub-horizontal discontinuities crossing the magnesite bed, which causes total instability of the roof and usually leads to collapse; excessive thinness of the magnesite beam in the problematic rooms; and the relaxation of the deepest (around 270 m in depth) zone of the mine because of existing faults."

Jan 1, 2015

By G. Russo

During the construction of the S.Martin La Porte access adit, 9m in diameter, extreme squeezing conditions occurred while crossing Carboniferous schistose formations, with diametric convergence up to 2m. Back-analysis of monitoring results shows that an acceptable simulation of the observed excavation behaviour can be derived using a combined analytical and numerical approach, with simple elasto-plastic laws but a progressive reduction of residual parameters and a relevant increase of dilatancy.

May 1, 2009

By L. H. Van Loon, E. G. Crayston, A. F. Heather

"IntroductionTHE PROPERTY of Madsen Red Lake Gold Mines. Limited. comprises fifty-eight claims located between Russet and Faulkenham lakes in Baird and Heyson townships of the Patricia portion of the District of Kenora. The mine is seven miles •by road from the hamlet of Red Lake. This latter centre is .connected to the Trans-Canada Highway at Vermillion Bay, Ontario, by provincial highway No. 105 which has a total length of one hundred and thirteen miles.The Madsen property went into production on August 11th, 1938, with a milling plant rated at 300 tons per twenty-four hours. Changes made in the crushing plant during October, 1938, allowed a •finer feed to be delivered to the mill and increased the milling •capacity to 400 tons per twenty-four hours. During 1948, steps were taken to add an additional 400-ton capacity unit to the milling plant. This latter work was completed by May, 1949, giving the present total milling .capacity of 800 tons per twenty-four hours.It was realized in the early 1940's that, eventually, considerable material suitable for backfilling in the mine would be required. As there were no large deposits of gravel or sand anywhere near the mine, officials of the Company considered it advisable to investigate the feasibility of using mill tailings for mine •backfill. Since Madsen mill tailings contain approximately 80 per •cent of minus-200-mesh material, it was not considered practical to use them for underground backfill without first removing the finer material. as the slimes present in a product of this fineness would without doubt preclude any drainage whatsoever. Table I gives a sizing analysis of Madsen mill tailings and Table II a quantitative chemical analysis of the ore."

Jan 1, 1954

By Jamel Sgaoula

During 1998 COGEMA Resources Inc. announced that operations at the Cluff Lake mine will be suspended indefinitely by the end of year 2000 as a result of deteriorating uranium market conditions. To remain profitable, Cluff Lake mine started an aggressive cost reduction program. A cost reduction team was formed to look for ways to lower production costs. The focus of this team was to screen all areas and to come up with ideas to improve the economy of the mine and ensure the viability of the operations over the remaining three years of operations. Among the areas where significant savings were achieved are: (1) replacement of portion of cement with fly ash in backfill; (2) backfill optimization; and (3) new shotcrete plant installation on site for bulk in-house shotcrete production. This paper details the research work carried out by Cluff Lake mine on these projects and results of its application.

May 1, 2001

By T. R. Yu, D. B. Counter

"Kidd Creek Mines primarily employs sublevel blasthole stoping for underground mining operations. During pillar recovery in No.1 Mine and panel mining in No.2 Mine, the free standing fill exposures may reach a height of 100 m or more. Consequently, a competent consolidated rockfill has been required to fill the mined openings.Since 1976, a total of more than fifty stopes and pillars have been filled with over four million tonnes of backfill. Fill exposures, up to heights of 100 m, have stood up well after blasting adjacent to the backfill.The fill material consists of a graded aggregate, Portland cement and ground slag. Sand has been occasionally mixed in the fill material to serve as avoid filler. The aggregate is prepared in a fill crushing plant, transported underground through raises, and hauled to the filling stopes by conveyors in No. 1 Mine and by trucks in No.2 Mine.Two mixing methods are currently in use. No. I Mine utilizes the free fall of gravity; the fill materials are sprayed with cement slurry as the aggregate departs from the conveyor into a mixing culvert. The other is the use of a rotary mixer on a batch basis in No. 2 Mine.This paper describes the fill systems, filling operation s, and the backfill quality control practice, together with various testing results related to improving fill quality as well as reducing fill costs.IntroductionThe Kidd Creek minesite, located 27 km north o f the city centre of Timmins, Ontario, produces 4.5 million tonnes of copper, zinc, lead, and silver ores annually from its No. 1 and No.2 underground mining operations. Both mines employ a sublevel blasthole stoping method."

Jan 1, 1983

By Ike Isagon

Backfilling has become an integral process in underground mining operations, with varying cost-effective forms of application. This study is a part of an overall investigation into the backfill practices at Copper Cliff North Mine (CCNM). The mine is well-diversified in terms of backfill methodologies and practices. The mine employs different fill methods such as cemented and un-cemented rock fill, slag fill, and cemented hydraulic fill (or a combination of them) in order to fill the voids created by mining activities. The history of backfilling at CCNM has been dominated by cemented and un-cemented slag fill. Each fill method requires operating and engineering guidelines for a cost effective quality control system. The purpose of this paper is to briefly discuss the backfill methods and applications, operating guidelines, and engineering requirements at CCNM.

May 1, 2011

By D. Bronkhorst, J. Rheault

"Backfill practices at the Williams mine have evolved significantly since underground mining began in 1986. The use of conveyors as a means of transporting fill material to stopes has been discontinued in favour of load-haul-dump equipment. The method of adding binder to the fill product has been refined along with the placement technique. Slurry preparation has been fully automated and flyash is now included as a binding agent. Surface preparation of the fill product has evolved substantially with the quarry operation coming into full operation and the construction of pennanent crushing, screening, and storage facilities. Company personnel have replaced all contracted labour involved in backfill operations. Some changes have been made to achieve a higher standard of quality control while others were implemented as a matter of cost reduction. BackgroundThe Williams mine is one of three mines operating at the Hernlo gold deposit, located 40 km east of Marathon, Ontario, along the north shore of Lake Superior. Daily production at the mine is 6000 Ud yielding an annual gold production of approximately 500 000 oz. The orebody extends from surface to a depth of 1300 m dipping north at 70 degrees. The average orebody width is 25 m, but varies from 45 m on the eastern boundary to about 10 m at the western probable limit and at depth. Proven and probable reserves currently stand at 32 million tonnes grading 6 g/t.The Williams orebody has been divided into four separate mining zones: the A-zone, the West-of-Dyke zone, the B-zone, and the C-zone (Fig 1). Ore production originally was from the A-zone open pit operation while the A-zone and B-zone underground workings were being developed. The mining method employed underground is blasthole open stoping with delayed rockfill. Mining block dimensions are nominally 20 m along strike by 25 m high. Stope extraction is sequenced to achieve a relatively short pillar life and 100010 ore recovery. Pillars are extracted as soon as the adjacent primary stopes are two cycles ahead. Primary stopes are filled with consolidated rock fill after mining. This allows pillars to be blasted right against the consolidated fill, but also requires that the fill be self supporting to a height of 30 m.The Williams mine places approximately 1.4 million tonnes of rockfill and 50 000 tonnes of binding agents each year, with about three quarters of the fill placed being consolidated. The consolidated fill placed at the Williams mine contains Portland cement and flyash as binding agents."

Jan 1, 1994

By F. E. Patton

"IntroductionAT NORANDA, the strength of the ore and wall-rocks per-mitted mining in large open stopes. While some of the orebodies were small and could be mined out in one stope, in most cases their size was such that it was essential that ore pillars be left during the first stage of mining. In these cases, backfilling has been necessary for ground control and to facilitate the later recovery of the ore in these pillars.This paper discusses the development of the •backfill and •its use in pillar recovery.GeologyThe ore bodies at Noranda occur as replacement deposits in a block of rhyolite flows and breccias, bounded on the north by the Horne Creek fault and on the south by the Noranda Andesite fault. Mineralization consists of pyrite, pyrrhotite, magnetite, chalcopyrite, sphalerite, gold, and various tellurides.The main orebody is a massive sulphide deposit designated the 'H' orebody. It is elliptical in plan, measuring roughly 400 feet by 800 feet, and is divided into two zones -Upper H, extending from surface to the 1,500-foot level, and Lower H, from the 1 ,500-foot level to just below the 3,000-foot level.Mining MethodsOpen stope mining in both Upper H and Lower H zones has been fully described in previous publications (1). A brief review only of the mining methods will be given here."

Jan 1, 1952