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By Gordon A. Gross

During the last half century major steel industries of the world have become dependent on iron and manganese resources derived from siliceous ironformation sediments. More than two billion years ago vast amounts of iron and manganese were deposited in ironformation sediments on all continents and until recent years their genesis was considered to be a uniquely Precambrian phenomenon. Research work has demonstrated that the origin of the banded ironformations was related to major tectonic features and associated volcanogenic processes that were active throughout geological history. Similar sedimentary-tectonic environments are being explored on the seafloor where metalliferous sediments are forming today. Ironformations are now considered to be a major part of the stratafer group of stratiform metalliferous deposits. They developed in sedimentary basins in areas where deep tectonic fracturing intersected high thermal gradients in the earth?s crust and gave rise to profuse volcanism and the hydrothermal effusive discharge of metals in the vicinity of the volcanic centres. Concepts for the genesis and metamorphism of stratafer deposits as revised in the last half century have had a profound impact in the development of iron ore resources and have had a dramatic impact in the use of metallogenetic concepts for guiding exploration for other mineral resources. They are based on the tectonic history and stratigraphy in the sedimentary basins, the role of hydrothermal volcanogenic processes , as well as environmental and sedimentation factors which influence the deposition of silica, iron , and other metals that were emphasized in the past. Topics of special interest in the genesis of ironformation and stratafer deposits and for guiding mineral exploration are: biomediation in their sedimentation and genesis, the significance of metamorphism in the processing and beneficiation of iron ore and in the genesis of iron ore by natural enrichment processes, syngenetic concentrations of Mn, Ti, Cu. Zn, Pb, Ni, Co, REE, F, Ba, B, As, Hg, W and Au in ironformations and stratafer lithofacies, and the mobilization of gold and other elements by alkaline solutions during diagenesis of ironformation and other stratafer metalliferous sediments.

May 1, 2003

By J. A. H. Church

My paper entitled "Spontaneous Combustion of Coal in Mines;" read some time ago before this Branch, consisted chiefly of extracts from the British Blue Book dealing with the same subject. Fallowing my article, Mr. de Hart submitted a paper dealing with the application of existing methods of working in Alberta to the conditions obtaining in our bituminous mines. I must confess that the perusal of his paper did not enlighten me either as to what might be a method preferable to those in vogue or even to Mr. de Hart's ideas on the subject, nor did the ensuing discussion of Mr. de Hart's paper afford any great clarification of the problem. The outstanding idea, supplied by Mr. Dick, seemed to be that a retreating system could not be employed, as no operator would feel justified in risking the capital necessary to driving to the boundary when the whole might be jeopardized by a fire breaking out in the waste. That a fire would break out in the waste of any adit-level working seemed to be conceded by all, and with that I agree; that an ignited waste could not be safely sealed off by means of stoppings in a coal barrier-pillar appeared to be the contention of Mr. Dick, or perhaps he attributed this idea to the Mines Branch-anyhow, it seemed to loom over the discussion like a miasma. With this aspect I do not agree, as ? reliance is placed wholly on stoppings in the thick seams of South Staffordshire, even in fiery mines. In the thick coal at Exhall colliery, or elsewhere in Warwickshire (34 feet 4 inches) containing hard and soft coal in layers, a fire assuming serious proportions is sealed off. By the way, retreating long-wall methods are employed. In Fifeshire, sealing off fire in the Dysart mine has been accomplished successfully. The seam averages about 20 feet thick.

Jan 1, 1924

By P. Ramirez

"To achieve an optimal metallurgical performance in the flotation process the bubble diameter should be between 0.5 and 2.5 mm. To control bubble size, frothers are added, reducing bubble size and conferring bubble stability against coalescence. However, the bubble generation mechanism determines predominantly the bubble size distribution produced. In mechanical flotation cells, bubbles are produced by the shear stress generated by a rotating blade on a gas stream. This mechanism produces a log normal bubble size distribution. In contrast, in column flotation, jetting gas injectors are commonly used in Chile and the bubble size distribution is bimodal with small and large bubbles. Bubbles larger than 2.5 mm mean a loss in surface area, while bubbles smaller than 0.5 mm move with the pulp stream and therefore collide less frequently with particles. In this paper, a new device based on a jetting sparger and a deflector disc (impinging jetting) is evaluated as an alternative sparger for column flotation.The new device was tested in a rectangular acrylic cell 1 m × 0.5 m × 3 m. High-speed cinematography was used to compare bubble generation with jetting and impinging jetting at different gas velocity in the nozzle (5–50 m/s). From the image analysis, it was observed that the deflector induces a shear stress, in analogy to a rotating blade but in this case the shear results from a high velocity gas hitting a static surface. The device was evaluated in a 0.5 m diameter 3 m high column. The McGill bubble viewer technique was used to compare the bubble size generated by jetting and impinging jetting. It was found that the Sauter mean bubble diameter reduced from 7.4 mm to 2.8 mm, when the jetting sparger was modified with the deflector. In the case of the impinging jetting sparger, 70% of the bubbles were in the effective range for flotation, while for the jetting sparger, only 10% were in the effective range."

Jan 1, 2016

By Laurier Juteau

"NOT TOO LONG AGO, the CIM Education Committee saw the need for an information campaign to promote interest in mineral industry careers among young people. Today, times have changed and the Committee has now placed emphasis on the definition and function of the mining engineer. This situation, we believe, is entirely normal and the change of emphasis is not due to a change in the needs of the industry but more to a diversification of these needs. In years past, the Canadian mineral industry, although a very important part of our economy, was looked upon as a specialized craft and, because of the richness of our resources, did not suffer too much from competition. Now, times have changed and we are increasingly at the mercy of the competition and damaging inroads of price fluctuations, etc. We will therefore answer the questions raised in this discussion in the light of the role played in the past and the role to be played in the future by the Canadian mining engineer. We will begin with the following premises : (a) If mining engineers are to be involved in the massive problems of society, their traditional background is not sufficient. (b) Councils for professional development are working on criteria for the accreditation of advanced degree programs in the United States as well as in Canada, and, simultaneously, three-year technology programs are being developed along with professional engineering degree programs. (c) The difference between European and American industries regarding philosophy of life is that in Europe they appreciate wisdom more than HPs while in America the HP seems to be more important. The Engineering Digest of February 1972 reported that in the United States, during June and July of 1971, the second highest rate of unemployment was for the 55 to 64 age group ( 4.1 per cent). This survey certainly applies to Canadian industry as well. (d) Degrees to be awarded in Canadian universities and technical schools in 1972 will be:"

Jan 1, 1972

By Yuki Koshino, Kenji Kaneko, Yasuhiro Aruga, Takuya Maeda

"In general, there is a strong correlation between dislocation characteristics and mechanical properties of aluminum alloys. In this study, dislocation characteristics of tensile deformed Al-Mg alloys were examined by transmission electron microscopy. An Increase in the solute concentration, ranging between 0.3 and 1.1%, Mg leads to increase in strength and decrease in uniform elongation. The accumulation of dislocations during the tensile deformation is observed in lower strain as the concentration of solute Mg in the alloy is higher. It is believed that the solute Mg in Al-Mg alloy promotes the accumulation of dislocation during the tensile deformation resulting in the decrease in elongation.INTRODUCTION The influences of the amount of Mg concentration on the tensile properties of Al-Mg alloys has been investigated by many researchers. It has been understood that the tensile ductility of Al-Mg alloy is higher for the Al-Mg alloys with higher Mg concentration beyond 2 mass% (Mukai, Higashi, Hirano & Tanimura, 1993). Uchida and Yoshida (1995) reported the formation of microbands during the tensile deformation due to the localization of dislocations, and the amount of microbands were reduced when the Mg concentration of the Al-Mg alloy was increased from 3 to 9 mass%. Their study suggested that the dynamic recovery of dislocation was retarded due to the increase of the Mg concentration during the tensile deformation, which resulted the increase in ductility. On the other hand, in the case of Al-Mg alloys with the Mg concentration less than 3 mass%, the reason why the ductility of Al-Mg alloys decreases with increasing Mg concentration has not been sufficiently investigated. In general, the Mg concentration used for automotive usage is less than 1 mass%, in the case of Al-Mg-Si alloys (Aoki, 2013). Therefore, to clarify the roles of Mg in the low concentration range on ductility of aluminum alloys and to elucidate its mechanism contribute to alloy design of Al-Mg-Si alloys having good formability."

Jan 1, 2018

By Ronald E. Pike

"PREFACEAt the CIM Annual General Meeting in Quebec City, R.E. Pike of The Toronto-Dominion Bank described how the Canadian banking industry worked with the minerals industry to assemble the large capital requirements for resource projects in Canada and overseas. He developed a profile of the mining industry and Canadian banking, both being important sectors of the Canadian economy and global in context. He also outlined the banking and financing partnership that existed in the past. In looking to the future and challenge that lies ahead f or each industry, Mr. Pike questioned whether the financing practices of the 1970s will be acceptable over the next decade. In that regard, he highlighted the risks that impact on each industry and illustrated how that financing partnership might evolve in response to those forces. The folio wing is a synopsis of his paper.IntroductionThe Canadian mining and metals industry is prominent in the domestic economy and respected internationally. Any measure of its performance illustrates that status-percentage of GNP, employment, exports, capital investment, size relative to other producing nations, etc.Energy, Mines and Resources Canada has forecast that the non-fuel minerals industry will grow in real terms by an average annual rate of 3.0% over the period 1981 to 1991. Using the GNP as reference, that growth in non-fuel mineral production will represent 7.7% of GNP. In translating that growth to capital expenditure, the EMR Canada estimates the aggregate amount at $43.8 billion ('79 dollars) or approximately 12% of total business investment in Canada during the 10-year period.About half of the $43.8 billion will be devoted to the ferrous sector-$8.4 billion in iron ore mining and $12.4 billion in the steel industry. Capital investment in nonferrous s projects is estimated at $13.3 billion, and it is projected at $6.7 billion for industrial minerals. The residual $3.0 .billion is exploration spending."

Jan 1, 1982

By V. J. Hucka, J. H. Tu

"A computer model has been developed for evaluating the performance of truck and shovel operations in an open-pit mine. The data required for describing the haulage network are simplified and shift startup and ending procedures specific to any given mine can be incorporated into the model readily . The model uses a FORTRAN-based simulation language SLAM and can be tailored to the conditions and practices of a particular mine.The model is used to analyze truck and shovel operations, evaluate different haulage layouts in the planning stage, or compare the expected performance of alternative truck dispatching systems.Stochastic simulation is used to model the random variation in time required f or the individual activities. This gives a more realistic representation of the processes under study. Simulation runs show good agreement with actual performance. The model is especially useful for examining alternatives in equipment acquisition , haulroad design, truck-shovel allocation and dispatching methods. The results of such simulations provide a sound basis f or making major management decisions.IntroductionTruck haul age is widely used in open-pit mining operations. It is also the largest cost item in the operating costs of an openpit mine, in some cases accounting for an estimated 50% of operating costs(1) , Truck replacements, which are necessary from time to time, involve large amounts of cap ital. Increasing depth of mining and consequent increase in stripping ratios require additional capital investment in the truck-shovel fleet. The sharp rise in energy costs over the last decade has also contibuted significantly to an increase in truck fleet operating costs . All these factors have led mine operators to search for ways to improve the effectiveness of the truck-shovel fleet, in order to lower costs and maintain a profitable operation in the face of declining markets and increasing world wide competition.Control of truck haulage in an open-pit mine is usually done in one of two ways."

Jan 1, 1985

By D. Sbarbaro

In many modem concentrators the use of supervisory systems in their grinding circuits has meant more consistent operations and increases in capacity. These systems rely on a series of specific algorithms that uses different pieces of information and drives actions over the process. One of these algorithms is the ball mill overload detection, which plays an important role in any supervisory strategy designed to maximize ore feed rate in wet grinding circuits. This paper introduces different overload detection algorithms, and compares them in terms of the time taken to detect the overload condition. This comparative analysis is based on highly noisy power draw measurements obtained in an industrial plant.

Jan 1, 2007

By L. M. Nolan

Chemical Mechanical polishing, or CMP, has four components in the polishing interface; namely the polishing pad, the polishing fluid, the abrasives in the fluid, and the surface being polished. Understanding the way that load is transferred between these components is key to understanding and predicting wafer polishing. Numerous models of CMP and the different load sharing arrangements which occur between these four components have been proposed. These include models by Luo and Dornfield (1) and Fu, Chandra and colleagues (2). Other models focus on just one or two loading arrangements, such as `hydroplaning' when all load is transmitted by the liquid in the case of Runnels and Eyman (3). In this work, we consider four possible combinations of load transfer from pad to wafer in CMP. These combinations are listed below in Table 1 and are illustrated schematically in Figure 1. It is important to note the role of scale in this process; the illustrations in Figure 1 are drawn to scale from actual profilometer scans obtained over a length of 5 with particles of 150 nm in diameter. Table 1 - Load Transfer Combinations Load Regime Illustration Wafer - Fluid - Pad Figure la Wafer - Particles/Fluid - Pad Figure lb Wafer - Particles - Pad Figure lc Wafer-Pad Figure ld The interest in load-sharing between the system components is because some configurations are anticipated to achieve higher polishing rates than others. For instance, in hydroplaning, or hydrodynamic lubrication, the lateral force exerted on the polished surface by the particles can be no higher than the fluid shear force developed in the gap between the pad and wafer. This is not expected to generate high rates of material removal from the surface (4). For contact that is largely between the pad and the wafer, such as the regime depicted in Figure ld, the lateral load is transmitted to the wafer by the pad and is hence limited by the pad's low rigidity. This regime is not expected to generate high material removal rates either. The highest levels of material removal are expected from regimes where the mechanical load exerted on the system is transmitted by the particles, in combination with either the fluid or the pad. Under what circumstances does the polishing regime change from one to another? What must occur in order to initiate or terminate contact between the system components? This is an ongoing question in the field of tribology, and has traditionally been viewed in terms of the pressure and velocity applied to the system, the thickness of the fluid film between the two surfaces, and the roughness of those surfaces (4). In this paper, we propose a new mechanism ? that the load on the particles may be sufficient to totally embed them in the polishing pad, at which point they are no longer able to transmit load and polish the wafer surface. This is illustrated schematically in Figure 1(d).

Jan 1, 2010

By O. A. Ojo, O. A. Idowu, N. P. Wikstrom, M. C. Chaturvedi

Transient Liquid Phase Bonding (TLP) of a Nickel base superalloy, Inconel 738, was performed to study the influence of holding time, filler gap size, and temperature on the joint microstructure. Insufficient holding time for complete isothermal solidification of liquated insert caused formation of eutectic-type microconstituents along the joint centerline region. An increase in the eutectic thickness with increase in initial filler gap size was observed in joints prepared at a constant temperature and holding time. In agreement with prediction by conventional TLP diffusion models, an increase in bonding temperature for a constant gap size, resulted in decrease in the time, tf, required to form a eutectic-free joint by complete isothermal solidification. However, a significant deviation from these models was observed in specimens bonded at and above l 175°C. A reduction in isothermal solidification rate with increased temperature was observed in these specimens, such that a eutectic-free joint could not be achieved by holding for a time period that produced complete isothermal solidification at lower temperatures. Considerable enrichment of the liquated insert with base metal alloying elements at higher temperatures is suggested to be an important factor contributing to the observed departure from conventional TLP prediction.

Jan 1, 2006

By R. Bruce Graham

"For the purpose of this paper, the Chibougamau area is taken as lying between lats. 49°20’ and 50°10’ N;-longs. 73°45’ and 75°30’ W. It comprises an area of some 4,000 square miles and is centred 300 miles north of Montreal.The first recorded report of metal deposits in Chibougamau was that of James Richardson in 1870-71 (I). In 1903 discoveries of asbestos, copper, iron, and gold made by Peter McKenzie stimulated considerable activity in the district. By 1909 interest had become so marked that considerable pressure was brought to bear on the Quebec Government to build a rail-way into the region. It was finally decided to have an authoritative opinion on its economic possibilities, and the Chibougamau Commission was appointed for this purpose (2). They examined the district in 1910 and concluded that at that time a railway was not warranted. Nevertheless, the district continued to be prospected and from time to· time gold and copper discoveries were made. Thus the Merrill Island deposit was discovered in 1920, the Campbell Chibougamau Cedar Bay deposit in 1922, Obalski in 1928, Opemiska Copper in 1929, and Norbeau in 1930.Difficulty of transportation was a severe detriment to the development of the region. Consequently the Quebec Department of Mines completed a road into Chibougamau in 1949. Further activity was encouraged. The Chibougamau Explorers, Quebec Chibougamau Goldfields, and Kayrand deposits were found in 1950. ** The copper-siderite zones on Copper Cliff ground were found in 1952 and 1953. Opemiska Copper, Campbell Chibougamau Mines, and Chibougamau Explorers were encouraged to prepare for production following favourable exploration results."

Jan 1, 1949

By M. P. Brameld, G. Archambault, R. H. McCorkell

"The rates. J. at which materials emit radon are of interest in studies of the sources of radon in mines. buildings. soils overburden and other places where it may indicate the location of ore or be a health hazard. Here J values were determined by measuring the radon concentrations by means of scintillation counters or alpha-particle track detectors in chambers in which specimens were sealed. Errors in the measurements due to leakage from the chambers. lack of secular equilibrium in the decay series of the radon measured, or the limited rate of diffusion of radon from powders or within the chambers were shown to be insignificant or could be corrected. Scintillation counter measurements are highly reproducible; track detector measurements also appear useable and may be made on many small samples using small, simple chambers. J measurements were made for uranium ores, drill cores, standard CANMET pulps. concrete and gypsum board. Powdering of ores had little effect on their J values.IntroductionIn the mining and mineral sciences, radon (222Rn) is of interest chiefly for two reasons: in geochemical exploration for mineral deposits its presence in soil. overburden and water in more than background concentrations may indicate the existence and locations of uranium deposits and, if present in sufficient concentrations, it is a possible health hazard in mines, mills and other places where people live and work. The latter aspect of radon's role has received considerable attention recently(1,2,3,4,5), and surveys of soil-gas radon concentrations have been used for some time in searches for uranium deposits(6,7,8,9,)."

Jan 1, 1981

By Howard J. Handewith

"Tunnel boring machines have been tried in underground mining applications as early as 1957. Although tunneling machines hardly threaten to supplant traditional drill-blast-muck operations for drifting, the latest modern machines offer substantial improvements in progress rates over traditional methods under certain conditions.An evaluation guide is offered which identifies conditions that favour machine boring. Different types of costs related to the use of a tunnel machine are discussed. The advantages of the tunnel machine over the traditional method are reviewed. IntroductionThe first tunnel boring machine ever used in a mine was in Quebec in the year 1957. Although less than successful at the time, this experiment served to illustrate the mining industry' s willingness to try new ideas. Since 1957, the manufacture of tunnel boring machines has developed into a full-size industry. Most contractors in Europe and North America will now consider the use of a tunnel boring machine first on any long tunnel that can have a circular section.The modern tunneling machine offers a great potential to the mining industry. Just as the blasthole and raise drill machines have revolutionized mining in the last decade, it is expected that tunnel boring machines will make an equal impact in the future.This is not to suggest that tunnel boring machines will replace explosives in the foreseeable future, but they can supplement drill-and-blast operations. In areas where long drifts, inclines, declines or cross-cuts are needed in a hurry, the tunnel boring machines offer the fastest way to a production return on investment. Figure 1 shows the range of production that two types of tunnel machines have produced as compared with conventional excavation methods in various ground conditions.Rotary hard-rock tunnel boring machines (Fig. 2) commonly produce progress rates in excess of 20 metres per day in uniform ground from 300 to over 2,000 kgl cm2 in unconfined compression strength . These averages are for the completed tunnels. One 7. I6-metre-diameter tunnel boring machine just completed a 1O.7-km tunnel in hard rock of 700 to 2,800 kg/cm>, The last eight months of production averaged 30 metres per day."

Jan 1, 1980

By Nicola Robjohns, Kenneth A. Grace, Jane Spooner

"The important volume end use sectors for the rare earth elements and yttrium continue to be in steelmaking, petroleum cracking catalysts and glass polishing. The market sectors which may provide future high volume demand, to replace the older sectors where demand has declined over the past decade, include magnets, automobile catalysts, ceramics and superconductors. Increases in production capacity in all major rare earth producing countries are expected to restrain large price increases for high purity rare earth oxides over the short term. A sharp increase in demand for yttrium in advanced ceramics and superconductors would lead to a shortage in separation capacity, however, and to a corresponding rise in prices for separated rare earths.IntroductionWhile visiting the Geology Department of the University of Toronto in 1981, Professor Lu from the Institute of Geochemistry of Academia Sinica at Gweiyang in China introduced one of his seminars with the observation that in China, rare earths are not rare at all.Indeed, while being rare in terms of abundance as measured by, say, the Clarke scale, the rare earth elements, like many other low abundance elements, are subject to geological, geochemical and mineralogical concentration.The rare earth elements comprise the fifteen inner transition elements of Group IIIB, Period 6, of the periodic table, from lanthanum to lutetium (the lanthanides), as set out in Table 1. The rare earths constitute a distinct group as they have extremely similar chemical characteristics due to the uniformity of their outer electron shells. The element yttrium is usually, by convention, included in the rare earth group as it occurs with the rare earths in nature and has properties very similar to them.The rare earths are divided into two subgroups, the ""light"" rare earths, made up of the first seven elements of the series, lanthanum to europium, and the ""heavy"" rare earths, comprising the remaining eight elements, gadolinium to lutetium, plus yttrium. These two subgroups are also referred to as the cerium subgroup and the yttrium subgroup, respectively. The subdivision is based on differences in the atomic radii resulting in somewhat different properties for the two subgroups."

Jan 1, 1991

By Adrianus (Adrian) Halim, Tomas Bolsöy, Stina Klemo

"This paper describes the ventilation system used in underground hard rock mines in Sweden and Finland, which is very different than the system used in other major mining countries such as Australia, Canada, South Africa, and USA. The system utilizes auxiliary fans with Variable Speed Drive to distribute primary airflow to working levels instead of regulators that are commonly used in other countries. Descriptions about Ventilation on Demand in this system and ventilation regulations in Sweden and Finland are presented in this paper. This paper also quantifies a comparison between the Nordic and the system used in other countries. KEYWORDS Auxiliary fan, Regulation, Ventilation on demand (VOD), Variable speed drive (VSD) INTRODUCTION The primary objective of an underground mine ventilation system is to provide sufficient airflow to dilute atmospheric contaminants that are produced by mining activities. Typically, a mine ventilation system consists of one to four primary fans, regulators to control primary airflow distribution to levels or panels, and auxiliary fans that distribute airflow to dead-end workings via ducts. Depending on circuit geometry, underground booster fans are sometimes used. This system is widely employed in many countries such as Australia, Canada, USA, United Kingdom, and South Africa. However, mines in Nordic countries such as in Sweden and Finland employ a different system, in which auxiliary fans instead of regulators are used to control primary airflow distribution, as well as distributing airflow to dead-end workings. This paper describes the general layout of Nordic ventilation system, its VOD and heating system, its comparison with the system that is widely used in other countries which is referred as “Non-Nordic ventilation system” in this paper, and the current ventilation regulations in Sweden and Finland. The comparison was quantified by doing simulations using Ventsim Design 5 software. DESCRIPTION OF NORDIC VENTILATION SYSTEM General System Layout Figure 1 shows a schematic of typical ventilation system used in Swedish and Finnish mines. The schematic was drawn by Franzen, Myran, Larsson, and Rustan from Stiftelsen bergteknik forskning (Swedish Rock Engineering Research Foundation) and the English translation is written next to the Swedish terminologies. The system is generally a push-pull system. None of Swedish and Finnish mines use exhaust system. As shown in Figure 1, the primary airflow is distributed to levels by auxiliary fans that are bolted to a bulkhead located in the access to intake raise/shaft. These intake fans are attached to ventilation ducts that distribute airflow to all working faces. The exhaust auxiliary fans, which are not attached to any ducts and also are bolted to a bulkhead located in the access to exhaust raise/shaft, then flow the exhaust airflow into the exhaust shaft. A small quantity of this exhaust air is directed into the ramp, which acts as another main exhaust airway. The deepest level is also equipped with booster fan(s) that draw air from the intake shaft/raise and direct it into the ramp to provide fresh air in the ramp. There is no free split in this system. Figure 2 shows a schematic of a level ventilation layout in the Nordic system. Figure 3 shows a photo of these auxiliary fans."

Jan 1, 2019

By P. Chryssanthakis

"The aim of this study is to give a better understanding for the behaviour of the composite material RRS and the synergy effect of its components while an excavation face advances during tunneling. The lower part of the Q-system scale (i.e. Q-values of 0.4 or less for a tunnel span of 10 m) is commonly used for tunnel applications in softer rock formations. This is implied in the updated Q-chart, where the dimensioning of rock support by using Reinforced Ribs of Shotcrete (RRS) comprises the use of fiber reinforced shotcrete, S(fr), rock bolts and shotcreted beams reinforced with steel rebars. The newest part of the lower part of the Q-chart gives more information on required support compared to older Q-system versions. The RRS can be described as a three part composite material with the 3rd component being the reinforced concrete beam (with no fibers) and with steel rebars that are placed in groups and sprayed with pneumatically applied shotcrete in layers. The RRS usually have a total thickness varying between 0.20 and 0.50 m and a width between 0.25 and 0.60 m respectively. They are placed at different distances along the tunnel axis depending on the rock mass quality.In this study, an effort has been made to simulate RRS in a numerical code and to study its behaviour in a real case study. A double tube, triple lane tunnel in weathered tuff has been chosen as reference material for testing RRS under changing load. This was done by writing a FISH subroutine which is implemented in the discontinuous numerical code Universal Distinct Element Code (UDEC) that takes into account the characteristics of the reinforced beam (height, width, number of steel rebars, and spacing) in the calculation process. A rather extensive parametric study has been run by using the same key input rock mass parameters and in situ stresses (?H/?v = 0.5), and changing only the RRS key input parameters in order to simulate the changing loads during the face excavation process in tunneling. The critical part of this numerical modelling work that is under investigation in this study, is the period just after the excavation: i.e 4 hours; 16 hours after excavation; and finally after the complete curing (28 days) of the cement based components. The RRS are installed in three consecutive steps as follows: a) the application of S(fr) on the rock surface in order to even it up for the RRS application; b) radial bolting; and c) installation of the reinforcing steel rebars, that form the reinforced beam.The results show that: a) the S(fr) is almost overloaded in the first 4 hours after application in both tunnels; b) on the S(fr) and steel rebars, the stresses, axial and shear loads, as well as moments are decreasing after 16 hours and 28 days; and c) the bolts loads are increasing with time. The observed reduction in stresses and loads in both the S(fr) and reinforced beam is surprising. A possible explanation is that tunnel arches are self-supported to a lesser extent as time progresses due to redistribution of stresses in the nearby rock mass through the radial bolts. The loads in radial bolts are increasing by almost 25%. This can be seen from the overall results as we observe an increase in maximum principal stresses, displacements, shear displacements along joints and hydraulic apertures compared to tunnel baseline that has no support of any kind."

Jan 1, 2015

By Eric D. Jamieson

"How well is the Canadian coal industry able to expand in the tough competition with other fuels in Canada and with other suppliers in the international, metallurgical and thermal coal trade? Coal production as a whole decreased between 1950and 1960 and, although there has been a tripling of production since then, underground production continued to decrease until 1975. This trend has been reversed to a certain extent as a result of the modernization of mines in Cape Breton, but in 1978 over 85% of coal production came from surface mining.Against this background, the paper reviews current Canadian mining operations as described in papers presented at the 1979 CIM Annual General Meeting. Conclusions are then drawn regarding some of the longer-term mining applications in relation to the latest coal demand forecast, which indicates a potential Canadian production requirement of 160 million tonnes per year by 2000. IntroductionThe Canadian 'coal industry is experiencing a period of expansion and a new record for production is set with each passing year. This growth pattern over the last IS years has been based on two main markets-the generation of electricity in Canada and the export of coking coal to Japan. Supply to those markets has entailed either the development of mining, coal preparation and transportation systems for the delivery of large volumes of products with predetermined specifications or the establishment of minesite plants designed specifically to suit the characteristics of the coal in the adjacent deposits (Fig. I).Coal is a low-value commodity in relation to its bulk, and the economics of coal use become progressively less attractive with increasing distance from the producing mine. The transportation component in the export trade and in the movement of western coal to Ontario is a major cost factor. In the latter case, it approaches the cost of coal produced at the mine. In addition, the .international competition among coking coal producers chasing diminishing markets in a worldwide steel recession as well as the competition from U.S. coals and other fuel s for Canadian thermal coal have introduced a tight set of parameters to govern mining practices in all parts of the country. Some appreciation of the required changes in mining techniques brought about by declining traditional markets and the rigorous demands of the new bulk markets can be gained by comparing production from surface and underground mines over the past 25 years (Fig . 2)."

Jan 1, 1980

By J. S. Godard

The mining developments in western Quebec have shown the ores of this district to be more complex in character than was at first expected. When the initial discoveries were made it was thought that the ores would be similar in character to those of Kirkland Lake and Porcupine areas. Developments have since shown that although the ores along the southern margin of the belt are somewhat similar, those within the belt, constituting by far the greater proportion, are vastly different, being complex mixtures of base-metal sulphides carrying values in the precious metals. These mixtures of base-metal sulphides of copper, zinc, and iron, present several problems to the metallurgist, especially in their concentration into marketable products. For the purpose .of this paper the following classification may be made of the ores of the district. Group I. - Milling Ores: (a) Ores in which~ arsenopyrite is the predominating sulphide. In this class of ore the gold occurs in fractures in the quartz and is associated to a lesser extent with the arsenopyrite and pyrite. Examples of this class of ore are the Granada, south of Pelletier lake, and the O'Brien and the Thompson in Cadillac township. (b) Ores in which iron pyrite is the predominating sulphide and gold is the valuable constituent. This may also be classed as a concentrating ore-Group III.

Jan 1, 1928

By Adrian Dance

Newmont Mining Corporation?s Carlin (?Carlin?) operation processes roaster feed through a double rotator, dry grinding mill following secondary crushing. Metso Process Technology & Innovation (?PTI?) was involved in a review of crushing and dry grinding circuit operations in August 2009. A number of opportunities for increased throughput were identified in the PTI study. Extended plant trials were conducted in July 2010 to provide justification for possible circuit modifications and if needed, additional equipment. The trials confirmed that a finer crusher product would reduce the build-up of material ahead of the coarse grinding chamber in the double-rotator. Also revealed were material transport and airflow restrictions in the dry grinding circuit that resulted in a high circulating load. The secondary crushing plant demonstrated it was capable of reducing the amount of +¾? material in the mill feed thus increasing dry grinding throughput. However, wet and sticky feed affected screening efficiency resulting in low utilisation of the existing crusher capacity. The open circuit configuration also made it difficult to control crushing circuit product size. For the dry grinding circuit, changes in grinding media size, transfer grate design and increased airflow were recommended to reduce circulating load issues and increase circuit capacity. This paper summarises the crushing and grinding circuit review, the extended site trials performed by PTI and the improvements in dry grinding operation that resulted from this study.

Jan 1, 2012

By N. Erarslan, M. Serati

"This paper attempts to provide a simple analytical solution for the calculation of the stress tensor induced in a two-dimensional elastic ring of a brittle material subjected to general loading on either its internal or external curved surfaces. Airy stress functions and Michell’s expression in polar coordinates are used, and parametric studies are performed to investigate the influence of loading angle and the geometrical aspect ratio of the ring on the maximum induced tensile stress. INTRODUCTIONA proper and accurate estimation of the tensile strength of a given material is a crucial step in many engineering projects; in particular in rock cutting, since rocks are much weaker in tension than in compression or shear. Both direct and indirect laboratory techniques such as the direct uniaxial pulling test, the Brazilian indirect tensile test (also known as the split cylinder test), and the ring test have been proposed and developed to estimate rock tensile strength. Among these, the ring test, initially suggested by Ripperger and Davids (1946), has become particularly appealing for its ease of sample preparation and its unique breakage mechanism in almost a pure tensile mode. Unlike the Brazilian test in which high shear stresses developed in the vicinity of the contact could interfere with the tensile breakage of the disc specimen, a tensile failure starts away from the loading platens in a ring test (Hobbs, 1964)."

Jan 1, 2015