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By Bong Sun You

AZ31 magnesium plate of 300mm in width and 40mm in thickness was produced by horizontal continuous casting machine, and its microstructures were analyzed on the surface and in the plate interior. The defect formation on the surface and subsurface, such as side cracks, buckling, pores and segregation, was very sensitive to the casting condition and processing parameters. Especially, casting speed and secondary cooling were the most dominant parameters to control the solidification behavior of magnesium melt, for example the position and shape of the solid/liquid interface in the mould. The cyclic marks on the surface were obvious because the plate were drawn by 4-step cycle and can not be removed. The average grain size of as cast plate measured as 50-65µm at the near surface and 300-400µm in the center of the plate was relatively small because secondary cooling water was applied to control the surface quality resulting in rapid solidification of the melt in the mould.

Jan 1, 2006

By S. A. Bunk, E. C. Eccleston, J. J. Amsden

Fluidized bed technology has been widely applied to mineral and metallurgical processes. Over the years, modeling of fluidization technology has progressed to a level at which basic design parameters can usually be defined based on empirical correlations. In addition, scale-up considerations have become better understood, and significant advances have been made in the area of computer simulation. Despite these advances, fluidization is a complex phenomenon, and modeling tools remain limited in their ability to simultaneously model both fluidization and chemical process parameters. As a result, almost all mineral and metallurgical applications of fluidized bed technology require pilot testing prior commercial scale design. This paper outlines pilot plant testing requirements and process development considerations for development of low technical risk commercial designs.

Jan 1, 2011

By Craig Haase

A. Introduction

Jan 1, 2003

By P. J. Mackey, Elsie Kossatz

The small town of Bruce Mines lies a few miles east of Sault Ste. Marie, Ontario on the north shore of Lake Huron. This was the site of the first copper mine. This copper smelter used a reverberatory furnace, based on the Welsh process, and operated in the period of 1849 to 1850, when the plant was destroyed by fire. During the period of 1850 to about 1869, crushed and sorted ore was shipped to England and later to the United States. In 1869, a roastleach plant, based on the Henderson Process, was constructed at Bruce Mines and was operated until about 1872. This short article describes the early development of this historic operation. Although the rich copper ore found in the vicinity of Bruce Mines was known to the Indians, its discovery in or about 184 7 was credited to Mr. Keating, who named the area Bruce in honour of Lord Elgin (James Bruce), Canada's Governor General. In 1848, the property was acquired by the Montreal Mining Company for 40 000 pounds sterling and the company immediately sent a small party to investigate the deposit. The party sank a shaft on a lode within a hundred yards of the lake and also reported two adjacent lodes farther inland, which were left undeveloped until the turn of the century. Initially, the foundation of the headframe and hoist was built above the frost line and when the machinery was set in motion it collapsed. Eventually, the shaft was to reach a depth of 300 feet. By 1849, the property was well established and the small settlement consisted of 77 miners, 65 labourers, 4 boys, 11 tradesmen, including blacksmiths and carpenters, 2 mining captains, 1 engineer, 2 clerks and 1 superintendent. The community, including the workers' families, numbered about 250.

Jan 1, 1999

By Nelson C. White

This paper summarizes the present state and future prospects of the potash industry, with special reference to its impact on production in Saskatchewan.

Jan 1, 1968

By Tony Breuer

Can international development NGOs and the mining industry find common ground and work together in the Third World? The Canadian Hunger Foundation/PARTNERS in Rural Development (CHF-PARTNERS) has, over the past forty years, implemented more than 800 rural community development projects in some 38 countries of Africa, Asia, and the Americas. Working with or for the Canadian mining industry offers both opportunities and risks. We can help a mining company ensure that the workers, their families and communities share adequately in the economic and social rewards from a mining operation in their area. We can contribute to the provision of sustainable livelihoods when an operation winds down. The company?s reputation as a good corporate citizen can be enhanced. On the other hand, mining companies and NGOs have generally very different values and missions. The risk of misunderstandings from conflicting motives and work practices is correspondingly high. Both partners must state their values and expectations clearly at the start of collaboration. The NGO must ensure that the company abides by acceptable ethical standards. It must not directly or indirectly subsidize a private company from its charitable donations. The mining company must satisfy itself that the NGO has the capacity, resources and experience to produce the results expected.

May 1, 2001

By S. Razavialavi

In the planning phase of every construction project, layout of temporary facilities is a crucial task; site layout can affect safety, travel cost and time, construction productivity, and space utilization. However, site layout planning can be a complicated problem, due to the interdependency of influencing factors. Although interaction among activities is one of the major drivers of site layout planning, it has not been properly addressed in past research. In this study, simulation is presented as a promising tool to address this gap. The capability of simulation technology to model complex processes in construction projects makes use of simulation tools in site layout optimization problems effective, while existing methods are unable to perfectly model these problems, in some cases. Additionally, the advantages and challenges of implementing simulation are assessed and a generic framework for simulation application in site layout planning is proposed.

Aug 1, 2013

By P. Van Marcke, X. Sillen, G. J. Chen, I. Troullinos, A. Dizier, J. Leysens, X. L. Li, J. Verstricht

"In Belgium, Boom Clay is considered as one of the potential host formations for the disposal of the High Level radioactive Waste (HLW). The URL HADES at 223 m depth at Mol site was constructed in order to study the feasibility of HLW disposal in Boom Clay. Since 1995, the URL R&D programme has focused on large scale demonstration tests like the PRACLAY experiment.The PRACLAY in-situ experiments include a set of three tests: The gallery and crossing test, PRACLAY Heater test and PRACLAY seal test. The main aim of the Heater test is to study at a large scale the thermal impact of a HLW repository on the host rock. The test is housed in the PRACLAY gallery. A hydraulic seal, made of compacted swelling and low permeable bentonite annular ring was installed at the intersection between the heated and the non-heated part of the gallery. The primary objective of the seal is to provide undrained hydraulic boundary conditions for the Heater Test.The PRACLAY gallery was constructed in 2007 and the hydraulic seal was installed in 2010. The bentonite is hydrated by natural and artificial process. The swelling, total pressure and pore pressure of the bentonite are continuously measured and analysed by numerical simulations to get a better understanding of this hydration processes. The performance of the seal is tested systematically by means of gas injection breakthrough tests and permeability tests at interface bentonite / Boom Clay and in the Boom Clay around the seal.The PRACLAY Heater will be turned on at the end of Oct. 2014, since a sufficient seal swelling has been attained to fulfill its role for the PRACLAY Heater test. This paper will focus on the PRACLAY Seal test: How we understand the process behind the hydration of the bentonite and how we test the performance of the bentonite seal."

Jan 1, 2015

By H. M. Roscoe

Fifty per cent. increased lump coal production and at the same time a reduction of from 20 to 40 per cent. in explosive consumption. These results were achieved by the use of the rock-dust cartridge method of stemming shots in the Ruhr coal mines, as announced by Dr. Alfred Stettbacher in the September 1923 number of the Explosive Engineer. Immediately, the explosive and the coal-mining industries on this continent became intensely interested. Considerable experimental work has been carried on in a number of American and Canadian mines during the past year, by which our engineers have become assured that the use of rock-dust cart-ridges for stemming shots increases the efficiency of explosives and promotes safety in coal mines.

Jan 1, 1924

By Sudarshan Martins

The measurement of acoustic signals produced by a tumbling mill has been previously used to determine the operating state of the mill and correlate the variations of the sound intensity with the operating conditions of the mill. Separately, the discrete element model (DEM), a software simulation of the mill, was developed to provide insight into the operation of the mill. Presently, there has not been much work reconciling the DEM simulation of the mill with the acoustic signal. Working from basic principles of linear, time-invariant systems, agreement between an acoustic DEM simulation and the measured signal is sought. First, the impulse response function of the mill is measured. The DEM provides the forces applied to the mill shell as a function of time. A convolution of these simulated shell forces and the transfer function gives the simulated acoustic signal. This signal is compared to the measured acoustic signal, both in the time domain and the frequency domain. Finally, a discussion of the results is given.

Jan 1, 2012

By G. Lafont

Aug 1, 2013

The dissolution of chalcopyrite has received a significant amount of attention. The predominant paradigm of investigation is that its rate of dissolution is limited or passivated by a surface coating or film of material. The common candidates for such a passivating coating are numerous: polysulfide coating, metal deficient layers, copper sulfide coating, bornite layers, jarosite precipitates, and elemental sulfur films. While this line of explanation, that is, the explanation based on passivation, seems to have won some form of consensus with researchers, there is another explanation for the results that can account for all the results observed. Crundwell (1988) proposed that the solid state properties of chalcopyrite, in particular, the semiconducting properties, are responsible for the observed behaviour of chalcopyrite. The purpose of this paper is to outline the semiconductor mechanism of dissolution, and to show how this mechanism accounts for all the observed phenomena recently reported in the literature. The paper outlines the semiconductor mechanism of dissolution, shows how this mechanism accounts for the dissolution

Jan 1, 2014

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. Chollet, A Constant

THE RESULTS of a previous investigation ( 1) on the work-softening of some heat-treated steels have led to the conclusion that the occurrence and intensity of the phenomenon is possibly governed by the work-hardening capacity of the material, and the relative amount of retained austenite. The first assumption is compatible with Polakowski's theory (2) that the absence of work-softening, in steels quenched and tempered at relatively low temperatures, may be attributed to the high rate of work-hardening which compensates for, and even out-weighs, the loss of internal energy associated with the simultaneous rearrangement of the original system of internal stresses. Indeed, Polakowski first considered the work-softening as a manifestation of a partial release (by cold-working) of the latent strain energy stored in

Jan 1, 1962

By A. F. Robertson

FOR several years, the rock-burst question has been given serious consideration by the operating staff at Wright-Hargreaves mine and, as a result of observations and experiences, several different methods of mining have been developed which are now in use in various parts of the mine. These are described in the following pages. It should be emphasized, how-ever, that these methods, and the views expressed in this paper, are based solely on conditions at Wright-Hargreaves mine. Such conditions may or may not be similar to those obtaining in other mines in the Kirkland Lake area, where ore occurrences, veins, wall-rock, faulting, intrusives, etc., may be entirely different. Size and shape of ore-bodies, composition of the ore-bodies and their surrounding rocks, locations of intrusives and sedimentary rocks, and direction, dip, and strength of faults, all introduce problems that are peculiar to each individual mine. Ground failure' in coal and metal mines is not at all unusual. A variety of systems of mining have been developed to delay movement of large or small blocks of ground, or, in some instances, to take advantage of the fact that ground movement, due to mining operations, is unavoidable. Ground failure may be either slow collapse or it may take place with explosive violence. It is principally with the latter condition that we are concerned in the mines of the Kirkland Lake district.

Jan 1, 1939

By L. Lamothe, C. H. The

"Pilot precipitator tests were conducted at Noranda's Horne Smelter in late 1978 to evaluate methods of improving converter dust collection efficiency. A bout 35 Nms/min. of converter off-gas were isokineticatly diverted from the converter flue to the pilot precipitator leased from Flakt (Canada) Ltd., consisting of three fields, one transformer-rectifier set per field and 77-m2 total collecting plate area. The effects of gas volume flow rate and conditioning the gas with sulphur trioxide and/or additional moisture were tested. II was found that gas flow rate and moisture level had the most pronounced effect on collection efficiency. Based on a techno-economic assessment of different options for improving converter Courell efficiency, waffle hoods and water sprays were installed on the converters to reduce air infiltration and increase the moisture level in the gas.BackgroundThe Horne Smelter of Noranda Inc. is situated in northwestern Quebec and, since the closure o f the min e in 1976, has been operating as a custom smelter. The concentrates treated at the smelter come from various mines with mixed sulphide ores , many o f them containing significant amounts of lead and zinc which makes dust collection in electrostatic precipitators mo re difficult than normal.As part of a continuing program on reducing smelter dust emissions, the No . 6 Cottrell was installed in 1972 up stream of Nos. 2 and 4 Cottrells (in parallel) in a n attempt to improve dust collection by increasing the over- all precipitator capacity ( Fig. I). However, it has never per formed a t its designed efficiency o f 94%. The average efficiency established in various test s was in the range of 60 – 65%."

Jan 1, 1985

By N. Hiroyoshi

In the leaching of chalcopyrite with sulfuric acid solutions, the copper extraction rate reaches a maximum at a given redox potential. This article reviews our studies on the determination of the optimum redox potential for chalcopyrite leaching. Shaking flask and column leaching experiments for a chalcopyrite concentrate were performed under various conditions, and the results were analyzed using a normalized redox potential defined from a reaction model assuming the formation of an intermediate CU2S from chalcopyrite. Factors, such as metal ion concentrations, solid/liquid ratio, and presence of iron-oxidizing bacteria, caused significant variations in the copper extraction versus time plots. However the copper extraction rate versus normalized redox potential plots were independent of the above factors and the copper extraction rate reached a maximum at the normalized redox potentials around 0.43. Converting the normalized redox potential to the solution redox potential gives the optimum redox potential for chalcopyrite leaching as a function of cupric and ferrous ion concentrations. This semi-empirical equation can be used to predict the optimum redox potential during leaching operation, and is useful to design the redox-controlled heap leaching for chalcopyrite.

Jan 1, 2007

By Erica J. Wanless, Ghislain Bournival, Seher Ata, Jose Hamilton Tavares

In froth flotation valuable particles are captured by bubbles and transferred to a froth layer where bubbles are in close proximity. Bubble coalescence frequently occurs in flotation, especially in the froth layer. The coalescence of two bubbles causes a release of energy, which may expel particles from the bubble surface. A series of experiments were carried out to determine the effect of bubble interfacial properties on the detachment of particles upon the coalescence of capillary-held bubbles of 2 mm in size. The interfacial properties of the bubbles were changed by the addition of low molecular weight alcohols, namely 1-pentanol or methyl isobutyl carbinol (MIBC). Using esterified glass particles (? = 43 °) with a 10 % and 90 % cumulative passing diameter of 47.2 and 92.5 µm, respectively, as model particles, it was found a certain concentration of frother, different for each additive, is needed in order to observe a reduction in the detachment of particles. In the presence of frother, the detachment was evaluated at approximately 65 % which was a considerable reduction in the detachment compared with the value of approximately 85 % found in the absence of frother. Frother molecules changed the properties of the airwater interface. Upon the coalescence of two bubbles, the release of energy caused the merged bubble to oscillate; effectively resulting in the detachment of particles. The addition of frothing agents augmented the elasticity of the bubbles, which appeared to be related to the detachment of particles. Whereas frothing agents are generally used to prevent bubble coalescence and stabilise the froth phase, only transient stability is imparted. The results suggested that an additional role of frothers in flotation is to decrease the detachment of particles from bubble surfaces. The type and concentration of frother can be strategically selected to improve metallurgical recovery as an economic incentive.

Jan 1, 2016

By M. P. Amsden

"Ecstall Mining Limited, a subsidiary of Texasgulf, Inc., Operates the Kidd Creek Mine, Concentrator and Zinc Plant at Timmins, Ontario. Two different ore types are mined and milled to produce concentrates of copper, zinc, lead and tin. The ores are processed at the Concentrator in three separate individual milling circuits, each with the capacity to treat up to 3500 tons per day. Two circuits treat copper-zinc ore 1t an average combined rate of 6500 tons per day and the third circuit treats 3500 tons per day of silver-lead zinc ore. Conventional crushing, grinding and flotation procedures are followed. The purpose of this paper is to describe Ecstall's experience with rubber mill linings in the grinding circuits.CrushingBoth ores are crushed to minus 6 inches at the Mine and delivered by rail to the Concentrator where they are reduced by secondary and tertiary crushing to minus 5/8 inch.GrindingThe crushed ores are prepared for selective flotation by grinding with rod mills and two stages of ball milling. flotation circuit middling products are reground in regrind mills prior to further treatment. A typical flowsheet of one of the grinding circuits is shown in Figure 1.All the grinding units are Allis-Chalmers Mills. The rod mills are 10 1/2 feet diameter by 16 feet long. Both the primary and secondary ball mills are 12 feet diameter by 18 feet long and all regrind mills are 6 feet diameter by 12 feet long. The primary mills are charged with 1 13/16 inch forged balls. Secondary mills are charged with 1 1/2 inch forged balls and the regrind mills are charged with 1 inch forged balls. All cyclones are Kreb units. The primary cyclones are Dl5B hydrocyclones, the secondary cyclones are D6B hydrocyclones and all regrind cyclones are D6B hydrocyclones."

Jan 1, 1974

By K. Kim

"All milling operations would like to achieve the best possible recovery in combination with the highest achievable throughput. Unfortunately, many operations have problems with low recovery, low throughput or both. In the majority of these cases investigated by the author, the solutions to improve recovery and throughput were simple. The solutions did not include implementation of new technology or major capital, but rather a return to fundamentals. In most cases, problems occurred because the operation was not following their own operating procedures, because the design of the circuit was changed without adequate change management or understanding of the implications of changes, or because basic metallurgical measurements and control were not taken or ignored. In this paper, examples of these problems and their solutions derived from long and varied experience will be discussed. The objective of this presentation is to provide the milling operator with practical and readily implementable solutions to common milling issues to improve throughput and recovery.INTRODUCTION Mill operations can be evaluated by many operating parameters, but the two best indicators are recovery and throughput. These two parameters summarize the wellbeing of an operation. The grade to the mill is also a good indicator, but this is usually controlled by the mine. The best recovery and throughput are achieved through a team work involving operators, maintenance workers and metallurgical staff. The milling operations include many unit operations, such as crushing, grinding, flotation, leaching, carbon in leach, Merrill Crowe, roaster, autoclave and refinery. The tools available to the operations are plant design, process measurements, assays, metallurgical balance, and operating strategy. Twelve real life examples of improvements of recovery and throughput are presented in this paper. EXAMPLE 1: CRUSHING CIRCUIT WITHOUT FINE ORE BIN At this operation plant throughput could not be maintained at a high rate because feed to the mill became coarse during the day. The cone crusher was feeding the ball mill directly. As the day wore on, the gap on the cone crusher would slowly increase, which would cause the product size to increase. This would increase the re-circulating load on the ball mill. The throughput would decrease accordingly. Furthermore, the throughput would decrease more with high frequency of maintenance on the cone crusher."

Jan 1, 2019