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By J. M. Harrison

FIELD PARTIES of the Geological Survey of Canada have, in recent years, mapped several areas east and northeast of Flin Flon (9, 10, 12, 13, 14, 15, 118) (1). In all these areas the mineral deposits show a close geographical relation to faults although the immediate control of most of the known ore deposits appears to be folded structures in the rocks (8, 11, 18, 19). Air photographs and published geological maps were studied in order to ascertain whether or not the mineral deposits in the Flin Flon-Snow Lake mineral belt as a whole show any spatial relationship to faults. It was found that most of them do show a distinct geographical association with faults, or with topographic features that may be faults. It is further apparent on air photographs that a major lineament, for composite of lineaments, follows approximately the contact between gneisses and schists of the Kisseynew complex on the north, and Amisk volcanic rocks on the south. However, this Kisseynew lineament is not continuous and, especially in the eastern half of the district, it is exceedingly irregular in plan. Parts of this lineament have been mapped as faults (9, 10, 12, 14), other parts have been shown as stratigraphic contacts (1, 2, 21, 27), and still other parts have not been studied except by reconnaissance methods. It is suggested that the Kisseynew lineament may be due to a deformed fault that extends easterly from a point about 25 miles west-northwest of Flin Flon to a point at least 10 miles northeast of Wekusko lake and that the mineral deposits.

Jan 1, 1951

By Ron Bradburn, Derek Perkins

"The Brenda mine is located in south central British Columbia, 22 km north west of Peachland, in the heart of the Okanagan Valley. Situated at an elevation of 1500 metres, the mine area straddles the divide between the headwaters of Trepanier Creek and Peachland Creek. These flow into Okanagan Lake, the chief valley water resource located 1100 metres below. The creeks are of great significance to the water management of the area as they constitute the water supply for the Municipality of Peachland, a thriving retirement community with a population of 5000. A location map is shown in Figure 1.The history of the Brenda Mine operation has been well documented previously. After twenty years of operation, the open pit mine shut down in June 1990 due to the depletion of its ore reserves. During the course of its life, it produced 109 million tonnes of waste rock which is thought to contain approximately 0.10% copper as chalcopyrite and 0.015% molybdenum as molybdenite stored in four rock piles around the pit. Similarly, approximately 200 million tonnes of tailings were produced grading 0.015% copper and 0.005% molybdenum. During operation, tailings were hydrocycloned to produce a coarse sand for dam building with the fine slime stored behind the dam. It is estimated that 40% of the tailings were used for the dam construction. The layout of the mine site is shown in Figure 2.The mine was operated under a zero discharge permit. This was achieved by diverting practically all surface water flowing towards the mine area around the site. Water affected by mine operations, was discharged to the tailings impoundment, and recycled to the mill for reuse. The recycle rate was 90% and seepage from the tailings embankment and saddle dam was pumped back to the tailings pond. Because acid drainage was not a problem with the low grade Brenda rock, runoff from the rockpiles was initially released to the local environment. However, due to the alkaline nature of the waste rock, dissolution of low concentrations of molybdenum in the site runoff waters became evident by the late 1970's. The molybdenum content of Peachland Lake, the water reservoir for the municipality of Peachland, approached the Canadian Drinking Water Standard of 0.25 ppm. By 1979, a series of diversion and collection ditches was completed to direct all runoff water from the rockpiles to the tailings impoundment area. This had a significant detrimental effect on the site water balance causing it to go into a considerable surplus condition. During normal operation, this oversupply was contained by the rate of tailings dam building. An economic shutdown of the mine during 1984 forced a limited discharge from the site to protect the integrity of the tailings dam."

Jan 1, 1995

By J. Andrew Cornier, Mike Cooper

"Upon successful completion of the AG/SAG Mill project (Cooper et al., 2001) the Brunswick Mine concentrator was configured with single line processing through the AG/SAG mill, followed by two parallel ball mill and flotation circuits. The Stage Grinding project was undertaken to utilize a pre-flotation stage on the AG/SAG mill primary cyclone overflow and simplify the flowsheet by consolidating the two ball mill and flotation circuits, resulting in single line processing fi-om AG/SAG mill to final concentrate. Concurrently, the reclaim water supply capacity was increased to enable flotation cleaning at lower percent solids.Benefits from the Stage Grinding project include revenue enhancement from increased zinc grade and recovery combined with additional lead recovery. Further benefits included a reduction in the number of process pumps, reduction in loss associated with spillage to sumps and process simplification through efficient use of equipment layout.The Stage Grinding project was implemented in six phases over a period of seven months in 2000. Each phase was commissioned during scheduled concentrator shopdays with minimal additional downtime required. One year after project approval the Stage Grinding project was completed at a total cost of $1 1.8 million.INTRODUCTIONThe Noranda Inc. Brunswick Mine, located near Bathurst, New Brunswick, is one of the world’s largest leadzinc mines. Milling of ore began in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore fkom the No. 6 open pit located several kilometers away. In 1980 the pit was exhausted and the concentrator rationalized to accept a design feed rate of 10,500 tomes per day from the No. 12 mine. In excess of 100 million tomes have been milled to date, representing over three-quarters of the known reserve. The concentrator employs differential froth flotation to produce lead, zinc, copper and leauzinc bulk concentrates. The ore is massive sulphide, with approximately 20% valuable minerals, 60% pyrite and 20% silicates."

Jan 1, 2002

By A. T. Robles, L. Buckingham

"When certain gold ores containing pyrite, graphite, humates and fumates or other organic compounds which are possibly adsorbed on clay minerals are treated with aqueous solution of alkali metal cyanide, the aurocyanide complex which forms is adsorbed by these ""pregrobbing"" species in the ore. Dissolution of gold proceeds until chemical equilibrium is attained between the cyanide complex adsorbed by the ore and in solution.Sodium lauryl sulfate (SLS) may be used to decrease the adsorptive activity of the prerobbing species in the ore. It has successfully improved recovery at Holt-McDermott Mine and is highly biodegradable.INTRODUCTION Precious metal-bearing ores containing preg-robbing species are poorly amenable to standard cyanidation techniques. The ""preg-robbing"" impurities with which they are associated tend to adsorb the gold cyanide complexes and carry them through the circuit to the tailings. Also, gold may be associated with these impurities in the ore and not be leached by the cyanide complexing agents used in conventional cyanidation techniques.Methods for dealing with the problem of preg-robbinq impurities in gold ores are:Carbon-in-Leach (CIL), in which granular activated carbon is added to compete with preg-robbing species for the gold cyanide complexes. This process can improve gold recovery in mildly preg-robbing ores especially when fine carbons with very high activities are used. However, with severely graphitic ore, the carbon associated with the rock may outperform the granular activated charcoal.Chemical Oxidation, in which the ore is subjected to preliminary chemical oxidation techniques to oxidize carbonaceous materials using such diverse reagents including ozone, sodium hypochlorite, calcium hypochlorite, chlorine gas, permanganates, perchlorates, chlorates and oxygen prior to subjecting the oxidized ore to standard cyanidation treatment to extract gold. This is generally a costly method and will not prove economical on a low-grade ore."

Jan 1, 1992

By Desmond M. Levins, Reginald K. Ryan

"Removal of radium from uranium tailings would substantially lessen their environmental impact because it would eliminate, or greatly reduce, the hazards arising from radon emanation, leaching of radium, dust dispersal and gamma radiation. A process is described to extract radium from tailings by leaching in 3M Nael solution. The radium can be rapidly leached at ambient temperatures by either multi-stage contact in agitated vessels or washing on a tailings filter cake. Extraction efficiencies ranging from 80 to 94% have been measured for tailings derived from a number of Australian ores.Methods for radium recovery from the leach solution by precipitation, ion exchange and adsorption have been examined. Precipitation as barium/radium sulphate appears to be the most suitable approach. A possible flow diagram for the process is proposed which involves recycle of the saline leach liquor.IntroductionDisposal of mill tailings is the most contentious environmental issue facing the uranium mining industry. The conventional approach to managing tailings is to store them in a dam above ground. Attempts have been made to stabilize some tailings piles by physical or chemical methods. In a few cases, the tailings have been covered with earth and revegetated. However, most tailings piles from abandoned mills are in a neglected state and subject to the erosive and dispersive forces of nature. The cost of remedial action is high. The United States Department of Energy (DOE) has estimated that it will cost up to $130 million to rehabilitate abandoned mill sites in the U.S .A..Environmental authorities are now insisting that uranium mining companies improve their methods of tailings disposal. In the United States, the preferred approach is burial, either in mines or in specially excavated pits. Tailings retention systems as operated in Canada are regarded as storage facilities until an acceptable method of disposal is found. In Australia, the Fox Commission recommended that tailings from the Ranger project should be returned to the open-cut mine. Returning tailings to the mine has considerable merit, but it is not always a viable option. In many underground operations, it is not physically possible to fit all the tailings in the worked-out mine. In other cases, the groundwater hydrology of the mine may be unfavourable."

Jan 1, 1980

By Axel Doelcken

?? Liebherr excavators in standard configuration are generally designed to operate in an ambient temperature range between -12°C and +55°C (10.4 °F and 131°F). ?? To operate effectively and efficiently in very low temperatures special consideration needs to be given to the design ??An inappropriately designed hydraulic system unable to withstand extreme changes in temperatures or varying working conditions will lead to: - premature wear and more frequent replacement of components - increased maintenance & unplanned downtime This will affect productivity & operating costs.

May 1, 2008

By S. Kelebek, E. Yalcin

"Diethylenetriamine (DETA) and/or triethylenetetramine (TETA) are reagents that can generally be categorized under ""flotation modifiers"". These polyamines are known to be effective for the deactivation/depression of pyrrhotite that is present in Canadian nickel-copper sulphide ores in large quantities. Recent bench scale testing on some of these complex ores indicated that chalcopyrite flotation can also be adversely affected under certain conditions involving polyamines, which may be taken to be an evidence of their specific affinity towards chalcopyrite. In general, DETA shows no significant depressant action on chalcopyrite while TETA's promotion of hydrophilicity for this mineral has been observed even in the presence of xanthate. This prompted an investigation on thermodynamic equilibria involving these polyamines and copper ions. Results serve a useful purpose to identify conditions for deactivation phenomena. CuDETA2+ and CuTETA2+ chelates as products of sequestration have large predominance areas covering a wide range of redox and pH conditions, where copper xanthate species are not stable, especially in the presence of TETA."

Jan 1, 2009

By Lam. E. K., J. W. Jarvi, B. P. Bharadwaj

"The Rabbit Lake Operation is located in Northern Saskatchewan on the western shore of Wollaston Lake approximately 800 km north of Saskatoon, Saskatchewan.The Operation is owned by a joint venture of Cameco Corporation and Uranerz Exploration and Mining Limited. Cameco is the operator.The Mill started production in June 1975 under the ownership of Gulf Minerals Canada Limited and over the past twenty years has increased the rated production capacity to over twice the original design.The Mill was originally designed to process 485,000 tonnes per year of ore with an annual production of 2.0 million kg of U308 (yellowcake). During the past ten years changes in ore sources and process modifications have led to increased milling capacity and production targets. The Mill has treated a total of 8.6 million tonnes of ore and produced over 40 million kg of U308 in the twenty year period, despite the shutdown and cutbacks in operating time since 1989.The 1995 tonnage throughput was 206,000 tonnes per year. Annual production though has risen to 3.6 million kg of yellowcake despite the departure from full-time to half-time operation in 1991 (7 up-7 down). The forecast for the coming years will see tonnage rising from 300,000 to 570,000 tonnes, production of 4.5 to 5.4 million kg of yellowcake and an anticipated return in the year 1998 to full time operation.This paper will give a brief history of the Mill and focus on the present flowsheet and recent improvements to the process."

Jan 1, 1996

By J. D. Bliss, G. J. Orris, W. O. Menzie

"Analysis of gold placer data from gold placers throughout the world suggests that gold grades and volumes cannot be used to distinguish between most types of gold placers. Only the alluvial plain and fan placers are significantly different among the types of gold placers considered. Gold grades and volumes change when working gold placers go from small-volume methods to large-volume methods. The small-volume methods are either of the surface or subsurface (drift mining) type. The odds that a gold placer will be dominantly worked using small volume methods at the surface are about 5:3. Once small volume mining has occurred, the odds against subsequent large-volume mining are about 4:1. If a placer deposit is suitable for large-volume mining, and the amount of gold produced from small-volume mining was reported, an estimate of the remaining gold (log10kg) can be made using the following equation:The relationship shown by this equation suggests that very large gold placers have over half their total gold content in the rich gravels appropriate for small-volume mining. These relationships in conjunction with other data and computer simulation can be used to estimate placer gold resources for an area targeted for mineral resource assessment. IntroductionGold placers have been an important source of gold throughout man's history and may have been one of the first sources-dating back to around 6000 B.C. (Gregory, 1980). From the primitive basket to the modern floating dredge, a variety of methods have been used to work gold placers. Placers are found in a wide spectrum of geologic and climatic settings. They have been classified by using a mixed set of criteria including age (Recent, Tertiary), climate (desert), fluvial position (modern channel, bench, alluvial fan), environment of deposition (eluvial, alluvial, residual, beach), type of transport (stream, wind, glacier), and whether the placer is buried or at the surface (Boyle, 1979; Wells, 1969). Several mining methods have usually been used to extract gold from anyone placer."

Jan 1, 1987

By E. Roman, M. Brissette

"For its end of life, Kemess Mine processed its difficult-to-treat hypogene ore from the east pit, which is a lower grade copper and higher pyrite content gold ore. Metallurgical testwork showed that improvement was possible with finer grinding. In July 2010, small grinding media were introduced into the regrind mill resulting to a nearly 50% decrease of the power draw with a finer grind, an increase of 1.1% in the final copper grade, an increase of 1.2% in copper and 2.9% in gold recoveries at the cleaners. The overall recoveries of copper and gold increased by 0.6% and 1.3% respectively.INTRODUCTION Kemess Mine was located at 430 km North West of Prince George in North Central British Colombia. Its production started in May 1998 by Royal Oak Mines and was acquired by Northgate Exploration in 2000. Kemess reached its end of life production with its last of production on March 9th, 2011. The concentrator nominal capacity was 56,000 TPD consisting of a primary gyratory crusher feeding two identical conventional SAB circuits totalizing 35,840 kW available for grinding. Rougher flotation consists of four parallel rows of eight 130 m³ tank cells to produce a concentrate of 2% Cu and 6 gr/t Au which was pumped to a 5 m x 10.8 m regrind ball mill. The regrind installed power was 4,480 kW to generate a final grind of 80% passing 40 microns. This product fed the two stage cleaner flotation circuit, first stage being two parallel rows conventional 17 m³ Denver cells and second stage, two 3.3 m x 11.0 m flotation columns in series, designed and constructed by Kemess, to produce a 22% copper concentrate grade. The cleaner flotation middlings were returned back to the regrind mill (Froehling, Mohns, Roman & Grady, 2005; Froehling, Mohns, Rajwani & Fullam, 2007). The large copper-gold porphyry deposit was composed of 15% of oxidized supergene and leach cap ore and 85% of hypogene ore having three major alteration zones, mainly characterized by their sulphur/sulphide content. Chalcopyrite is the main copper mineral species with minor bornite. Gold was mainly present as free gold or electrum. The first alteration zone consisted of a phyllic alteration characterized by moderate and intense sericite alteration which had the highest mill throughput and twice the mine grade when in contact with the supergene horizon. The second zone consisted of a potassic alteration characterized by more competent groundmass and moderate to intense quartz stockwork which had a slightly lower mill throughput. Finally, a third alteration found beneath and on the east side of the deposit having almost no quartz stockwork and hardened groundmass, sulphide content is two to four times more. This zone had the lower mill throughput with significant drop in the mine average gold and copper grade. Through its mine life, the average life of mine copper head grade was 0.21% Cu and 0.64 gr/t Au. Head grades began to decrease from 2008 to be substantially lower in 2011 as shown in Table 1."

Jan 1, 2012

By Xiang-dong and Zhu, Xia-jie Hao, Yi-lin Fan, Quan Jiang, Lian-xing Hu, Xia-Ting Feng, Shao-jun Li

"Columnar jointed basalt (CJB), which typically originates from lava flows, is characterized by a special geological structure. The joint network inside the CJB is self-organized into a rough hexagonal arrangement. The field test and measurement during the unloading excavation showed that the columnar joint sets and other internal joint sets not only are important for determining the mechanical properties of the CJB, but also induced its asymmetric unloading deformation and failure. A constitutive model, which considered mechanical deterioration of both the joint sets and rock blocks, was also developed to simulate the unloading behaviour of the CJB. The crack-confinement supporting method for surrounding CJB was developed to overcome the failure of columnar jointed rock mass in the diversion tunnel. Application to the CJB segments of the diversion tunnels in the Baihetan project indicated that this supporting method was reasonable and efficient. INTRODUCTIONColumnar jointed basalt (CJB) rock mass, which typically originates from lava flow, is characterized by a special joint structure. The joint network of the CJB is self-organized into a roughly hexagonal arrangement and leaves behind an ordered colonnade. The widely accepted current view regarding the formation of columnar jointing involves a thermally induced contraction–cooling mechanism (Peck and Minakami, 1968; Degraff and Aydin, 1987; Budkewitsch and Robin, 1994; Lore et al., 2000; Goehring et al., 2006). Although the columnar basalt formation mechanism is fairly understood, the geotechnical characteristics of such rock mass are not well established. In practice, joint sets always induce the multi- failure modes of surrounding rock in the tunneling space (Ibarra et al., 1996; Waltham and Swift, 2004; Mahendra and Bhawani, 2008).According to conventional wisdom in rock mechanics, structural joints affect the mechanical response of rock mass by introducing strong anisotropic effects (Barton and Bandis, 1980; Hudson and Priest, 1983; Brady and Brown, 1985). The columnar jointed rock mass should be considered an anisotropic material given that the joint sets are distributed with some principal directions and characteristic spacing distributions (Cundall and Fairhurst, 1987; Hoek et al., 1998; Wu and Wang, 2001; Budetta and Nappi, 2011; Fortsakis et al.,2012). Columnar jointed rock mass would result in structurally controlled failure patterns, even serious disasters during underground excavation because of the low shear and tensile strengths of intensely joints (Gencer,1985 ;Ishida and Uchita, 2000; Jia and Tang, 2008). The CJB rock mass in the Baihetan hydropower station has exhibited serious stability issues, including sliding, fracturing and even collapse (Shi et al.,2008). Therefore, the geotechnical and geomechanical characteristics of such rock mass should be assessed through in situ investigation and numerical analysis to minimize future accidents during underground unloading excavation."

Jan 1, 2015

By Derek J. Cyr

Climate has an immense impact on the properties of Athabasca oil sands and the ease or difficulty with which it is excavated and extracted. During the winter months, exposed oil sands surfaces freeze. When hydraulic and electric shovels used in the various operations in the Fort McMurray area excavate the benches, large frozen lumps are created. These lumps are either destined for a lump dump where increased costs are incurred due to rehandle, or they are sent to the primary crushers where they can cause significant downtimes by jamming the crusher. The crusher down time can sometimes be in the order of hours of lost production. Studies on data from two consecutive winters at Syncrude Canada?s North Mine show a high correlation between climate and the generation of frozen oil sands lumps. Temperature, ground cover, traffic, bench exposure time, oil sands grade and moisture content all contribute to this problem. A 1D frost penetration model was used to predict the depth of frost expected in oil sands and the corresponding likelihood of frozen oil sands lumps. To combat the challenge of frozen lump generation, a number of proactive and reactive measures can be taken: artificial snow, shallow ponds, scheduling, blasting or ripping.

May 1, 2004

By Qidong Xu

Abstract - The metavolcanic rocks of the Dahongshan area have been subjected to both widespread hydrothermal alteration from circulating seawater and to regional metamorphism, such that their mineralogic and chemical compositions have been changed significantly. Studies of immobile elements show that these rocks were originally tholeiitic basalts with geochemical characteristics of T-type MORB formed in a mid-ocean ridge environment. The accompanying Cu-Fe-rich VMS mineralization is in keeping with the original petrology and tectonic setting of these metavolcanic rocks. Regional alteration evident in K-Mg enrichment is related to sericitization and chloritization. Local alteration associated with Cu-Fe mineralization appears to have been accompanied by Si enrichment and Na depletion. P enrichment identifies metasomatism related to Cu-Fe reconcentration in the mineralized zone.

Jan 1, 1999

By D. Bois, P. Poirier, B. Bussière

Environmental desulphurization is an attractive alternative for the management of acid generating tailings. Bulk sulphide flotation is the most commonly used method to produce desulphurized tailings, as shown from many previous works. This process can reduce the volume of problematic tailings to manage by making a sulphide fraction concentrate. The desulphurized tailings can then be used as construction material for an engineered cover to prevent acid mine drainage. In this paper, the results from different laboratory and pilot scale flotation tests are presented. These results confirm that low sulphide non-acid generating tailings can be produced by froth flotation. It is also demonstrated that sulphide concentrates obtained from desulphurization mixed with the appropriate binders can be used as feed material for paste backfill. From previous work on the evaluation of operating and capital costs of the desulphurization process for different ore types, the use of desulphurization as a tailings management approach was studied for several scenarios. The desulphurization process was then compared with other technologies (underwater disposal, engineered soil cover) currently used to prevent acidic drainage (AD) for a new mining operation. Results show that partial or total desulphurization are technically and economically viable options to manage tailings during mine operation. Mine waste managers should consider desulphurization as an efficient technology.

Jan 1, 2005

By J. Patterson

"Acid Mine Drainage (A.M.D.) formed from oxidation of disposed pyritic wastes presents a potential pollution hazard for Equity Silver Mines Limited and the mining industry in general. Failure to collect and treat these acidic effluents can eventually affect the quality of receiving streams and endanger aquatic life forms. Treatment and discharge of A.M.D. to the receiving environment must comply with stringent anti-pollution laws.Wastes mined at Equity Silver Mines Ltd. are predominantly acid generating and must be handled according to a materials management plan to minimize impact on the environment. Effluent collection and treatment facilities have been constructed to address the short term environmental concerns with reclamation and research programs being developed to address and monitor abatement measures for control over the long term.The purpose of this paper will be to offer a brief overview of these measures. Key topics for discussion will be: kinetics of Acid Mine Drainage, acid generating potential, water treatment and sludge handling, special wastes, waste dump construction, waste rock amendments, bacteriocide research, simulations, and reclamation measures. Programs designed to evaluate impact on the receiving environment, although they form an extensive part of the program, will not be discussed in this paper.INTRODUCTIONEquity Silver Mines Limited is located in the Central Interior of British Columbia, within the Omineca Division, 35 kilometers southeast of Houston Mining and approximately 575 kilometers by air north-northeast ofVancouver (Figure 1).The deposit consists of three economic mineralized zones: The Southern Tail, Main and Waterline Zones. Mining of the Southern Tail commenced in April, 1980 and was completed in the first quarter of 1984. Main Zone production was initiated during the last quarter of 1983 and the Waterline Zone remains undeveloped to-date. Remaining reserves are in the order of four years contingent on favourable metal prices. The three deposits (Figure 2) occur within a window of interbedded volcanic and minor sedimentary rocks. The older volcanic sequence is bounded on the west by a small granite-like stock and on the east by a gabbro-monzonite complex. All are overlain by a series of younger lava flows."

Jan 1, 1988

By V. Casaletto, L. Widdifield

"Equipment capable of higher production requires more sophisticated maintenance programs. Reduction in fleet size through use of modern technology requires higher utilization and greater availability.Caterpillar's maintenance control system program will display the capabilities of computer technology in a maintenance environment. Utilizing such resources will provide increased availability, lower inventory requirements, fewer last minute disasters, and lower cost per ton for maintenance. High production from premium quality equipment can only be realized by good planning and care, executed by maintenance and operating personnel acting cooperatively.IntroductionAccurate information is the key to accurate decision making. In today' s competitive industry, mines' teeter on the brink between profit and loss. Their survival will be predicated on their ability to retrieve, analyze, and formulate accurate information regarding the operating costs of mobile equipment. This paper deals with the impact of managing equipment operating costs and introduces a computerized control system designed for monitoring and controlling these costs.Caterpillar has produced several computerized modelling programs designed at taking the guesswork out of equipment selection and management. The most basic of these is Equipment Investment Analysis (E.LA.). This model compares two or more alternatives over the projected life of the machine. It deals with comparisons in the owning and operating costs of equipment, as well as, their productivity over a specified period of time.In order for a computer model to be credible as a decision making tool, it must be simple to use and the results measurable and accountable . Caterpillar has introduced several computer models to assist the industry in selecting and managing an equipment fleet. These programs allow the user to predict and monitor the operating costs and production capability of mobile equipment. They are often used in the machine selection process to select equipment which best suits a mine's long-term cost per tonne objectives. After the optimum selection is made, Caterpillar's Maintenance Control System provides an efficient management tool to monitor and verify the data used in the selection model."

Jan 1, 1992

By Christoph Broeckmann, Andrew Gardula

"The adoption of high-pressure grinding throughout the cement industry has yielded enormous savings in energy and wear material costs. There are also a number of installations, which are applied in iron ore and kimberlite processing operations. High pressure grinding offers the same potential for energy and media savings, as well as opportunities for improved, downstream metallurgical performance, in hard rock mineral processing. This paper focuses on developments in machine design and wear surface engineering for high-pressure grinding of the hard rock ores in the mineral processing industry.INTRODUCTIONIn the early eighties 3.4 % of the world electric energy was consumed for crushing and grinding operations of minerals and 0.6 % in the ore dressing industry. During that time using a laboratory press Prof. Schönert found that interparticle comminution at high pressure requires remarkably less energy as conventional grinding technologies. In grinding of cement clinker he expected energy savings between 33 and 50 % compared to ball mills [1]. Piston presses are limited in throughput. Thus, the principle of interparticle comminution was adapted by the use of high pressure grinding rolls (HPGR) to a continuous plant process. Figure 1 shows the main components of a roller press: The roller nip of width s is formed between the floating roller (1) and the fixed roller (2). The diameter of the rolls will be denoted with D, the width with B. The total press force F can be adjusted using a hydraulic system with two or four pistons (3). In order to compare machines with different roll geometries the specific pressing force p is introduced:"

Jan 1, 2005

By R. Abel

During the last eight years, the Ventanas electrolytic refinery has been improving its results by applying a series of new operational practices, such as the reduction of electrode spacing from 115 mm to 105 mm., raising the current density to 305 A/m2, modification in the operators shifts to about 70% of the workers, reducing the pulling cycles by two hours, and the erection of a new anode preparation machine. These improvements allowed an increase in the refinery capacity from 323,000 to 382,000 metric tons of electrolytic copper per year. In the same period it has obtained certification under the ISO 9001:2000 and ISO 14001:2004 standards. All of these developments have been achieved using conventional technology with copper starting sheets. The continuous improvement that is covered in this paper has allowed Ventanas to optimize its production, costs, and operational results. For the next years, it is planned to follow the same tendency, bringing the productivity of the plant to higher levels.

Jan 1, 2007

By A. Gillis

"In order to obtain the equipment necessary to automate the equipment it was decided that the Department of Mines and Milltronics would jointly survey the circuit. This survey involved a metallurgical study by the Department of Mines and Milltronics took sound recordings during manual operation, checking in circuit bottle necks and determining the number of sensors that could be applied.The conveyer is driven by 2 PIV drives, each has a variable speed of 3/ 1. The cross conveyer that feeds directly into the rod mill feed chute will contain an automatic sampler and will eliminate the necessity of stopping the belt feeders to take samples. At present, hourly samples are taken from this point and the circuit is upset and cycling results.The Adequate scale is very close to the rod mill to totalize and thereby give a continuous indication and the second purpose is to take a signal from it and control the solution addition to the rod mill.A pneumatic positioner was obtained and put on the scale which produces a 3 tp 15 psi signal which is proportional to the actual tonnage going across the scale. This signal was fed to a bias loader which was mounted on the control panel. From there the operator can set the amount of solution that is wanted by the rod mill. No matter what the feed rate is, the signal coming from the bias loader will follow it automatically. This in turn goes to the pneumatic solution valve (1 ½” ) and the solution goes in with the feed to the rod mill. This is only part of the solution required for dilution since the rest of it is used as a force jet for the feed to the rod mill."

Jan 1, 1965

"This paper presents results of ongoing research being carried out at the University of New South Wales into improved methods for estimating rock mass strength. In this paper the rock mass strength of Hawkesbury Sandstone is assessed. An estimate is first made based on an extensive database of intact rock strength data being collated by the author and the assumption that the rock mass strength can be determined following the GSI approach. The resulting strength envelope is then compared with nine documented case studies where stress induced failure has occurred in tunnels excavated within Hawkesbury Sandstone. The case studies cover a range of conditions:• construction dates from 1880s to 2006• spans from 3.2 to 12.5 m• depths from 17 to 124 m• excavation methods of drill & blast, roadheader and TBM.Finite element modelling is used to assess the likely magnitudes of the stresses present at the time of failure. The finite element models are interrogated to provide the values of major and minor principal stresses in and around the mapped failure zone. Failure and stable points are then compared to the rock mass strength envelope(s) of Hawkesbury Sandstone derived using the Hoek-Brown criterion and spalling criteria. The Hoek-Brown rock mass strength envelope differentiates between the ‘failed’ and ‘stable’ zones remarkably well.INTRODUCTIONStress induced failures of rock masses are fortunately rare in civil tunnelling. However, when they occur they provide an opportunity to calibrate an assumed rock mass strength criterion. This paper compares an estimate of rock mass strength for the Hawkesbury Sandstone with the following cases where stress induced failure occurred in tunnels excavated within Hawkesbury Sandstone.• Cataract & Devines Tunnels of the Upper Canal (Upper Nepean Scheme)• Malabar Outfall Tunnel• Sydney liquefied petroleum gas (LPG) Storage Cavern• Northside Storage Tunnel• Cross City Tunnel• Lane Cove TunnelTwo-dimensional finite element modelling is used to assess the likely magnitudes of the stresses present at the time of the failures. The principal stresses are then compared to the assessed rock mass strength envelope(s) of Hawkesbury Sandstone derived using the 2002 Hoek-Brown criterion (Hoek et al., 2002) and the spalling criterion suggested by Diederichs (2007)."

Jan 1, 2015