Search Documents

By Q. J. Xu

In blasting sensitive areas, the use of traditional blasting for rock fracture is restricted. Soundless cracking technology has been presented to achieve rock fracture instead of blasting. However, when the diameter of a borehole is larger than 50 mm, it is very dangerous that the soundless cracking agent (SCA) will have a tendency to gush from the borehole, which also leads to cracking failure. In this paper, a new large scale borehole soundless cracking technology is presented, using a new enlarged drill bit and mechanical plugging, which could solve the hole gush problem. Experiments were made to measure the expansive pressure of the SCA working in boreholes of F40 mm and F100 mm. The difference between two sizes of cracking phenomenon was compared and analyzed to show that large scale borehole soundless cracking has larger expansion pressure and faster reaction speed than traditional soundless cracking. Based on the experimental data, finite element analysis was utilized to study the stress concentration and stress distribution between two neighboring large boreholes in several media during the cracking process. Using the data from experiments with finite element numerical simulation, the brittle material fracture condition was analyzed to design the parameters of the borehole mesh. With the theoretical analysis, the technological process and reference parameters of large scale borehole soundless cracking were proposed, including the borehole spacing, burden distance, depth of the borehole and so on. Then a single rock was cracked with the new enlarged drill bit and new plugging, and the results proved the feasibility of the new large scale borehole soundless cracking technology, which achieved the expected goal of improving the cracking ability and reducing the cracking time. This paper presents a new efficient rock cracking method for blasting sensitive areas, and which could apply to mining engineering, demolition engineering, excavation engineering and other engineering applications where there is a demand soundless and non-vibration fracturing of rock.

Aug 1, 2013

By Q. Li, X. -L. Lei

Accompanying global energy shortage and high energy prices, shale gas has been the focus of exploration and it has attracted great attention from both academia and industry. The combination of horizontal drilling and hydraulic fracturing enables the extraction of huge quantities of shale gas from impermeable shale formations. During the process of hydraulic fracturing, great amounts of water, sand and chemicals must be injected into the shale formation at a high pressure, which may trigger seismicity. Such events could occur due to the presence of faults or the reactivation of these potential faults. Shale is the main component of source rock in the Sichuan Basin, China, and has the greatest contribution to the natural gas of the basin. In order to investigate the fracturing mechanism of shale in the Sichuan Basin, China, a series of rock fracture tests under triaxial compression were carried out in the laboratory in order to investigate the hydraulic fracturing behavior in relation to the seismicity. Based on the high multichannel waveform recording system, the detailed spatio-temporal distribution of acoustic emission due to micro-cracking was used to examine the deformation and fracture behavior of shale. In addition, the effects of confining pressure on the deformation and fracture in shale under triaxial compression by acoustic emission monitoring and X-ray computed tomography were discussed in the paper. The experimental results indicate that the shale samples demonstrate brittle fracture behavior. The confining pressure has a great effect on the complexity of the fracture planes and the spatial distribution of acoustic emissions, but few contributions to the location of the major hydraulic fracturing plane. The foliation of shale plays a dominant role in the rupture process and the geometry of the major hydraulic fracturing plane. The high confining pressure increases the strength and restrains the deformation of shale during compression process. These results shed some light on the understanding of onset of hydraulic fracture initiation not only for shale gas exploration and production, but also for CO2 geological storage and underground energy storage.

Jan 1, 2015

By S. E. Malouf, W. Thorpe

"The Chibougamau Explorers property comprises 35 claims or 1,726 acres in La Dauversiere and Rohault Townships along the west shore of Norhart Lake. The property is serviced by a twelve-mile all-weather high-way leaving the main St. Felicien-Chibougamau highway at mileage 121. This road junction will be a railway siding on the C.N.R. St. Felicien-Chibougamau branch line in early 1957. The property lies 30 miles due south from the town of Chibougamau. The mine-sponsored Chibougamau Explorers town has a population of 250 with 28 residences and a well equipped recreation centre.The original showing was found by Rider Storn, a prospector financed by Val d'Or interests during the 1950 field season. This showing did not stand up to detailed prospecting but it did lead to the discovery of the main zone by A. Oakley during the following winter.The company was incorporated in I 950 under the presidency of E. Ayrhart. Financing following the initial drilling involved an increase of capitalization from 3 to 5 million shares and the advent of financing and management control by Anacon Lead Mines Limited.Exploration and development to Jan. I, 1956, has included 42,331 feet of surface diamond drilling; 31,363 feet of underground drilling; a 3-compartment shaft to 1,245 feet, with 8 levels at 150-foot intervals; and I 1,054 feet of drifts, crosscuts, and raises. Ore reserves to the 900-foot level, proven and partly developed, total 546,725 tons averaging 0·30 ounces gold and 0·93 per cent copper. A mill with capacity of 500 to 750 tons will be installed by January 1956.GeologyThe Chibougamau Explorers property lies along the belt of Keewatin volcanics striking through Bachelor Lake, Opawica Lake, Doda Lake, Coapatina, Norhart Lake, and Rohault Lake, where it is in close proximity to the paragneisses of the Grenville series. The boundary of the paragneisses, a gradational contact, lies within a mile of the Chibougamau Explorers shaft collar."

Jan 1, 1949

By H. S. Bostock

"The study of the geology of Mayo district began in 1900 with R. G. McConnell's exploration of Stewart river; and has been carried on intermittently for the Geological Survey of Canada by J. Keele; D. D. Cairnes; W. E. Cockfield; C. H. Stockwell; and the writer. It has revealed a thick group of schists; quartzites; and crystalline limestones; almost certainly Precambrian in age; placed in the Yukon group; and a group of younger intrusive igneous rocks varying in composition from granite to gabbro. These include bodies of many sizes and forms; from stocks 8 miles across to small dykes and sills.Prospecting in the district began about 1897 and has disclosed the presence of gold, silver, lead, tungsten, tin, zinc, antimony, copper, arsenic, manganese, and mercury, both in placer and lode deposits . Of these metals gold, silver, and tungsten have been produced from placer deposits, and silver, lead, and gold from lodes. This paper deals only with some structural relationships of the lode deposits, generally referred to as the silver-lead veins, which have yielded the greater part of the mineral production of the area. From 1938 to 1941 the geological mapping of the district for publication on a scale of 4 miles to 1 inch was carried out by the writer, and the ideas expressed here are based on this work. No comprehensive detailed study of the silver-lead veins has been made, but the mapping has brought to light several important facts. The rocks of the Yukon group, although of a striking sameness throughout their section, can be divided into a number of distinguishable members by means of which the main structures can be traced . The most important and readily distinguishable rock in the group is referred to here as the quartzite member. This is composed of quartzite and schist in about equal proportions. The quartzite is resistant to erosion and forms most of the outcrops of the'member and the peaks of the Gustavus mountains and mount Haldane (Lookout)."

Jan 1, 1954

By A. L. MacAllister

"Keymet Mine lies on the Elmtree River 4 miles west of Petit Rocher and about 18 miles northwest of Bathurst, Gloucester County, New Brunswick. A good gravel road runs through the property.The deposit was discovered in 1880 when a vein 7 to 8 feet wide was re-ported. In 1883 the first shaft was sunk as part of an exploration program which was carried out sporadically until 1897, when the company dissolved. Activity on the property was intermittent until 1946 when the property was again staked and during the next year diamond drilling intersected the vein in several places. In December 1949 Coulee Lead and Zinc Mining Company optioned the property and held it until its sale to Keymet Mines Limited in June 1952. By July of the same year work had begun on a three-compartment shaft.Early development work was concentrated on the 150, 300 and 450 foot levels but the shaft has since been sunk to I,228 feet to allow for eight levels at intervals of 150 feet. Little lateral work has been done to date below the 600-foot level.The first mill constructed to treat the ore was to go into production in the spring of 1954, but was destroyed by fire. A second mill, having a capacity of 200 to 300 tons a day, was officially opened October 25th, 1954. Present shipments are zinc and copper-lead-silver concentrates.The rocks of the area are shale, conglomerate, limestone, and quartzite with minor interbedded volcanic rocks, all probably of Silurian age. Small plugs of Devonian acidic intrusions occur 3 miles east and about 5 miles northwest of the mine. Minor dykes of both acidic and basic rocks occur on the mine property. In many of the fine-grained sedimentary rocks the more brittle laminre have been brecciated and give the rock the appearance of an argillaceous conglomerate."

Jan 1, 1949

By D. Monder

Fuel cells are promising sources of electrical energy because of their exceptionally high efficiencies and low environmental emissions. Solid oxide fuel cells (SOFC) are high temperature, ceramic electrolyte cells, which unlike most other fuel cells, can work with a variety of fuels besides pure hydrogen, including hydrocarbons and hydrogen sulfide. In this presentation I present the models we have built for a single H2S fuelled SOFC. The goal is to model the electrochemical reaction coupled with mass transfer, fluid flow and current/voltage distribution in an yttria stabilized zirconia electrolyte fuel cell assembly operated between 750 and 850 °C. Two models of the above SOFC are discussed in this work, 1) an algebraic model derived by integrating a one-dimensional model along the thickness of the cell, and 2) a two-dimensional finite element axi-symmetric model. Important physical and kinetic parameters in the 1-D model are fitted using experimental data sets at three different temperatures and reasonable fits are obtained for the data sets at 800 °C and 850 °C. However, the third data set at 750 °C does not give a good fit using this model. We need to apply the curve-fitting to more data sets to investigate it further. The main conclusion that can be drawn from the present study is that the ohmic resistance of the cell is very high. Contact resistances are believed to be the reason behind this although there could be contributions from resistive interphases. The 2-D model also gives reasonable fits to data sets. One of the publications by the laboratory group doing the experimental work in our department presented very interesting data that shows the cell performance being a strong function of fuel flow-rate. We want to model this behaviour and will have to depend on the 2-D model as the 1-D model does not have flow-rates as parameters.

Jan 1, 2005

By A. C. Bonnier

"INTRODUCTIONI have been asked to make available our method of determining the ""Live Inventory"" within a gravity type thickener. The ""Live Inventory"" being that tonnage compatible with the optimum underflow density that satisfies the process requirements.One prerequisite of the calculations is the proper determination of the final pumpable underflow density. A further requirement is that, prior to shutdown, the thickener has been operated properly, particularly that it has not been operated for storage capacity.Until recently I had few chances to check the method. However, I was most fortunate to have had a recent opportunity covering several months on a 300 ft diameter thickener at a plant in the U.S.A., wherein the method was tested and verified many times. The method proved accurate in that our calculated tonnage checked with the actual dried physical inventory of the drying facility.At the precise moment that the feed to the thickener was shut down, a per cent solid determination was made at the critical place and the tonnage in ""Live Inventory"" determined. Let us say that at the moment the thickener U' Flow solids was 20%. The thickener then was allowed to cperate until the U' Flow solids dropped below 20%, at which time the thickener was deemed to be relieved. Anything much less than 20% was not suitable to the subsequent centrifuge operation.Most large metallurgical thickeners have a configuration of three slopes at the thickener ""bed"". These slopes are dictated by the raking blades and are shown on the manufacturers' erection drawings. The first slope is from the outer periphery and generally takes up a distance of about 60% of the radius, the second slope generally taking up the remaining 40%. Within this 40% area, there is usually another slope taken up by the cone on ""B"" type thickeners or an actual steeper inner slope on ""C"" type thickeners."

Jan 1, 1989

By C. Barnett

"The mill at Noranda Mines Limited, Geco Division in Manitouwadge, Ontario went into production in 1957 treating 3,500 tons per day of copper, zinc, silver, lead ore. It was not until 1964, however, that the zinc concentrate drying plant was constructed. In 1970, milling capacity was increased from 4,000 to 5,000 tons per day. The zinc content of the ore is quite variable so that the tonnage of zinc concentrate dried varies from 200 to 650 tons per day.The drying plant was designed to dry 25 tons of zinc concentrate, dry basis, per hour from an initial moisture of 8.5% to a final moisture of 2%. Major components were selected so that capacity could be increased to 35 tons per hour with minor modifications.The rotary dryer is a Hardinge type ""XH-12"" with a diameter of 80"" and a length of 50 feet. The all welded construction shell is 1/2 inch thick mild steel plate with 3/4"" thick reinforcing bands under the tires arranged for operation at a slope of 1/2 inch to the foot. The 3/8"" x 12"" radial flights are welded to the shell. There are no flights in the last ten feet of the shell in order to reduce dusting. This unit is driven by a 50 HP motor through a Hamilton type CD250 double reduction helical gear speed reducer. The shell speed is 5.8 RPM.The standard Peabody horizontal air heater has a rated capacity of 10,000,000 B.T.U. 's/Hr. The combustion chamber is lined with high temperature refractory brick and is air cooled. The conical outlet section of the air heater is lined with insulating fire-brick. The burner is a type AT wide range mechanical atomizer fired by a type ""CR"" gas-electric ignitor.The Bunker ""C"" fuel is supplied by a Compower duplex pumping, heating and straining unit, including tank hotwell, Powertronic oil line heating system with a 550/24 volt transformer."

Jan 1, 1972

By Angela L. Paveley

Economic viability of base metal mining is complicated by a volatile market that can cause fluctuations in the growth and stability of the mining industry. Nickel is mined primarily for use in manufacturing industrial stainless steel and plating materials. It is essential that nickel production be sufficient to respond to current and future market demand. A Post-Pillar cut and fill mining complex is currently producing nickel sulfide in Northern Manitoba. From 2006 this area has shown a steady increase in the annual production, rising from 11 million pounds to 25 million pounds in 2010, and accounting for 33% of the total production for the Manitoba Operations. It is expected that fine-tuning the mining process can safely enhance ore recovery, improve production consistency, and increase economic viability of the deposit. The Post-Pillar Optimization study was initiated by the local technical and operational departments at the mine and analyzed from July 2010 to March 2011. The purpose of this study was to determine if portions of non-recoverable pillars can be slashed without jeopardizing pillar strength. This ideal case would optimize ore recovery and reduce variability in the ore mining rate. Numerical modeling and geo-mechanical material properties analyses, completed by the rock mechanics group, define the maximum drift width at 32 feet and the minimum pillar width at 18 feet. Thus, it was suggested that up to 3 feet could be slashed from two sides of the standard 24 foot pillars without damaging ground integrity. The field data collected indicates that additional ore could be recovered without compromising the safety of the workplace and the integrity of the complex. Consequently, technical and operational groups work jointly to achieve pillar optimization. Approximately 500,000 pounds of nickel was extracted from the complex via pillar optimization from July 2010 to March 2011. Economic analyses show the pillar optimization method developed for the mining complex is a proven, value-added opportunity. The mining complex remains a safe workplace and has retained ground integrity.

Nov 1, 2011

By A. J. Basu

Air quality in underground mining operations can fluctuate significantly during the working day, but until recently the only way to ensure a mine had sufficient air quality was to over-ventilate and, with significant delay in adjusting the volume or direction of the airflow, working conditions were tolerable but rarely ideal. Underground ventilation management systems have been considered to achieve improved energy efficiency since at least 1995 in Canada. However, Mine Ventilation Optimization, (MVO) and integrating it with Ventilation on Demand (VOD) is only becoming more of a reality in the last seven (7) years. A systems approach to MVO takes into account all the elements from the point the power is distributed into the ventilation fans through to the actual quality and quantity of fresh air to the working face. Opportunities exist throughout the system for energy savings but not all of them are cost-effective. Depending on the application, these opportunities can be easily realized by properly sizing, streamlining, maintaining and operating the ventilation system using premium mine proven energy efficient components throughout. Implementation of VOD is a key step in achieving the MVO objective. However, VOD is a concept, and it requires various tools, procedures, and strategies for cost-effective and operationally viable implementations. This paper provides a historical background in the development of VOD and provides a framework for proper and cost effective implementation. In addition, an optimized ventilation system will also allow increased airflow by reducing resistances and shock losses. This turn in would allow production target improvement, which could be an attractive objective compared to energy savings. Mine ventilation is the life support system of an underground mine, and contributes to sustainable development and corporate social responsibility goals assuring health and safety aspects of strategic planning. The proposed framework addresses implementing mine ventilation optimization integrating ventilation on demand as a vehicle for energy, health and safety improvement and overall productivity gains in the mining cycle efficiency while significantly reducing carbon footprint of a mine.

Aug 1, 2013

By Caitlyn J. McKinley, Ahmad Ghahreman, Maziar E. Sauber

Ever-growing global needs for critical metals, such as rare earth elements (REE), and base metals, such as copper and nickel, and their environmental impact motivated the pursuit of sustainable flowsheets. Regeneration of hydrochloric acid in chloride-based leaching flowsheets as a critical unit operation is an example of such endeavours. This study presents an efficient hydrochloric acid regeneration method from a magnesium chloride solution waste by addition of sulphuric acid in relatively low temperatures. The effects of temperature, initial magnesium chloride concentration, amount of sulphuric acid added, and airflow are examined. Experimental work are focused on temperatures between 80 and 95°C, initial magnesium chloride concentrations of 2 to 5 M, 100% to 220% stoichiometric sulphuric acid additions, and airflows between 200 to 600 mL/min. The developed technology recovers hydrochloric acid at a high rate and concentration with a relatively small environmental footprint. INTRODUCTION The mining and metals sector is under increasing environmental and economic pressure to make more efficient use of chemical reagents and energy in the production process. This affects several areas of process and flowsheet design such as the selection of the most appropriate unit operations and the overall design of the process steps paying heed to sustainability. Owing to the inherent inefficiencies of physical and chemical processes used in the mining industry, there will be materials and streams generated that are not a desirable part of the envisaged flowsheet. Thus, methods are implemented to recover and recycle such effluents. If the recovery and recycling are not economical, then procedures are followed to neutralize and dispose these effluents. In many cases, these effluents are potential contaminants or hazardous materials, and their disposal should be looked at in a much larger context. The safe management of the contaminants and hazardous effluents is a daunting problem in the mining industry and is fraught with technological, economic, social, and political challenges.

Jan 1, 2019

By J. F. Perreault, J. Paraszczak

"The profitability of an underground mine depends to a great extent on the production cost.Random equipment failures and their consequences have significant influence on the total operating cost of a mining system. Having this in mind, a reliability study has been initiated for loadhaul- dump machines in an underground mine. The main purpose of this study is to identify the critical machine subsystems, determine their failure pattern and help work out an effective maintenance strategy. In the first part of this study, the failure data of diesel powered LHDs, gathered during an I8-month period have been analyzed. For the purpose of the analysis, the machines have been treated as systems comprising seven subsystems connected in series. The most critical subsystems, i.e. electrical and hydraulic, were further investigated, and analyzed by graphical and analytical methods. Different failure distribution models were applied and discussed. Basic Defmitions and ConceptsIn order to avoid confusion, several terms used in this paper are defined below, as compiled by O'Connor (1991). The acronymsBS and MIL-STD stand respectively for: British Standard and US Military Standard.Failure: the termination of the ability of an item to perform a required function (after BS 4778).Time Between Failures (TBF): the length of time between two consecutive failures.Observed Mean Time Between Failures (MTBF): For a stated period in the life of an item, the mean value of the length of time between consecutive failures computed as a ratio of the cumulative observed time to the number of failures under stated conditions (after BS 4778).Maintenance, Corrective: The actions performed, as a result of failure, to restore an item to a specified condition (after MIL-STD-721 B).Maintenance, Preventive: The actions performed in attempt to retain an item in a specified condition by providing systematic inspection, detection and prevention of incipient failure (after MIL-STD-721B)."

Jan 1, 1994

By Robert B. Parsons

"The junior mining company's bank account is depleted, but there are high hopes for the company's property. Where to get the necessary development funds is a common problem facing many juniors.This month, Tax Notes discusses briefly the principal income tax implications of several alternative development plans that are typically available to the junior mining company and its shareholders.IntroductionAt some point in time, many junior exploration companies face a common dilemma-a promising prospect has been proven up and is read y for further development, but alas, the company is flat broke! In this situation, a number of alternatives is available to the company, including the possibility of ""going it alone"" by raising capital through a public offering, or by striking a deal with a ""major"" or some other substantial investor.This article focuses on the latter possibility, and discusses in broad term s the main income tax implications of various arrangements between a typical junior exploration company and a larger public company for the development of a particular property. It will be appreciated that a discussion of other important considerations, such as those pertaining to securities exchange regulation s and company law requirements, is beyond the scope of this article.Jointly Owned CorporationPerhaps the simplest way of dealing with the problem of financing the development of a property is to sell the property outright, pocket the proceeds, and run! Clearly, though, this alternative is unacceptable to most prospectors and developers, who are predictably reluctant to completely relinquish their equity interest in a promising property. An additional strike against such an outright sale is that the sale could trigger an income tax liability. Generally speaking, in the case of a simple sale of a mineral property, for income tax purposes the proceeds are applied to reduce the taxpayer' s unclaimed balance of Canadian development expense (""CDE""), with any unapplied excess being brought into income and taxed at ordinary corporate tax rates."

Jan 1, 1983

By J. A. Eisele, D. L. Pool

"The Bureau of Mines has developed a particle agglomeration technique to improve the flow of leaching solution through heaps of low-grade gold-silver ores which contain clays. Benchscale experiments and industrial practice have shown that the percolation rate of cyanide leaching solution is greatly enhanced by mixing the ore with Portland cement as a binder, moistening the mixture, mechanically tumbling to form agglomerates, and aging before heap building and leaching. A slight modification ofthe technique allows its application to precious-metal-bearing tailings, which cannot normally be treated by heap leaching because they are too fine. IntroductionMany low-grade gold and silver deposits and tailing materials from former mining operations in the western United States are potential sources for precious metals. These sources are too low grade or too small to support the capital cost of a conventional cyanidation plant but have stimulated the development of lowcost procedures, such as extraction of the precious metals by heap leaching with cyanide. The Bureau of Mines has investigated methods to improve the recovery ofgold and silver by heap leaching these low-grade materials.For heap leaching to be successful, the crushed and stacked ore must have good permeability so that uniform distribution of the cyanide leaching solution through the heaps can be achieved. Some of the most difficult gold and silver ores to treat by heap leaching are those containing excessive amounts of clays, or fines generated during crushing. Excessive amounts of slimes, generally minus 50-micrometre or 270-mesh Tyler sieve size particles, in the heap feed will slow the percolation of the leaching solution, cause channelling, or produce dormant or unleached areas within the heap, resulting in long leaching periods and low extractions. In extreme cases, the clays or slimes can seal the heap and cause the leaching solution to run off the sides of the heap rather than through it."

Jan 1, 1987

By J. -P. Tshibangu, M. Ramos da Silva, F. Descamps, C. Schroeder, J. -C. Verbrugge

"The mechanical behaviour of rock materials is largely influenced by the stress state and pore pressure; this needs to be considered in many applications, such as petroleum engineering, tunnelling or stability of underground excavations.Conventional triaxial tests (d2 = d 3) are generally used to assess the influence of confining and pore pressure. In this work, those tests are performed on a porous rock named Sorcy limestone, along with more advanced stress paths like triaxial extension (d 2 = d 1) and purely deviatoric tests with independent pore pressure control. Those tests are combined to a true triaxial study on the dry rock in order to investigate the effect of true triaxial stresses and, hence, build three-dimensional limiting envelopes accounting for the influence of pore pressure.Three-dimensional limiting envelopes of dry Sorcy limestone are built in meridian and octahedral planes. The shape of the octahedral sections changes with the mean stress from triangular to hexagonal and then quasi circular. At even higher mean stresses, the evolution continues towards a triangle with apexes oriented in the direction of triaxial extension stress states. In meridian sections, the deviatoric stress first increases with the mean stress. The transition between this deviatoric mechanism and the cap surface occurs at different mean pressures depending on the Lode angle.Terzaghi’s effective stress is found to govern the yield and plastic behaviour of saturated Sorcy limestone. Triaxial tests show that the limiting envelope does not depend on the loading path (conventional compression and purely deviatoric stress paths). In triaxial extension, the brittle-ductile transition is observed for a higher mean pressure than in triaxial compression, as previously found for true triaxial tests on the dry rock.Finally, it is shown that upon appropriate normalisation, the dry and saturated yield envelopes come very close one to each other. This suggests that only the size of the yield envelope changes between dry and saturated state (pore pressure = 5 MPa); its shape does not seem to evolve."

Jan 1, 2015

By Simon Houlding

"Part 1 (CIM Bulletin, August 1991) of this article identified the limitations of the traditional modelling methods and demonstrated the approximations and errors introduced by their application to complex mining situations. The objective of Part 2 is to describe the evolution of new, proven technology which eliminates these deficiencies and provides precise and practical alternatives. These promote the use of realistic geological and mining shapes and eliminate altogether the rectangular approximations of traditional methods. Limitations of the Traditional MethodsThe limitations of the traditional modelling method s (viz. simple block modelling, gridded modelling and their more recent variations) were discussed in detail in Part I . These methods are all based on the use of rectangular discretization to represent the boundaries and volumes of ore deposits and mining excavations. As a result they are unsuited to the representation of nonrectangular geological complexities such as folding, faulting and irregular surfaces . They are equally unsuited to the representation of realistic excavation limits, particularly in the case of underground mining.Recently these methods, with limited variations in technique, have been applied to increasingly more complex geological and mining situations. The resulting potentials for error and conflict within the modelling pro cess were demon strated in Part I . The effects can be summarized as follows:• volume and boundary approximations introduced by therepresentation of irregular three-dimensional boundaries andvolumes by rectangular modelling elements ; • geostatistical errors and conflicts caused by neglect of the relationships between grade estimates and the geometric/ spatial aspects of irregular volumes and associated observations (samples), including an inability to produce precise estimates with meaningful confidence limits;• a compounding of these effects by using rectangular element sizes other than the geostatistical optimum, and by using differentelement sizes within a model; and• general inefficiencies introduced throughout the modelling pro cess by the requirement to represent geology, grade distribution and excavation limits with a singular modelling method and model."

Jan 1, 1991

By D. A. Stokes

"In October of 1980, Fording's 15 000-ton capacity clean coal silo was put into operation. The silo was unique in that it was designed to allow the mass flow of material, coupled with a sophisticated computer-controlled dynamic loadout/weighing system.On May 1, 1982, an explosion, attributed to a methane buildup, devastated the top of the silo. .As a contingent plan to minimize the effect of the production loss the conveyance to the silo top was reconstructed and the silo' was back in operation by mid-June 1982, although the silo top was left open to prevent further methane buildup. As dust emissions occurred during loading ofthe silo, an extensive study was undertaken to establish dust control for the silo, while at the same time providing methane detection and control.The study, conducted for Fording by the Cambrian Engineering Group of Calgary, involved an extensive search and investigation of existing coal silos. The study concluded that there IS no universally accepted method of controlling methane emissions in coal silos.The study also produced a unique conceptual design for handling methane, while at the same time, protecting the environment . Fording's Clean Coal SiloThe silo, located at Fording Coal's Fording River metallurgical coal operation, is utilized for the storage and load out of cleaned thermally-dried metallurgical coal.The silo was constructed in 1979 and 1980 to reduce unit train loadout times and to facilitate the handling of “special shipments"" of coal. The silo went into operation in October 1980, under manual control and utilized computerized, automated loading by February 198I.The silo is 21 m in diameter and 76 m high, with a nominal capacity of IS 000 live tonnes. The coal storage portion of the silo is 21 m in diameter and 48 m high, utilizing 2 steep-angled hopper outlets allowing mass flow of the stored coal during train loading. Each hopper has a control gate and a discharge chute, one chute for bulk loading, the other for trim loading of coal into the rail cars."

Jan 1, 1986

By P. K. Chatterjee

"The operations of a large-scale open-cut mine are simulated to produce a user-oriented interactive digital model. SIMULA is chosen as the programming language and a sophisticated data base system is designed to consider the complex information network of large mines. The main features of the program are the mine planning and mine evaluation modules. These modules are responsible for the computation of truck-shovel performance and other equipment details in a mine. Stochastic methods are employed to generate event times. The use of a Data Base Management System and the data base provide an efficient method of information storage and retrieval. Truck allocation techniques have been examined in detail and comparisons between random dispatching and conventional methods are made. The model is intended for mine operators and engineers and the program is well structured and very readable. It is hoped that these features will make the model readily acceptable.IntroductionThe Department of Mining and Metallurgical Engineering at the University of Queensland has been conducting research into computer simulation methods in mining since 1972. The research has been mainly applied to actual operations, with both open-cut and underground mining being equally emphasized. The results of the research to date have not only strengthened the methodology of the approach to complex mining problems, but have also increased confidence in the use of modelling and computer simulation technique s in mining.The model described in this paper translates a complex LHD situation into a user-oriented interactive digital model using SIMULA. The present methods of evaluating truck dispatching and trolley assist are tedious and laborious. This model provide s the engineer with readily acceptable means of rapidly assessing the relative merits of these techniques in a large-scale open-pit operation. Poor utilization figures in many current operations (for example, 40% for trucks, 70% for shovels) are of serious concern to mine officials and computer-controlled dynamic truck dispatching systems have been considered by some of these mines to improve equipment utilization ."

Jan 1, 1981

By Frank Lemy

This paper presents an application of digital face mapping in an underground mine. Based on photographs it was possible to construct discontinuity maps, which were used to characterise the rock mass. The results were compared favourably with values obtained through scanline and area sampling mapping for the same site. The case is made for digital mapping as a safe, fast, and efficient way to characterise rock mass structure for engineering purposes.

May 1, 2001

By E. H. Mulligan

THE ERRINGTON B-I mine is located on the Errington orebody of Steep Rock Iron Mines, Limited, at Steep Rock Lake, Ontario. In 1946, this orebody entered the shipping lists of the Lake Superior iron ranges as an open-pit operation. During the period 1946 through 1953, the mine produced a total of 9,000,000 tons of iron ?ore. In 1950, as the mine approached the economic limits for open-pit operations, ground was broken for an underground development to exploit the deposit at depth. The mine is designed for an ultimate production of 1,500,000 tons per year by the block-caving system of mining. The initial shaft is 1,200 feet deep and will produce or-e from the 700 and the 900 levels. The 1,100 level will be used as a drainage level and as a haulageway for waste rock, which is hoisted to surface by skip. Ore from the two production levels moves by a system of ?ore passes to a crushing plant located below the 900 horizon and is hoisted from this point to surface by conveyor belt. At the present time, all material handling on the levels is by tramming, but as tram distances increase conveyor haulage will probably be installed. The initial stage of the Errington B-1 development is expected to produce 6,000,000 tons of ore over a period of four years. The second stage will require deepening of the shaft and development of lowe1? production levels. The conveyor hoisting plant will be extended to serve these new levels. We visualize the mining of this orebcdy down to 2,500 feet by these methods.

Jan 1, 1956