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By H. Harzer

Statistics Canada record s indicate that in recent years 70 to 80% of the entire Canadian coal production originated from Western Canadian mines, in which about 55 to 65% of the coals are bituminous. Until recently, less than 20% of the latter were extracted by underground methods , the remainder being surface-mined. However, it is estimated that within the foreseeable future up to 80 or 90% will have to be gained by underground techniques. Although Western Canadian coal resources are considerable, they are unfortunately locked in severely distorted, steeply pitching and faulted seams of variable thickness. Consequently, their economic extraction poses serious engineering problems. One possible solution to the problems is to apply hydromechanical methods to underground mining. This is a well-proven method shown to be economically viable under Western Canadian conditions. This paper endeavours to widen the scope of published data on hydraulic coal mining and is based on practical engineering and cost data gathered through contacts with West German designers and operators of hydraulic mines.

Jan 1, 1978

By V. Eu. Anufriev

Despite of significant cost advantages and the positive expierence roof bolting is limited up to 15-20% of all workings at the Cuzbass mines. The low level of application of the technology is the result of specific mining and geological conditions and defects of organization and technical support. When analysing mining and geological conditions of development working maintenance and geomechanical aspects the necessity of changing of technological schemes of preparation and extraction of mine sections is obvious. The prospects of large-scaled transferring from frame support to roof bolting are connected with introduction the technology of support with the help of non- expanding bolt classes including bolts with mineral filler fixation. The design of bolters with mineral filler fixation is given here. The results of testing of bolt load characteristics in artificial holes and in situ are proved the technology efficient. For increasing of drilling works efficiency it is necessary to master the production of oilless compressor and heading machines with the manipulator. For bolt dimension forecast it is recommended to use the relations of measured parameters of "dome" to roof shifts. Effective monitoring of rock pressure in the junctions is connected with the use of support of conventional bolts together with bolts of deep installation with simultaneous supporting of face ends with special mechanical sets, which compensate face length change.

Jan 1, 1996

By R. G. Gillies

The pipeline transport of paste, backfill or thickened tailings slurries is an important component of the tailings disposal process in many mines around the world. These mixtures contain significant fractions of fine (clay) and coarse (sand) particles, causing them to exhibit non-Newtonian flow behaviour and high frictional pressure gradients. There is often economic incentive to operate these pipelines in the laminar flow regime. Irrefutable evidence has shown that coarse particle settling can occur during the laminar flow of these slurries even when static settling tests show no indication of such behaviour. Previously, a ?rule-of-thumb? minimum frictional pressure gradient of ~1.5 to 2 kPa/m was suggested to ensure effective sand transport, based on the results of tests conducted using viscous Newtonian oils. In this paper, we examine the results of laminar flow tests of non-Newtonian, coarse-particle laden slurries conducted with pipeline loops of 50, 100, 150 and 250 mm in diameter. The results indicate that the ratio of the mean wall shear stress to the mean surficial particle stress is a better indicator of laminar flow settling tendency than frictional pressure gradient.

Jan 1, 2007

In 2005, CANMET-MMSL used its pool of expertise in ground control, open pit and underground mine engineering, mineral processing, information technology and applied mineralogy to launch a mine-to-mill research project. The objective of the project was to optimize the distribution of rock breakage energy among blasting, crushing and grinding activities, thereby maximizing the efficiency of energy utilization within comminution processes. Tests were designed to demonstrate that this approach could reduce energy consumption, thus contributing to the reduction of green house gas (GHG) production in the Canadian mining industry. The project was accomplished in collaboration with DynoConsult and Dyno Nobel Canada Inc., a world leader in commercial explosives. Data analysis from 12 blasts (six with electronic and six with pyrotechnic detonators) in three different geological formations of a quarry (soft, medium and hard rock) shows that electronic blasting technology increases rock fragmentation by 15% (in hard rock) to 20% (in soft rock). A more uniform rock fragment size distribution is also obtained. The impact of the improved quality blast product on downstream processes of the quarry is a productivity increase of 5% in rock excavation and transportation, an energy reduction of 10% at the primary crusher and an increase of 15% of the primary crusher throughput. Extended to the Canadian mining industry, such results would represent a GHG emission reduction of about 0.05 Mt/y. A soft-sensor prototype, based on primary crusher data (i.e., energy consumption, throughput, feed and product size distribution) and a simplified computerized process model, was developed for on-line measurement of the rock operating work index. Additionally, 385 rock samples and specimens were collected and prepared to determine the tensile and compressive strength of the rock material, to evaluate the aggregate crushing work index and to examine the development of micro- and macro-fractures by means of binocular and SEM microscopy. Blast-induced rock damage was evaluated at the end of the laboratory-testing program. Even though a reduction of rock hardness of around 30% was observed after blasting, the impact of the blasting method (i.e., use of electronic or pyrotechnic detonators) on the induced rock damage could not be observed. It is either absent or not perceptible with the accuracy of the measuring techniques and the number of samples analyzed.

Jan 1, 2008

To optimise the haulage of the large tonnages produced from stopes in this high production gold mine, an ore pass and feeder system was introduced. This ore pass was designed and developed in 2006 and the plate feeder installed and commissioned soon after. Cavity Monitoring Surveys (CMS) have been undertaken of the ore pass since it was commissioned. The original thickness that was surveyed of the pillar between the ore pass and decline in 2006 was approximated to be 19 m. Throughout its use large blocks of rock were reporting at the plate feeder. This was the first indication that the ore pass was failing somewhere between the tipping level of the 9410 level and the plate feeder at the 9330. Even so CMS surveys were not taken systematically to firstly identify the area of failure or for monitoring. In August 2008 after implementing management systems, large blocks were still reporting to the plate feeder. The decision was made to empty the pass enough to carry out a CMS survey of the area. Subsequently it was discovered that the area of failure was between the decline and ore pass and that since the first survey the pillar thickness had been reduced to approximately 9.7 m. Further management controls were put in place to slow movement of the material and to limit the use of the 9410 tipping level, but this did not alleviate the rate of failure entirely. The results of an initial inelastic numerical model indicated that most of the pillar was damaged and that failure into the decline likely if no changes were made. The numerical modelling predicted that even if material was no longer passing through the ore pass interaction between the voids was unavoidable and would most likely result in instability. The ore pass and plate feeder are an integral part of the haulage system for the remaining life-of-mnie. Two solutions were sought to firstly reinforce and stabilise the pillar between the ore pass and decline and secondly to rehabilitate the ore pass or find an alternate design that could still utilise the plate feeder. This paper summarises the analysis and resulting rehabilitation that was used to accomplish this.

Nov 25, 2010

By Michel Aubertin

Les digues de retenue de résidus miniers devraient être conçues de façon à supporter les charges les plus défavorables qui puissent être anticipées. Les nombreux incidents survenus au cours des dernières années à travers le monde montrent toutefois que la stabilité de ces ouvrages reste encore problématique. Les cas de rupture les plus courants sont reliés aux glissements le long des pentes, à la liquéfaction des rejets, au débordement des bassins, ainsi qu?à l'érosion régressive et de surface. Après avoir rappelé la nature et l?importance des certains incidents récents, l?origine de ces problèmes est brièvement présentée. On aborde ensuite le lien entre les critères à adopter pour la conception des ouvrages et les risques associés à leur rupture. De nouvelles avenues pour aider à contrôler certains problèmes sont enfin discutées, en mettant l?emphase sur la codisposition des roches stériles dans le parc à résidus miniers.

Nov 1, 2002

By S. Frimpong

The cable shovel excavator is used for primary production in many surface mining operations. A major problem in shovel excavation is the variability of material diggability, resulting in varying mechanical energy input and stress loading of shovel dipper-and-tooth assembly across the working bench. This variability has a huge impact on the shovel dipper and tooth assembly in hard formations. In addition, the geometrical constraints within the working environment impose production limitations resulting in low production efficiency and high operating costs. A potential solution to the above problems is the application of an intelligent cable shovel excavation(ICSE) technology. In this paper, dynamic models of the cable shovel excavator are developed with numerical examples. These models form the basis for developing dynamic shovel simulators for an intelligent shovel excavation technology. The dynamic equations are developed using the Newton-Euler techniques. The path trajectories for efficient excavation, dynamic velocity and acceleration profiles, and dimensioned parameters for optimum feed force, torques and momentum of shovel boom-bucket assembly are modeled and used for efficient excavation. The optimum digging forces and resistances for the cable shovel excavator are modeled and used to predict optimum excavation performance.

Jan 1, 2003

By Christian Neumann, Josef Arnold, Albin Reinhart

The Channel Tunnel Rail Link (CTRL) is the most significant railway infrastructure development project in the United Kingdom since the construction of the Channel Tunnel between France and England. The key element of this project is the 2-mile-long(3.2-km) North Downs Tunnel, which will be the largest and deepest twin-track railway tunnel ever constructed in the United Kingdom. This paper describes the contractual agreement, the design and construction (excavation and lining) of a twin-track rail tunnel with a 166-m 2 face excavated through chalk using the New Austrian Tunnelling Method (NATM). The importance of collaborating with a tunnel contractor/designer who not only is expert in the New Austrian Tunnelling Method, but who was also responsible for the design and installation of the primary lining, will be highlighted in this paper. This paper will also deal with the benefits of partnering in tunnel design and construction, as well as the resultant added values of mechanizing conventional tunnelling methods (Value Engineering).

Jan 1, 2001

By X. Bai

A new heuristic sublevel mining stope optimizer is presented. The optimizer seeks the best locations and lengths of a series of vertical raises that, together with the blocks linked to each raise, define a mining stope. Five design constraints ? the footwall angle, the hangingwall angle, the number of raises, the maximum distance of a block from a raise, and the minimum width required to move the farthest block towards the raise ? allow the shape of the sub-stopes associated with each raise to be controlled. The optimization is done on the locations and lengths of raises using a genetic algorithm to efficiently sample the parameters? space. For each raise, a local network is defined in cylindrical coordinates around the raise such as to impose the design constraints. A maxflow algorithm on the local network is used to determine the optimal sub-stope for each raise. All substopes are combined to define the global stope for the entire deposit. The best global stope is obtained using a genetic algorithm to find the raise parameters providing the best profit over the entire deposit. Two synthetic cases and one real deposit are used to evaluate the new algorithm and compare the results with the single-raise optimizer. The multiple raises approach leads to significantly improved economics compared with the single-raise stope optimizer, and the dilution is also substantially reduced compared to the single-raise case.

Jan 1, 2014

A survey has been undertaken of the petrography of North Island greywackes with a particular focus on the occurrence in them of smectite (swelling clay) and chlorite, two minerals that are closely associated, chemically similar, and commonly interlayered with each other. Smectite is the mineral which is a major cause of greywacke aggregatesÆ poor performance. Smectite occurs in post-metamorphic shear zones and in surface weathering profiles over greywackes. This study focuses on the petrography of fresh greywacke rocks.   The occurrence and abundance of smectite and chlorite and their relationship in greywackes is largely a function of the chemistry of the sediment and the maximum temperature that the sediment has been subjected to during diagenesis and/or very low grade (anchizone) metamorphism. Rock chemistry, particularly total iron and MgO, determines the smectite and/or chlorite content in fine-grained greywackes (argillites and siltstones) but chemistry is less influential in sandstones, where at low grades of metamorphism much of the Fe and Mg of the rock is held in unreacted detrital minerals and the mineral constituents of lithic grains.   Four zones which represent different levels of diagenesis and very low grade metamorphism, based on the crystallinity of illite and chlorite in the greywackes, and the proportion of smectite interlayered with chlorite, have been mapped in the North Island. In the anchizone smectite does not appear as a discrete mineral, it is always interlayered with chlorite. In the upper anchizone chlorite contains approximately ten per cent of smectite interlayered with chlorite. In the lower anchizone about 30 per cent smectite can occur in chlorite.   In addition to the smectite and chlorite inherent in the greywackes as the result of the thermal overprint, smectite as a discrete mineral and smectite-chlorite interlayered minerals with high levels of smectite, occur in abundance on shear zone surfaces and in alteration zones adjacent to shear zones; these are the main contributers to the high levels of smectite recorded in many greywacke aggregates.

Jan 1, 2006

By Doug Smith, Gary Taylor, Paul Ogryzlo

"Uranium mining and processing typically requires large quantities of high-grade reactive quicklime for the neutralization and clean-up of contaminated mine water and mill process waste streams. Currently the Saskatchewan uranium industry’s lime requirements are being supplied from deposits located in Manitoba and Alberta, resulting in transportation costs, which equal or exceed the market value of the product. During the summer of 2001, the northern village of Pinehouse located 220 kilometers south of the Key Lake uranium mill, staked two previously discovered limestone deposits. The relatively shallow deposits, known as Missi Island and Sleeping Giant are located in close proximity to one another, approximately 35 km south of Pinehouse adjacent to highway 914. This paper describes the results of a scoping study carried out by AMEC for the development of the two deposits based on a seasonal limestone mine and a continuous 40,000 to 60,000 tpy wood fired calcining plant designed to supply all of the quicklime needs of the Saskatchewan based uranium industry. Also described are the results of calcining and customer acceptance testwork on samples of lime prepared from recent Pinehouse limestone drill core. The potential benefits of the project include a significant new source of revenue for the community of Pinehouse, lower lime costs for the Saskatchewan uranium industry and employment for northerners in the mining, processing and wood harvesting operations.IntroductionInterest in the Pinehouse/Massinahigan River area for limestone was precipitated by the results of a 1977 industrial minerals drill hole program by the Saskatchewan Department of Mineral Resources. These results led to the acquisition of land in the area by Missi Island Mines Ltd who carried out extensive exploration during the period 1978-1982. They discovered a deposit on the southwest side of the Massinihigan River thought to contain approximately seven million tonnes of limestone with an average grade of 94.5% CaCO3. Missi Island carried out preliminary feasibility studies and prepared an Environmental Overview Statement. However, in spite of positive recommendations, the project was dropped in the mid 1980’s."

Jan 1, 2003

By H. Naghavi, M. Karimi

"In self-aerated flotation machines, the gas rate depends on operational variables (e.g. froth depth and impeller speed), pulp properties (e.g. solid content and viscosity), and reagent addition (e.g. type and concentration of frother). The gas rate has a strong correlation with the flotation performance by influencing the gas dispersion properties and froth retention time. A factorial experimental design was used to study how the gas dispersion properties, the froth retention time, and the flotation performance respond to changes in froth depth and impeller speed (as the most common operational variables). An in-depth understanding of the effects of impeller speed and froth depth on the gas dispersion properties, especially the bubble surface area flux and froth retention time, is necessary to improve operating strategies for self-aerated flotation machines. All experiments were carried out in a 50 m3 selfaerated flotation cell in an iron ore processing plant. The results showed that the froth depth affected the metallurgical performance mostly via changing the froth retention time. The impeller speed had two important impacts on the metallurgical performance via varying both the froth retention time and the bubble surface area flux in the froth and pulp zones, respectively. The interaction effects of the froth depth and impeller speed were also established. This allowed us to develop a strategy for operating selfaerated flotation machines based on varying the froth depth and impeller speed with regard to the cell duty. IntroductionA mechanical flotation machine may be divided into two distinct zones. namely the pulp zone and the froth zone (Goodall and O’Connor, 1989; Yianatos, Bergh, and Cortes, 1998; Rahman, Ata, and Jameson, 2015a, 2015b). The overall flotation recovery, Ro, including the true flotation (particles attached to the bubble lamellae) and entrainment (particles recovered in water held in the bubble plateau boundaries), is a function of the recoveries in the two zones. It is calculated as follows (Dobby, 1984):"

Jul 1, 2019

Heavy mineral deposits are becoming more complex in terms of their compositional variation, particle mineralogy and size distributions, and present challenges for the operation of separation equipment to achieve the required grades and recoveries. Changes in equipment design and concepts potentially provide new opportunities for the beneficiation of heavy mineral sands deposits. The development of a framework for the fundamental understanding of separator performance issues is a first step to develop new machines. This paper will present a first step analysis of electro-static and magnetic separation machines with their application to real separation scenarios.

Jan 1, 2010

By Pefer F. Stacey

This paper considers some of the trends in slope design technology that are emerging as we approach the 21st Century. While the ability to design stable slopes has unproved greatly over the past 20 years, trends towards steep short-term slopes for small highgrade orebodies, and very high slopes for mega-pits centered on low grade deposits are starting to indicate limitations in available technology.

Jan 1, 1994

By M. E. Hawkins

This Bureau of Mines report provides information on all of the 329 proved salt domes in the Gulf Coastal States and offshore area of the United States. It includes specific data related to location, depth to the cap rock and to the salt, storage facilities for LPG, and production of salt, sulfur, petroleum, and rock related to or associated with the dome structure in the subject area. Also included is a brief history of the salt, sulfur, and petroleum industries as related to the salt domes. Gulf Coast salt domes were the scene of the first discovery of a large deposit of rock salt in North America, the birthplace of the Frasch sulfur industry, and the location of the first large oil gushers that ushered in the modern liquid fuel age.

Jan 1, 1966

By K. L. Whitehead

An examination of 60 mines, 11 illumination hardware manufacturers, and Q mobile face-equipment manufacturers was conducted to determine the state of the art of mine-illumination hardware, define lighting system installation and maintenance problems, and assess personnel acceptance of lighting. Longwall and mobile face-machine illumination were studied. Major problems on room and pillar installations include damage of exposed components, electrical failures, and reduced lamp life, with extreme variability in the degree of these problems. Most longwall systems had low or moderate maintenance requirements, cable damage being the most significant problem. Most mobile face-equipment operators liked their machines' lighting systems, but many complained of glare and other problems. Longwall lighting was unanimously accepted by face personnel, and most complaints were minor. Based on the survey, general installation and maintenance guidelines have been established, addressing hardware design aspects, approaches proved effective to minimize problems, and improper practices noted which are detrimental to effective lighting.

Jan 1, 1981

Splashing and wave formation has been studied experimentally by using one-seventh water model of CLU (Creusot-Loire Uddeholm) reactor with two different tuyere configurations by using paraffin and water as simulated slag and simulated metal respectively. Experimental results have been discussed in terms of total splash volume, relative slag volume in splash (dispersed-slag phase holdup in splash), wave frequency and wave amplitude with respect to gas injection rate and tuyere configuration at bath heights between 0.25m and 0.38m. Gas flow rates ranged from 0.00599m3/s to 0.01312 m3/s. The results of splash measurements showed that splash volume increased with increasing gas flow rate and decreasing bath heights for both tuyere arrangements. However, off-centre configuration resulted in relatively higher amounts of splash. The slag volume in splash (Dph in splash) was also found to increase with decreasing bath heights and increasing gas injection rates. The high speed photography revealed that the frequency and wave amplitude increased with gas injection rate. The experimental results also showed that the frequency and vertical displacement of the rotating wave increased when the tuyere orientation became off-centered. This might be used to explain why off-center configuration resulted in generating higher splash volumes than the center configuration.

Jan 1, 2004

By R. L. Zick, J. D. Byars, T. P. Mucho

Recently, a large Pittsburgh seam longwall mine was nearing depletion of a major portion of its current reserves and had begun to develop in a different direction from its original portal area to access additional reserves. This meant that a large area of the mine would be abandoned and sealed. However, continued use of the original mine portal area required that the water accumulating in the abandoned mine would eventually need to be pumped to the surface. This would be in addition to the discharge water associated with the new portion of the mine. Several pumping and sump options were investigated to handle the quantity and quality of the anticipated discharge water. This paper describes some of these options, their advantages and disadvantages, and the final engineering decisions. Some problems and unanticipated outcomes, as well as the eventual solutions are discussed, including: (1) estimates of pump requirements, (2) water pool size, (3) eventual water quality, (4) anticipated inflows, and (5) integration into the overall mine water system.

By Jacques Deraisme

The technology of percussion drilling is well mastered. The lower cost of percussion drilling allows more holes to be drilled for the same budget. However significant differences are often observed between percussion and diamond drill assay data. This is due to the increased level of contamination associated with the percussion drilling process. Statistical studies on twinned drillholes show that diamond and percussion assays cannot simply be merged in a traditional kriging approach. The question then arises on how to use both types of data to produce reliable resource estimates. To take into account the lower reliability of the percussion assays, a noise characterized by a variance is added to these data. The methodology consists of two steps: 1. Cokriging of the equivalent diamond value at the percussion assay locations. 2. Block kriging using diamond assays and equivalent diamond values, calculated previously, adding to these data a noise, whose variance is the cokriging variance of step 1. This particular way of achieving kriging is called kriging with variance of measurement error. The paper illustrates this innovative methodology with an application on an iron deposit. The results of this method is compared to block kriging using only diamond assays and the gain in precision is quantified. An extension of that approach to multi-element cokriging is also presented.

Oct 1, 2009

By W. W. Adams

"The eighth National Safety Competition, that for 1932, conducted by the United States Bureau of Mines, was participated in by 322 mines and quarries operating in 34 States. Tables 2 to 6 show the relative standing of the various plants. Bronze trophies, known as the 'Sentinels of Safety,' which were provided by The Explosives Engineer magazine to be awarded annually by the Bureau of Mines, were awarded to five companies, each company having established the best accident-prevention record within one of the five classes into which all companies were divided, - namely, anthracite-mining companies, bituminous-coal-mining companies, metal-mining companies, operators of nonmetallic-mineral mines other than coal mines, and quarries or open-cut mines. The trophies are awarded for one year and are then transferred to the winners of the succeeding year's contest.The trophy for bituminous-coal mines, which has been in New Mexico for the past year, was won by a mine in West Virginia. The metal-mine trophy goes to New York State for the first time in the history of the contest. Previously it had been won by mines in Minnesota, Michigan, Kansas, and Oklahoma. The trophy for nonmetal mines remains for another year at Retsof, N. Y.; that for quarries and open-cut mines is transferred to a quarry in Michigan. The trophy for anthracite mines, of course, remains in Pennsylvania, as all anthracite mines in the contest are in that State.In determining the relative standing of the various companies enrolled in the competition, all mines and quarries were graded according to their accident-severity rates, - that is, according to the number of days of disability of employees on account of accidents in proportion to the number of man-hours of work done by all employees at the plant. Reports of accidents and of man-hours worked were furnished by the companies to the Bureau of Mines. These reports were carefully examined and classified to insure uniform weighting of accidents according to the number of days of disability to be charged and to insure uniformity of methods of calculating the accident-severity rates for the various mines and quarries. When the relative standing of all plants had been determined in this way, a summary of the records of the companies having the lowest accident-severity rates in their respective groups was sent to the members of a committee of award for examination in accordance with the rules governing the safety contest. The members of the committee of award were W. H. Cameron, managing director of the National Safety Council; William Green, president of the American Federation of Labor; James F. Callbreath, secretary of the American Mining Congress; T. T. Read, Vinton professor of mining engineering of Columbia University; C. B. Huntress, executive secretary of the National Coal Association; A. J. R. Curtis, assistant to the general manager of the Portland Cement Association; and A. T. Goldbeck, director of the bureau of engineering of the National Crushed Stone Association. When two or more mines within the same group operated without an accident during the year, the contest rules provided that the trophy for the group should be awarded to the plant whose record showed the largest number of man-hours worked."

Jul 1, 1933