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By Andrew Cushing, Joe Carvalho, Jon Hurt

"INTRODUCTIONThe shales of the Greater Toronto Area (GTA) have been known to experience time-dependent deformations (TDD); those initiated by excavation-induced stress relief which progress with time as a function of rock porewater salinity, access to freshwater (or air) of lower salinity, clay and calcite content, and the buildup of swelling pressures within the rock mass. These deformations can induce long-term pressures on shaft walls and tunnel linings, especially if the permanent works are constructed soon after excavation, with very little time delay. This paper presents measurements of time dependent deformation of recent shaft and tunnel projects constructed in the GTA, and draws conclusions regarding the main influences on the TDD. The key input parameters for the numerical predictive model are critically assessed. Projects which are discussed include the Billy Bishop Airport Pedestrian Tunnel, Hydro One Midtown Tunnel and Hanlan Feedermain (Contract 3).MECHANISM OF TIME DEPENDENT DEFORMATIONExplanation of MechanismThe Georgian Bay Shale unit consists of typically moderately weathered to fresh, grey to dark grey, fine to very fine grained fissile shale interbedded with slightly weathered to fresh grey, fine grained calcareous siltstone and limestone interbeds. There are two distinctive features of the shale in the Greater Toronto Area (GTA). One is a high horizontal stress regime, and the second is long-term time dependent swelling behavior which develops when the following factors occur (Lo and Micic, 2010):• Stress relief of the rock mass• Outward salt concentration gradient from pore fluid of the rock to the ambient fluid• Availability of fresh waterThe swelling is a consequence of the reduction in confined stress in the rock which occurs upon excavation in combination with a differential gradient in salinity between the saline rock porewater and freshwater or even humid air. Osmotic and diffusive processes result in a decrease in the salinity of the rock porewater achieved by an overall increase in the water content, resulting in volumetric expansion of the shale rock over time. The development of this time dependent deformation (TDD) relative to the time"

Jan 1, 2016

By E. Poeck, R. Garvey, U. Ozbay, K. Zhang

"A coal bump is characterized by the sudden failure of one or more pillars and an associated release of kinetic energy. Although the geologic conditions surrounding coal bumps are often similar, their occurrence and magnitude are difficult to predict. This paper presents the development of an approach to assess the potential for coal bumps in room and pillar mines through the use of energy concepts and special consideration of the interface properties between the coal and overlying rock. Back analyses were performed on the widespread collapse of a room and pillar mine with an associated 3.9 local magnitude seismic event. Pillar and mine-scale models were constructed using the distinct element method to simulate the mining stages leading up to the collapse. A variety of loading conditions, material properties, and interface properties were evaluated for their effect on unstable failure of single pillars, and the mine-scale model was used to study the evolution of failure during progressive mining. The extent of unstable failure was quantified in these models by calculating the total kinetic energy released during each mining stage. Through the parametric analysis of single pillar models, it was found that softening parameters in either the coal or the coal/rock interface can individually facilitate unstable failure of pillars, but the combination of the two in a single model produced a much higher release of energy during failure. The results of the mine-scale model further illustrated the trend as the combination of softening parameters in the coal and coal/rock interface produced a series of failures which most accurately reflected the collapse event at the mine. The method of room and pillar stability analysis demonstrated in this study effectively explores the potential for large coal bumps. INTRODUCTION The analysis of coal pillar behavior is challenging because natural variability in geologic conditions requires estimation of material strengths and predicted loads for a given mining situation. Large width-to-height (w/h) ratio pillars and deep overburden present an additional challenge, as the failure behavior of such pillars is more complex and may involve unstable, sudden collapse. Empirical methods have improved the success of coal pillar design for a wide range of conditions, but for deep cover and squat pillar geometries, less accuracy has been achieved in predicting pillar performance (Mark 2000). As with any empirical approach, more back analysis and the consideration of new variables may help improve understanding and accuracy of future designs. This paper presents a study of pillar stability through the back analysis of the Crandall Canyon Mine collapse, which occurred in August of 2007. Pillar-scale models are utilized for the calibration of material properties and assessment of potential instability under varying conditions, and mine-scale models are constructed in an attempt to reproduce widespread failure and correlate with observations at the mine. Special consideration is given to the properties of the coal and the coal/rock interface, and kinetic energy is calculated throughout the course of the simulations to gauge the level of instability when pillar failure occurs. Models were constructed using the Universal Distinct Element Code (UDEC) (Itasca, 2014)."

Jan 1, 2016

By Y. Xi

In this paper we present an availability approach for determining the capacity of a major bucket wheel excavator (BWE) system to be used in Yuabaoshan surface coal mine in China. Based on previous time studies on other BWE systems for delay times due to internal and external causes, and a four-level maintenance and repair schedule, we develop a model taking into account all random failures and present a fast and simple solution to this highly complex problem. A procedure employing the Markov renewal theory for determination of the distribution of random failure times and failure rate of the whole system, as well as some other factors useful for calculation of the overall system capacity are discussed in the paper.

Jan 1, 1993

By John E. Good

A new coal boom is under way in western Canada. As recently as the mid-1960's, the coal industry in western Canada was considered stagnant and, by some people, dying. But Japan's growing need for coking coal, which became more pronounced in 1966, sparked great interest in western Canadian reserves.' As a result of a rapidly expanding export market and growth in domestic demand, Canadian coal production spurted to 16.6 million tons in 1970. Production in 1965 was 11.5 million tons. Today, about 60 percent of Canada's output comes from western Canada, but by 1975, 87 percent (23.1 million tons) is expected to come from this area, and 96 percent (68.3 million tons) by 1990. Opening, developing and operating new coal mines to meet the needs of a rapidly expanding market for western Canadian metallurgical coals pose some unique and difficult problems. Covering nearly every aspect of opening properties in remote areas, the problems are met in exploration, transportation, construction, mine operations, coal preparation, manpower and ecology. Paul Weir Company has been involved as coal consultants on various projects in western Canada for some years. In presenting this paper, we make no attempt to present an in-depth or detailed analysis of any one specific situation. Each has, or would have, its own peculiarities or characteristics, both physical and economic. Instead, the objective is to provide for those not intimately

Jan 1, 1972

By Ray Nivens, Umit Turunc, Caroline Sui, Nick Morrison, Douglas Steiner

"In most copper mines and concentrators, mill water is continuously recovered and reused, with fresh water added to make up for losses. When mixed, these water chemistries can lead to severe scaling, especially in pipes, tanks and pumps on the mill water supply side, reducing flows and causing significant production losses.At KGHM Robinson Nevada Mining Company, increases in the fresh water addition ratios resulted in severe scaling problems and premature pump failures thereby jeopardizing plant production. What initially appeared to be a few isolated scaling events eventually unfolded into a recurring pattern of water supply-related production losses with increasing severity. Through completion of detailed root cause analysis, chemistry modeling, laboratory work, in-plant testing and system engineering, the mechanisms of scaling were identified and addressed.Advancements in fresh water addition practices and antiscalant application methods resulted in a stepwise reduction of the mill water scaling. Cost savings and production benefits were significant to the plant’s financial metrics.IntroductionThe production of mineral and metal concentrates using the flotation process requires large quantities of water. Flotation is utilized as the primary means for beneficiation with sulfide ores. Mined ores are crushed and ground into a slurry of 20-55% solids. For each ton of ore, 400-900 gallons of water1 are needed to supply the grinding mill. Additional water is needed in the flotation circuit as well. With many modern mills operating at2 11,000-165,000 tons/day, water supply volumes can be upwards of 3,000 gpm to 100,000 gpm of flow.For plants operating at or above capacity, disruptions and restrictions in water supply flow can have a direct impact on the production and recovery of the metal. Water supply flow can be limited for a wide variety of reasons but of interest are those limitations that are due to mineral scaling. Mineral scaling is the selective point precipitation of inorganic species above saturation concentrations. Precipitated species form flow-restricting deposits in the mill water system (tanks, pumps, pipe and valves), that will eventually impact water supply to the concentrator."

Jan 1, 2017

By C. J. Monahan

INTRODUCTION With the introduction of large diameter hole drilling to underground mining, a need for the development and utilization of new blasting technology has arisen. This new technology has resulted in more efficient blasting and in easier and cheaper mining. The application of spherical charges to the crater blasting method is an im¬portant part of this technology. Spherical charges, or their geometric equivalent, make for highly efficient use of explosives in the cratering application. Their geo¬metrical configuration, normally a length :diameter ra¬tio of <6:1, limits their charge weight size. For exam¬ple in a 165-mm (61/2-in.) hole a spherical charge weighs approximately 34 kg (75 lb). These dimensional and weight restrictions require careful engineering design and control in the production application. Strathcona mine, a large tonnage mechanized cut¬and-fill and blasthole stoping operation located on the north rim of the Sudbury Basin, has successfully em¬ployed the crater blasting method in production blasting for the past two and a half years (as of 1979). As a result, this method, using spherical charges, has made possible the development and implementation of a verti¬cal retreat blasthole mining method which is being used for rib pillar recovery. The crater method used at Strathcona essentially in¬volves blasting off horizontal slices of ore into an under¬cut while retreating vertically to the pillar overcut elevation. Loading and priming of the holes is achieved from the overcut and the blasted muck is recovered by load¬haul-dump (LHD) units through a drawpoint system. Valuable experience on the behavior and performance of explosives in large diameter holes has been gained in both the operating and engineering fields from the appli¬cation of this blasting method. It has been used in drop raising and is currently being evaluated for use in pri¬mary stoping operations. Also, serious consideration is being given to its use for the recovery of post pillars in mined-out and backfilled cut and fill stopes. The method has significant advantages over conven¬tional methods of pillar recovery from economic, ground control, and safety points of view. These include cost benefits due to a minimum stoping development, less ground support, reduced labor requirements, and faster mining rate. Research on the blasting mechanism indi¬cated the method should result in less damage to fill walls than would be normally expected with conven¬tional longhole or large diameter hole cylindrical blast¬ing (benching). This is primarily due to the limited charge weight per delay because of the size dimensions of a spherical as compared with a cylindrical charge and also because the large diameter holes experience less de¬viation than standard-sized longholes. Experience has shown that wall damage has been minimized with the use of this method. The safety aspects of the working environment are also improved as there is no require¬ment for manpower to go under exposed backs after blasting, which is a significant improvement over the cut-and-fill method. CRATERING THEORY General The term cratering, in blasting, is applied to the for¬mation of a surface cavity in a material as the result of detonating an explosive charge in that material. This blasting concept was initially used as a tool in the eval¬uation of explosives performance; however, it has been utilized more recently on surface and to a lesser extent in underground blasting operations. Explosive charges used in crater method are nor¬mally spherical or the geometric equivalent, as research into the application of this breakage mechanism to rock indicates that spherical charges or their equivalent pro¬duce optimum results. In blasting practice, spherical charges have been defined as having a length to diameter (L: D) ratio of 4:1 or less, and up to, but not exceeding a L: D = 6:1. Thus, for holes 165 mm (61/2 in.) diam, a charge 165 mm (61/2 in.) diam, and 990 mm (39 in.) in length would constitute a spherical charge. Crater Testing Crater testing for underground application is nor¬mally accomplished by drilling smaller than production sized horizontal holes into drift walls, placing relatively small charges of explosives in the holes at various depths of burial, firing the charges, and measuring the results. The results are scaled to develop charge weight size and depth of burial parameters for production hole sizes. In theory, this procedure is valid and is appli¬cable for surface blasting; however, in underground applications where the craters are inverted, there is some question as to whether extrapolation of the test results is appropriate. As well as the fact that craters are in¬verted with resultant gravity stresses, testing is often done in a waste rock drift where the properties differ from those in ore rock. There is also some question as to whether a 51 or 76-mm (2 or 3-in.) diam charge behaves similarly to a 165-mm (61/2-in.) charge in a confined borehole. Crater testing at Strathcona involved drilling 76-mm (3-in.) diam holes at various depths ranging from 1 to 3 m (3 to 10 ft). The test holes were drilled in the wall of a predominately waste drift as well as in a stope wall. Charges were placed at various depths of burial, were confined, fired, and the results measured. Although the results were somewhat erratic, possibly because of the aforementioned reasons, they did provide guidelines for basic design for production application. MINING METHOD APPLICATION FOR PILLAR RECOVERY Following initial crater testing, assessment, and tech¬nical evaluation, it was decided to use this blasting tech¬nique in a production application, and to incorporate

Jan 1, 1982

Since it was first introduced, the Coal Mine Roof Rating (CMRR) has been widely accepted as a tool for geologic characterization and mine planning. This paper discusses the application of the CMRR to another practical ground control problem. Extended cuts (cuts greater than 6 m (20 ft) in length) are commonly used with remote control continuous miners. Extended cuts can greatly increase productivity, but they have been associated with a number of fatal roof fall accidents. When extended cuts are attempted in weak roof, the roof may collapse before it can be bolted, causing hazardous conditions. Until now, it has not been possible to predict where conditions may not be suitable for extended cuts. In this study, data on the CMRR and extended cut experience were collected at 36 mines in 7 states. It was found that when the CMRR was greater than 55, deep cuts were routine in nearly every case. When the CMRR was less than 37, extended cuts were almost never taken. Between 38 and 55, extended cuts were feasible sometimes but not others. The data also shows that extended cuts are less likely to be stable if either the entry span or the depth of cover increases.

Jan 1, 1998

By Dean Treweek, Anthony Harding

"Creating stable, safe, constructible and cost effective support to openings in segmental linings is often an important component of successful delivery of a tunnel project. The many constraints on successful implementation include operational, spatial, geotechnical, structural, and constructability. Failure to address these adequately can lead to significant impacts on cost, schedule, and safety. With reference to past project experiences, this paper provides a guide to the most commonly deployed solutions and their pros and cons to help designers quickly identify the most appropriate solutions for a particular project, or act as a starting point for developing new and/or project specific solutions. It highlights the challenges that designers and contractors must address together for successful support of tunnel openings, and the constraints that arise. It then describes the common structural forms of support, and how each form can mitigate the constraints, allowing an effective opening support system to be implemented. INTRODUCTION While only a small portion of the total excavated volume of a typical tunnel, cross passages are often one of the most complex elements of a bored tunnel to construct. Cross passages in soft ground typically involve ground improvement, groundwater control, and multiple temporary support measures to ensure safe construction in ground that has been overstressed due to the construction of the running tunnels. Permanent lining construction requires careful detailing of waterproofing and reinforcement to support both long term ground and water loads and support the broken out rings of the main lining. While there is not a great deal of literature specific to cross passage construction, the main issues are discussed in some detail by Lawrence and Taylor (2011). However, the temporary support of the running tunnel segmental linings during cross passage construction is not always given the attention it deserves. The authors have been unable to find much literature on the design of such systems, much less the considerations that have led to a particular solution. Lawrence and Taylor only mention two potential methods of support. It may be easy to dismiss the opening support as an inconsequential item, but selection of the wrong support type can have significant impacts on both cost and schedule."

Jan 1, 2016

By Jaime Adasme, Héctor González, Benjamín Celada

"INTRODUCTION Santiago de Chile is a 7 million inhabitant city which has a 103 Km long subway network operated since 1975 by Metro S.A. This network is used on a daily basis by 2.3 million people. In 2012 Metro S.A. started the biggest enlargement of the network in its history, through the design and construction of the new Line 3 (21.7 Km) and Line 6 (16.5 Km). The detailed engineering contract for civil underground works of Line 6 (L6) was awarded to the Consortium integrated by ZAÑARTU Ingenieros Consultores and GEOCONTROL S.A., in which Geocontrol was in charge of the design of 11 stations and 14.5 Km of tunnel. L6 crosses L3 in Ñuble and Ñuñoa Station, L2 at Franklin Station and ends, intersecting L1, at Los Leones Station as shown in Figure 1. LOS LEONES STATION DESIGN Location and Characteristics Los Leones Station is placed at the North end of the L6, next to the same named station of the L1 and it has four construction units, as is shown in Figure 2: I. Access Shaft (49 m x 47 m). II. Access Gallery (34 m long). III. Station Tunnel (120 m long). IV. Connection Gallery with L1 (104 m long). A singular feature of Los Leones is that the western section of the station tunnel, built from the access gallery, is located below an underground parking lot which is located in Providencia Avenue, as shown in Figure 3. It is important to highlight that the distance between the vault of Los Leones Station Tunnel and the foundations of the underground parking lot is only 6 m."

Jan 1, 2016

By W. A. Haley

Wheel loaders became important bulk materials handling tools in the early fifties. As they grew in stature, productivity, reliability, and acceptance, the demand for larger, more rugged machines evolved. In the early sixties, the introduction of a 6-yard model by a major equipment manufacturer led quickly to widespread experimenting and rapid worldwide acceptance in loading blasted rock in stone quarries and smaller metal mines. As the wheel loaders grew in popularity for shot rock loading, larger models, 12- to 15-yard size were developed and saw widespread use by 1970. In the past 20 years, at most quarries and small to medium size mines, most cable operated power shovels have been replaced by large wheel loaders and more recently, hydraulic excavators as the primary loading tool. Wheel loaders larger than the 15-yard class are not used widely. Until about 1970, very few overburden haulage trucks were used in domestic coal mines. About that time, many states started enforcing stringent mine relcamation regulations and in 1977, even more stringent federal legislation was passed. These controls brought about drastic changes in coal surface mining methods in the mountainous areas of Appalachia, in smaller midwestern mines and in the large mines of the western states. Basically, simple material casting methods were replaced at most eastern mines by "haul-back" and "mountain top mining" methods using controlled placement of overburden as it is removed from the coal. These methods dictated much more extensive use of loaders/ shovels and trucks to the point that most coal mining in those areas is now by loader/shovel trucks methods. And those areas have become the home of the country&apos;s largest population of large wheel loaders and trucks in the 35- to 85-ton size class with a trend to even larger trucks and shovels. To maintain equal production or provide for growth, the new methods have caused a sharp increase in machine intensity to the extent that at some mines, up to three times as many machines were required to mine a million tons of coal in the late seventies as compared to the machine needs in the sixties. Mining costs and coal prices also took a big jump in the mid-seventies, paralleling the drastic price increases in oil and other forms of energy.

Jan 1, 1984

By J. J. Carlson

An important factor in filtration is the zeta potential of the particles. Since zeta potential is a function of pH, it was hypothesized that adding CO would lower the pH and zeta potential. A reduction in zeta potential should enhance the flocculation and increase the filtration rate. This paper describes the effects CO on the zeta potential of both pure hematite and samples of hematite concentrate from an iron ore processing plant. A significant shift in the zeta potential, as a function of pH, curve was observed. These results point to an alteration of the particles mineralogy, not just a shift in pH.

Jan 1, 2010

By D. L. Shenck

The Navajo Mine is exclusively a multiple seam surface coal mine located in the northwest corner of New Mexico, and entirely on the Navajo Indian Reservation (Figure 1). It is owned and operated by the Broken Hill Propriety Co., Ltd. of Melbourne, Australia. The mine began production in January 1963 with a long term fuel supply agreement with the Four Corners Power Plant. The 2175 MW nominal generating capacity of the Four Corners Power Plant consumes 8-8.5 mm tons of coal annually. With estimated reserves totaling 1.1 billion tons of strippable coal, only one third is presently committed to the generating station. As development has continued over the years, mining operations have advanced into more complex geologic areas to satisfy the utilities demand for coal.

Jan 1, 1990

By James D. Cooper

Long range projections of demand for nonmetallic minerals are presented, together with data on the domestic resources which are available or which can be made available to meet the requirements through the balance of the century. Past successes in solving nonmetallic mineral requirement problems are briefly reviewed to illustrate the capability for solving present and future problems of a similar nature. Major problems threatening the future supply of specific industrial minerals are described together with possible methods for assuring the necessary raw materials for future needs. The growing importance of environ- mental degradation, land use conflicts, and other important factors which affect several or all of the nonmetallic mineral industries is covered, and the necessity of planning for efficient and orderly extraction and rehabilitation of mineral deposits is stressed. The present Bureau of Mines methods and tools for long range planning are briefly discussed.

Jan 1, 1969

By Sidney V. Gold

Should Gold Producers Hedge Their Positions Through : Sales on Forward Markets? Purchases of Put Options? Sales on Spot Market? Forward Markets Assumptions 1. Producer has a known amount of Gold to sell in year ahead. 2. Effect on price is the same whether sold Forward or Spot Market. 3. None of us have knowledge of Future Price Trend. 4. Contange Market prevails in which Forward price premium is approximately equal to rate of interest.

Jan 1, 1983

By Duk-Wom Park

INTRODUCTION he most important information required for the engineering design of structures in rock is its inherent rock properties. Without providing adequate parameters on the rock property, high quality of the engineering design cannot be expected. Many laboratory rock testing methods have been developed, but they are mostly for testing intact homogeneous rock without appreciable discontinuities. In reality, the intact rock properties do not represent the behavior of rock masses because rock masses include irregularities such as joints, cracks, interlayers, variations in mineral compositjon, etc. In general the type and intensity of the rock defects may be much more important than the type of rock which will be encountered (Terzaghi, 1977; Bieniawski, 1984). In-situ rock testing can provide better information on rock parameters, but it takes a tremendous amount of effort, time and cost. In recent years, numerical modeling techniques are gaining popularity as a design tool for structures in rock bodies due to the enormous progress in computer technology (Wang, 1985; Kripakov and Melvin, 1983; Park and Ash, 1985), but it is quite dangerous to use numerical techniques without having proper input data. Bieniawski (1984) states as follows: Some design methods such as numerical techniques have outpaced our ability to provide the input data necessary for the application of these methods. In order to provide an adequate amount of input data, an extensive testing program may be required, thus major efforts in time and expenses are commonly spent, In this paper, a new rock testing method, which utilizes the principles of laser holographic interferometry, is introduced. This method requires only a few minutes to measure the modulus of elasticity and quality of a rock core sample, which does not have to be cut or ground. The Poisson&apos;s ratio can also be measured in a few minutes from a sliced rock specimen. HOLOGRAPHIC INTERFEROMETRY Principles of Holographic Interferometry In 1948, Gabor invented holography, which is the technique of reviving three dimensional images using a monochromatic light source. A full demonstration was not made then because a clean monochromatic light source was not available. Later, in the early 19601s, the laser was adopted as a monochromatic and coherent light source (Leith and Upatnieks, 1963). Since then, applications have been made in many different fields ranging from crime prevention to three dimensional television. In this technique, an image is usually recorded by means of constructing a hologram which is a record of the interference pattern formed on a high resolution photographic system as shown in Figure I. A beam splitter divides the laser beam into two beams: (1) an object beam which is expanded while passing through a beam expander and illuminates the object, and (2) a reference beam which is expanded and illuminates the holographic plate. Subsequent to an exposure, the holographic plate is developed and illuminated by the expanded reference beam. The original scene of the object is revived in a three-dimensional image. The entire instrument should be placed on a vibration free environment. The wave fields reflected by the object in two slightly different configurations can be superimposed to form an interference fringe pattern. Thus, it is possible to measure the displacement that took place between the two configurations. This method is called holographic interferometrv. There are two methods of achieving this phenomena. The first is the double exposure method, wherein two exposures of the object are made prior to developing the film. The first exposure represents the original condition and the second, the disturbed condition. The second method is the real time holographic method. In

Jan 1, 1986

By P. C. Miller

The paper reports laboratory results from the bioleaching of chalcopyrite ores and concentrates using a mixed moderate thermophile culture. Copper dissolution in excess of 95% was achieved in all cases with the majority of tests indicating copper dissolution in excess of 98%. Aspects of downstream processing are discussed in relation to iron rejection by precipitation. The laboratory results indicate the need for careful control of the pH value to prevent co-precipitation of the copper with iron. A comparison of capital and operating costs is given indicating potential differences between the use of toll smelting and bacterial leaching on site. A discussion is given highlighting the sensitive cost areas of these two process routes.

Jan 1, 1997

By Christina Dahnér, Kristina Jonsson, Savka Dineva

The iron-ore underground Kiirunavaara mine (Sweden) has been seismically active since 2008. In 2021 the production is on levels 1079 and 1108 m, at the deepest. The seismic activity has changed in character over time with one step change in 2015 when the number of large (ML>2) seismic events increased, and another one when the mine experienced a seismic event with Mw 4.2 in 2020. In 2017, the mine evaluated Hydraulic Fracturing (HF) as a possible mitigation technique for larger seismic events and a one-hole trial was performed. The trial proved promising, and it was decided to go forward with a large-scale trial covering an entire production area in the mine. The large-scale trial consists of 25 fracturing holes, 50 – 100 m long, running along 500 m out of the ~4500 m long orebody and spanning between production levels 1108 m and 1252 m. The fracturing process continued for a year, resulting in ~38,000 seismic events with −3.3<ML<1.4. The breakdown pressure varied between fractures in the same hole as well as between holes in the same area. Average breakdown pressure per hole was between 12 MPa and 50 MPa. The seismic response to HF in 2021 varied greatly from hole to hole and between fractures in the same hole. The variation in response could be related to the geological conditions and the infrastructure, such as large caving ore passes. In this paper some details about the experiments and broad preliminary analysis of the effect of the HF by using seismic source parameters of the induced seismic events are presented: 1) Distribution of seismic activity, 2) the ‘equivalent’ magnitudes for each hole calculated from the cumulative seismic energy and seismic potency released during the fracturing, compared with the local geological conditions; 3) the predominant failure mechanism of the seismic events caused by HF was defined. A difference in the behaviour between larger and smaller events was noticed, as ML>0 events seem to be compression type failures and ML<0 events seem to be related to slip on pre-existing structures. No seismic events related to creation of a new fracture have been identified yet.

Apr 26, 2022

By Ihsan Berk Tulu, Mustafa Can Suner

The synthetic rock mass (SRM) approach and discrete fracture network (DFN) model, generated from the field surveys, are novel methodologies to study fracture rock mass behavior. In this paper, SRM and DFN methodologies were used to understand the influence of natural fractures on the mechanical response and failure mechanisms of stone mine pillars. First, a multi-stage up-scaling procedure with the homogenization process is established. Later, a stochastic sampling process on the SRM model to have further insight on the stone mine pillar mechanics is performed. Two-dimensional Universal Distinct Element Code (UDEC) is utilized to represent intact rock with the Voronoi-Trigon discretized blocks and to capture fractured rock mass behavior. The laboratory-size limestone mine rock specimens are systematically up-scaled to the average width of the field-size stone mine pillars. Then, the DFN model, generated from field data, is sampled to construct the SRM of the field-size pillars. By doing so, the effect of discrete discontinuities on pillar stability is studied by assessing the pillar strengths and failure mechanisms in various width-to-height ratios.

Jul 8, 2022

By P. R. Thomas

This paper addresses the relative long-term cost and resource position of new and existing ferrochromium production in selected Market Economy Countries (MEC&apos;S). The market impact of new ferrochromium production in nations such as India and the Philippines and the continuing change in international trading patterns are also evaluated. This paper draws upon the data and analyses of a recent Bureau of Mines, Department of Interior, Minerals Availability Program appraisal that evaluated the cost and availability of chromite and ferrochromium production from the demonstrated resources of 80 producing or potential mining operations in 10 market economy countries.

Jan 1, 1984

The purpose of this talk is to report the answers to the-&apos;two following questions: I. What unique services does the mining industry expect of mining and mineral processing engineers? II. What suggestions does industry offer to educators to help revamp curricula to meet the needs of a changing industry? As the result of many questionnaires filled out in person by presidents, managers, superintendents, chief engineers, metallurgists, and other responsible individuals in operations and management of large and small metal and nonmetallic mining companies in the U. S., and to some extent abroad, we now have some, specific answers to the above questions. Although answers to questionnaires continue to arrive at this writing (February 10) the volume already received is sufficient to clearly justify certain conclusions and observations. A copy of the tally of the first 117 questionnaires received accompanies this report, and I am certain that a compilation of additional questionnaire data will merely reinforce, the trends in thinking that have been&apos; established thus far. Here are the results of the survey:

Jan 1, 1961