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By Pentti Seppanen

The activities of Outokumpu Oy in the mining industry span all branches from prospecting to the production of refined metals and the manufacture of semifinished products. Products include copper, zinc, mercury, cadmium, nickel, cobalt, sulfur, gold, silver, iron ore, stainless steel, selenium, ferrochromium, lead and chromium concentrates, as well as chromite foundry sand. In addition, the activities of Outokumpu Oy extend to development and production of instruments, planning of industrial projects and instrumentation as well as the manufacture of specialized machines for ore concentration. The company now operates ten mines, eight of which are underground and all of which are in sulfide deposits except one. The fully automated monitoring and control of underground mine ventilation at the Pyhasalmi mine are an important technological contribution to the industry. Development work on the ventilation control was started in 1969, and since May 1973 it has operated in a fully automated computer-controlled mode. The Pyhasalmi mine has been in production since 1962, having begun as an open-pit operation. Development of underground mining began in 1967. The present annual production is one million tons all from underground mining. The underground mine is equipped with a main shaft and an auxiliary shaft, and two raises for ventilation. There is also a rampway from the surface to a depth of 500 m. The main shaft is for ore hoisting and the auxiliary shaft is used for waste-rock hoisting and for downcast ventilation. One of the raises is for downcast ventilation and the other is for upcast ventilation. The mining methods employed at Pyhasalmi are sublevel stoping with backfill, sublevel caving, and sublevel shrinkage stoping. Copper and zinc concentrates are produced. The Pyhasalmi mine is highly mechanized with about 4000 installed diesel horsepower underground. The maximum capacity of fresh air is 150 m3/s.

Jan 1, 1979

By W. S. Versluis, G. W. R. Kuzyk

Atomic Energy of Canada, Ltd. (AECL) is constructing an Underground Research Laboratory (URL), a major component of-the Canadian Nuclear Fuel Waste Management Program. In July 1988. AECL and the US Department of Energy commenced a cooperative research program that included extending the URL shaft by 190 m (623 ft) to the 443-m (1,453 ft) depth. The shaft extension, 4.6-m- diam (15-ft-diam), was excavated by full-face blasting using a burn cut. A Galloway stage accommodated the research activities and the sinking contractor's requirements. Blast designs incorporated controlled blasting techniques to minimize blast-induced damage on the shaft walls. The full-face method has been very effective, with 3.5-m (11.5-ft) rounds being advanced.

Jan 1, 1990

By R. M. Clowes

LITHOPROBE is a collaborative, multidisciplinary research project that is providing new insights into the tectonic evolution of the North American continent through its work within 10 study areas across Canada. For the mining industry, such results provide a well-constrained regional framework and better understanding of the tectonic processes that lead to the development of ore deposits. In mining camps, seismic-reflection methods have proven effective in imaging shallow to moderately dipping lithologic contacts as well as structural features and, in some cases, have been used for the direct detection of orebodies. Here, short case histories are provided for the Abitibi greenstone belt (VMS deposits) in Quebec and the Thompson nickel belt in Manitoba.

Jan 1, 2002

By Jack W. Eloranta, Michael J. T. Orobona, Peter R. VanDelinder

Ore blending for process optimization is becoming an increasing factor for managing cost and productivity. Economic mining and processing a very low value ore in an extremely competitive market results in constant cost-cummng initiatives. The inherent conflict between the lowest possible mining cost and the lowest possible milling cost is reviewed here in the context of blasting for autogenous mills. Iron ore blending focuses on a number of properties-liberation values, weight recovery and concentrate silica being the most important. Other chemical and metallurgical factors are also commonly controlled. Recent research has fueled an emerging interest in blending on physical ore properties as well. Fully autogenous mills rely on the large, competent fraction of crushed feed to act as grinding media. Previous test work indicated that Hibtac mills require specific amounts of 6- to 10-inch ore. High recirculating loads require a steady influx of large rock fragments. Blast designs, therefore, are presently characterized by wide patterns and low powder factors. Current efforts are aimed at improving mill throughput while reducing overall energy costs. Based on research and published case studies and the advent of optical measuring devices, Hibbing Taconite Company has embarked on a blast optimization program to identify optimum mill feed and to design blast fragmentation goals for each geological unit and each mining area. This paper outlines the preliminary findings of an ongoing, broad-based team effort which requires close cooperation of geologists, mine engineers, crushing and milling personnel. Three specific areas of investigation are reviewed: I) Historical relationships between powder factor and mill performance, 2) Blast fragmentation modeling using the Kuz-Ram model, and 3) Drill

Jan 1, 2004

By G. Fehr, B. B. Panda, R. Bansberg, K. Arnold, T. Freiman

To evaluate the potential for process solutions to contaminate groundwater outside of the area of the Oxide Pit, it is necessary to predict the ultimate maximum pit lake elevation once pit-related mining activities have ceased. A water balance model was developed based on the existing pit geometry, the groundwater flow equation, and assumed hydrologic input parameters. The model assumed that at any point in time, total inflow into the system (i.e., direct precipitation, surface water and groundwater) minus outflow from the system (evaporation from the lake surface) was equal to the change in the pit water volume. The pit lake depth over time was predicted by assuming a succession of steady- state water balance conditions over short time increments. An analytical method using a spreadsheet and computer model was employed in this study to model the filling of the lake in the Oxide Pit. A sensitivity analysis was performed to evaluate uncertainty associated with the input parameters in the model. The sensitivity analysis indicated that the model was most sensitive to values of precipitation and evaporation and to the static water table.

Jan 1, 2003

By J. O. Marsden

The 1980s saw a revolution in gold extraction technology, paralleled only by the implementation of the cyanide leaching process in the late 1800s. This revolution was led by the widespread adoption of carbon adsorption processes for gold recovery, heap leaching of low-grade ores and development of a variety of oxidation processes to treat refractory ores. In the late 1980s, gold production in almost all regions of the world increased, with the notable exception of South Africa. This geographical variety is creating a greater diversity of challenges for gold metallurgists. This is due to a wider range of ore types, new and varied environmental regulations and legislation, and a requirement to select the most appropriate technology to suit local needs. North America has led this technological revolution in the gold industry and is responsible for commercialization of most of the major innovations: carbon-in-pulp, heap leaching, whole ore pressure oxidation and whole ore roasting. Process selection trends A general trend in precious metals extraction technology has been the treatment of ores with progressively lower grades. The lowest cost treatment, dump leaching of material at run-of-mine size, has been increasingly applied. ZortmanLandusky, in Montana, is successfully leaching material above a cutoff grade of 0.2 g/t (0.006 oz/st) at an average rate of about 28 kt/d (31,000 stpd). At this scale of operation and in the absence of any cyanide-consuming species in the ore, process operating costs for a ge¬neric run-of-mine leaching operation are estimated to be between 60 and 75 cents/ t (66 and 77 cents/st). One of the largest heap leach opera¬tions treating crushed ore is the Round Mountain Mine in Nevada. It processes about 45 kt/d (49,000 stpd) at an average grade of 1.1 g/t (0.032 oz/st) at a nomi¬nal crush size of 80% -19 mm (-0.7 in.). Round Mountain also treats lower grade ore above a 0.2 g/t (0.006 oz/st) cutoff by dump leaching at run-of-mine size.

Jan 1, 1993

By Robert S. Ondrey, Robert A. Haney, Kenneth G. Fields

As part of an MSHA program to assess the effectiveness of dust control parameters for longwall mining operations, respirable dust studies were conducted on six longwall sections. The purpose of these studies was to evaluate the influence of airflow and production on longwall dust control. The sections were sampled for 3 to 35 shifts. In addition to dust, ventilation and production measurements, other dust control and operational parameters were measured. The water spray system at each operation conformed with generally recommended practices. Therefore, improvements to the dust control system would be most appropriately achieved through improvements to ventilation along the face. Two methods of assessing the influence of ventilation and production were used. For two of the six mines where extensive sampling (22 and 35 shifts) was conducted, multiple linear regressions were used to relate the dust concentration versus airflow and production data. For these two mines, statistically significant correlations existed between concentration and both production and airflow. The multiple regressions fit to the data indicated that for every 0.47 m3/s (1,000 cfm) increase in airflow, there would be approximately 0.04 to 0.07 mg/m3 reduction in dust concentration; and, for every 1,000 ton increase in production, there would be approximately 0.15 to 0.50 mg/m3 increase in dust concentration. For the four mines with limited data (3 to 10 shifts), as well as for the two mines with extensive sampling, an average ratio of airflow per ton of coal mined was calculated by proportioning the airflow to production and the designated occupation personal respirable dust exposure to a 2.0 mg/m3 concentration. The airflow-to-tons ratio required to maintain a 2.0 mg/m3 dust concentration ranged from 0.0038 0.0071 m3/s per ton (8 to 15 cfm/ton). Both of these data treatments were used to calculate airflow requirements based on normal production rates. For the two mines where both the regression and airflow-to-tons ratio were calculated, results from the two methods agreed within 25 percent.

Jan 1, 1993

By E. C. Speers

The Ray Mines Division has a long and illustrious history beginning in 1911, as one of the early underground block caving copper mining operations. The tonnage mined from the beginning of operations, increased from about 8, 000 TPD in the early years to the present daily tonnage of 22, 500 tons, mined by open pit methods. The Ray basic milling practice has been relatively unchanged since 1927 when the sulphide-flotation process was introduced at Ray. The leach-precipitation-flotation process introduced in 1956 has been the only major process change. The change from underground to open pit' mining was accompanied by a renewed drilling and exploration program which developed large tonnages of low grade copper ores both in the porphyry and the diabase. The ore bodies as outlined were of such a grade and size that it soon became apparent that economical operations could not be developed with the lower tonnage rate. Also that in order to develop these newly proven ores, it would be necessary to move the Ray crushing plant, the Ray. supporting shops, offices, a substantial part of the Ray townsite, and would also involve the rerouting of Mineral Creek, a typical Arizona stream which is dry most of the year but can flow as high, as 8, 000 second feet in the rainy season. This paper describes those changes in flow sheet made necessary both by the increased tonnage .and the requirement to move as much of the crushing facilities from the Ray area as practical in order to free the mine for the extension of its physical limits.

Jan 1, 1961

By Peter J. Szabo

Introduction There has been much said recently about the sophistication of Japanese management in turning out reliable, inexpensive products, such as cars and steel. Techniques include consensus management and quality control circles. These have been touted as prime reasons for inexpensive Japanese products. An often neglected, but major factor in the equation of cheap export goods is the role played by Japanese trading companies. These venerable institutions are a vital factor in Japan's overall strategy of keeping the price of raw materials at the absolute lowest levels. This is critical to the country's economy since the nation must import more than 85% of all its coal. A country that is so desperate for imported raw materials would appear, at first glance, to be at the mercy of its suppliers. In Japan's case, however, the opposite is true. With the exception of brief' windows in 1974-75 and 1980-81, Japan has been in the driver's seat in negotiating contract terms. Due to its dominant role in world coal trade, Japan's situation can be described as a de facto monopsony. By definition, a monopsony is a market situation where the product, in this case coal, is offered by many sellers, but is sought by one buyer. This has been due to the role of Japanese trading companies in dealing with suppliers. Table I depicts trading companies by size in 1982. Mitsubishi and Mitsui are overshadowing in terms of size. The others, though less significant, play a vital role in orchestrating the marketplace. Three critical building blocks on which the strategy of the trading companies are based include intelligence networks, exceptional training of company traders, and shrewd contract negotiation practices. Intelligence Networks All success is derived from these three key components. Of the three blocks, intelligence gathering is the most important. With all the sophistication of the Central Intelligence Agency, the Federal Bureau of Investigation, Interpol, and the Surete', the Japanese trading company's method is every bit as competent and far more gentlemanly. Friendly competition is the key to survival. Although one Japanese trading company appears to be in ardent competition with its rival Japanese trading companies, when studied closely, a totally different picture emerges. The end user nominates certain trading companies as market channels and informally guarantees they will receive compensation regardless of which company wins a specific order. With this concept in mind, the pressure is off. The trading company can aim at getting the best price for the end user. Therefore, bargain, beat, and share information. Do whatever is necessary, but get the lowest price for the end user. Even with the game so gentlemanly, no victory could be assured without an adequate intelligence network. Proper intelligence is only as good as the sources drawn from. But how is one assured of getting the right information? The answer is verification. Verification is comprised of numerous independent sources each cross checked at a central processing unit in Tokyo. Each major Japanese trading company receives input, not only from its own trader, but from other trading companies as well. In addition to their own compilation, trading companies rely on the The Coal Manual, published annually in Tokyo by TEX. This manual covers every major coal producing country. Mines are located and seams identified, mining recovery, equipment lists, washing yield, railway distances

Jan 1, 1985

By John J. Stetler

Introduction Samarco Mineracao S.A. is a Brazilian corporation jointly owned by S.A. Mineracao da Trindade (SAMITRI) and Utah International Inc., with 51% and 49% interests respectively. Headquartered in Belo Horizonte in the state of Minas Gerais, Samarco owns and operates an open-pit mine capable of producing 10 Mt/a (11 million stpy) of iron ore; a concentrator rated at 7 Mt (7.7 million st); a 12-Mt (13.2-million st) capacity pipeline transportation system; a 7 Mt (7.7 million st) filtering and pelletizing plant; and a harbor facility. The mine, concentrator, and initial pipeline pump station are located at Germano, near the town of Mariana, about 300 km (186 mi) north of Rio de Janeiro by air. The pipeline terminal, pellet plant, and harbor are at Point Ubu, 70 km (43 mi) south of Vitória the capital city of Espirito Santo State, and about 300 km (186 mi) by air east of the mine. The 396-km (246-mi) pipeline uses 508-mm (20-in.) pipe except for a 53-km (33-mi) section where 457-mm (18-in.) pipe is used to absorb and control slack flow through increased velocity in one of the areas along the route. Iron Ore Slurry The hematite iron ore concentrate produced at the Germano plant is processed by grinding the average 53% run-of-mine itabirite iron ore to -150 µm (100 mesh) for gangue liberation. It is treated further by cyclones to remove the 10 µm (1250 mesh) slimes to prepare the ore for flotation. During the multistage cationic flotation operation, the silica is reduced to make low silica (-0.9%) concentrate for the direct reduction (DR) iron-making process or to make normal silica (-3.5% +2.5%) concentrate for the blast furnace (BF) iron-making process. The concentrate then is reground further to meet the particle size requirements for pelletizing. The reground hematite concentrate is screened in closed circuit with the regrind mills for close control of the maximum particle size, not to exceed 3.5% +75 µm (200 mesh) for pipeline wear protection. Thickened concentrates at 70% solids are pumped to four 5-kt (5,510-st)(dry) agitated slurry storage tanks that hold the feed before it goes to the first pipeline system pumping station. Pipeline Capacity The pipeline system was designed to transport 12 Mt/a (13.2 million stpy) at 65% solids with an average minimum safe operating velocity of 1.5 m/s (5 fps). The pipeline will transport 5 Mt (5.5 million st) in 1982. The system, therefore, is moving the slurry in 13-kt (14,330-st) batches, dispatched at intervals to move 20 kt/d (22,050 stpd). In operation, the pipeline has exceeded its 93% availability design factor. Pipeline Facilities Pump station No. 1 is located at the Germano minesite 1 km (3,330 ft) above sea level. It has the following principal equipment: • four 5 kt (5,510-st) (dry) agitate slurry storage tanks; • two centrifugal charge pumps (one stand-by); • a 217-m (712-ft) safety loop; • seven mainline positive displacement triplex plunger pumps (189 m3/h, or 832 gpm, at 14.9 MPa, or 2,162 psi, maximum output pressure); • four corrosion-erosion spools; • a control and command center. Operation The output at pump station No. 1, with 66% solids, is pumped 151

Jan 10, 1982

By Ava Walker

The first form of a surety bond was an honest handshake between two or more parties agreeing to a decision, and their personal guarantee of following through to completion. This has all changed. In 1895, the United States Treasury Department disclosed that $35 million worth of personal surety guarantees on file had become worthless. A law was then passed requiring all surety bonds, in favor of the United States, must be corporate surety bonds. Today, a surety bond is one of the nation's leading fi¬nancial instruments used to guarantee the completion of an obligation. One hundred-fifty years before cor¬porate surety bonds, outcrops near Richmond, VA, were being mined. These surface mines became under¬ground mines. Today, both are still being used to obtain a leading energy source, coal. Since that time, more than 5956 km2 (1.47 million acres) of coal mined lands have been reclaimed in the US (Coal News, August, 1987). More than 25,000 state and Federal mining permits cover¬ing 14,164 km2 (3.5 million acres) have been issued during the last 10 years. Coal production has exceeded 7.2 Gt (8 billion st). More than 5665 km2 (1.4 million acres) of mined land has been reclaimed by active coal mining opera¬tions and 267 km2 (66,000 acres) of abandoned mined lands have been re¬claimed by others. Reclamation bonds were freely writ¬ten in the 1970s. So why are they almost extinct today? Contract bond losses have been severe in the past several years, sureties reporting losses exceed¬ing 110%. But the same proof is not provided on coal reclamation bonds. A surety bond is generally required when one party has an agreement with another party, and that party wants a guarantee from a third party that a job will be completed according to the terms and conditions of a written contract. In the case of a reclamation bond, the surety company (surety) is the financial party. It extends the credit and a guarantee to the obligee (the State, Commonwealth, or Office of Surface Mining) that the principal (the coal operator/company) will complete the terms and conditions of the contract (the rules and regulations governing mining and reclamation). The Surety Association of America defines a surety bond as an agreement providing for monetary compensation should there be a failure to perform specified acts within a stated period. Webster's Dictionary defines surety in the following ways: "l: the state of being sure; 2a: a pledge or other formal engagement given for the fulfillment of an undertaking and; 3b: one who has become legally liable for the debt, de¬fault, or failure in duty of another." Banks never make accommodation loans. Either the loan is good or no loan is made. Surety companies feel the same way about surety bonds. In fact, the average person does not realize the dangers of becoming surety for another. How does a surety determine if a company is worthy of receiving surety credit? How is a coal account under¬written today for an obligation that may not be released for many years? When a surety company executes a reclama¬tion bond, it financially puts its name behind the coal company and its guaran¬tee that should the coal miner default on the obligation, the surety will fulfill the contract. How is the surety to determine if a coal company may still be in business five, 10, 20, or 40 years from today?

Jan 1, 1989

By D. M. Range

Exxon is reclaiming an 80-hm2 (200¬acre) Wyoming uranium tailings basin in accordance with US Nuclear Regula¬tory Commission (NRC) requirements. The reclamation plan is designed to control the tailings and their radioactive emissions for at least 1000 years. Rigorous modeling and analyses of potential erosion by wind and water were used in planning. The use of riprap was minimized and the need for long¬term reclamation maintenance was eliminated. The resulting design is simi¬lar to regions stable from erosion cre¬ated by nature. The concepts used in the design and implementation have wider reclamation application. Highland history Exxon operated the Highland uranium operations from 1972 until mid-1984. Open pit and underground operations mined the roll front, sandstone ore body. The 2.5 kt/d (3000 stpd) mill employed conventional crushing, grinding, sulfu¬ric acid leach, CCD, SX, precipitation, and drying circuits. The mill produced 10 Mt (11 million st) of tailings. An¬other operator began commercial scale in situ leach operations in 1988. The tailings basin was constructed by deposition of mill tailings in a natural ravine dammed with a 30-m (100-ft) high by 1050-m (3400-ft) long clay core earthen dam built entirely downstream of tailings. Figure 1 shows the layout of the tail¬ings basin in 1984. Tailings were de¬posited along the dam face on the east side of the basin as well as in the west end and along the south edge until the early 1980s. This left a deep water pond in the middle of the basin. The deep pond was largely filled with tailings during the last four years of mill opera¬tions by the construction of two tailings causeways and the operation of a float¬ing tailings line. The pond water volume was mini¬mized by recycling tailings water to the mill and by operating a 32-L/s (500¬gpm) evaporation enhancement spray system during the last few years of mill operations. Regulatory requirements Several conceptual tailings basin rec¬lamation plans were prepared during the mine life. In September 1984, Exxon began preparation of a detailed plan in response to an NRC request. Water, Waste and Land Inc. (WWL) of Fort Collins, CO was selected to help pre¬pare the plan that was submitted to the NRC in December 1984. The design had to meet the require¬ments found in NRC regulations (NRC, 1985). These were based on regulations prepared by the Environmental Protec¬tion Agency (EPA, 1983) at the man¬date of federal law (UMTRA, 1978). The NRC regulations contain strict primary standards and a somewhat more flexible set of secondary standards. The primary standards include requiring that: • radiation from tailings be reduced to background levels; • radon gas emitted by the tailings be reduced to 20 pCi/m2/s from the surface of the reclaimed tailings pile; • tailings not be dispersed into the environment; and • these conditions endure without routine maintenance for at least 200 years and for 1000 years to the extent reasonably attainable. The lengthy time requirements pre¬sented a large engineering challenge. Radon is a radioactive, chemically inert gas with a 3.8 day half-life. It is released from uranium tailings by the radioactive decay of radium, which, in turn, is a radioactive decay product of uranium. Radon emissions from ura¬nium tailings can be controlled with an earthen cover thick enough that the radon gas decays to its solid, short-lived radio¬active daughters before the radon dif¬fuses through the cover material into the air.

Jan 1, 1989

By B. G. McNish

Neutron Activator Laboratory Analyzer used for Rapid Analysis for Ore Blending, Stripping, and Mine Control BAUXITE is a rock and not a mineral. Since it is a rock, its mineral content varies depending upon the parent material, the topographic location of the deposit and the severity of the weathering and oxidation processes which produced it. At Bauxite, Arkansas the major mineralogical constituent of the bauxite is gibbsite with ~IL- purities of silica, iron oxide, and titania minerals. Minor amounts of other minerals are found. Chemical composition varies from 45% to 60% alumina, 5% to 25% silica, 1% to 15% ferric oxide, and 1% to 3% titania. The Arkansas bauxite region is located entirely within the north- central area of the West Gulf Coastal Plain near the northwestern boundary with the Interior Highlands. The region trends northeast-southwest and is about 42 kilometers long. It is divided into two districts; one in Saline County and the other in Pulaski County. The Saline District is the most important since it has produced most of the bauxite. This district is located at Bauxite, Arkansas and the deposits are found around the flanks of the Saline Dome. It is roughly square-shaped and is about 190 square kilometers in area. The source rock for the deposits is a Late Cretaceous, grey, nepheline syenite.

Jan 1, 1979

By H. Colijn

Weight may be defined as the gravitational attraction or earthpull on a body (mass). Thus, from a practical viewpoint, a scale or other form of weighing device is a mans for determining mss. Weighing devices fall into two principal categories: (1) gravimetric; and (2) non-gravimetric or nuclear devices. At present, after many centuries of use, the effect of gravitational force as a meas¬ure of mass in industry and commerce remains by far the most common method for mass (weight) determination. Non-gravimetric methods are just becoming established in the industry. The general trend is toward weighing materials on the Trove. The predominant advantage of continous weighing over batch weighing operations is its ability to handle large tonnages without interrupting the material flow. The accuracy and reliability ability of continuous weighing equipment have improved greatly over the past decade. However, static weighing equipment still outsells continuous weighing equipment, because it is considered more accurate. Batch weighing can also produce a continuous weighed delivery through automation where the process requires it. In caparison with continuous weighing devices, such as belt scales, the batch-type scales have same inherent advantages affecting their accuracy. For example, weighing 300 lbs. per minute on a belt scale may mean that only 3 lbs. will be on the scale mechanism at one time. In order to obtain a scale accuracy of 1/2%, the sensitivity must be at least 1/2% of 3 lbs. or 1/4 oz. On the other hand, a batch scale may discharge a 300 lbs. load every 60 seconds through a timing mechanism. An accuracy of 1/2% can be easily obtained with a sensitivity of 1/2% of 300 lbs. or 24-oz.

Jan 1, 1970

Many investigators believe that column flotation cells offer significant advantages over standard mechanical machines for the flotation of fine particles. However, because of their unique design and operation, conventional techniques for flotation cell scale-up and design cannot be applied to columns. In an attempt to help alleviate this problems a population balance model based on first principles has been developed for fine particle flotation in a column. Two different terms have been considered in the model, i.e., transport and rate. Transport terms, incorporating fluid flow and bouyancy, are used to describe the movement of air bubbles, unattached particles and bubble-particle aggregates along the length of the column. Rate terms, which describe the bubble-particle attachment process, have been derived from first principle considerations. Because the model is based on first principles, it can be useful for the design, control, optimization and scale-up of column flotation cells.

Jan 1, 1987

By B. Simser, T. Butler

Glencore’s Nickel Rim South Mine, located in Sudbury, Ontario, Canada, is a primary/secondary blasthole operation with ore zones lying between 1,100 and 1,700 m below surface. One of the main benefits of the primary/secondary mining method in high stress ground is that up to 50% of the mining will take place in secondary stopes that should be in a yielded rock mass state. Every future secondary mining stope begins as a primary stope abutment pillar. As with many mines, the stress loading history and rock mass conditions can vary considerably in space and time. Safe mining practices cannot allow for an assumption that the planned secondary stopes are indeed yielded. An incorrect assumption can result in mining development and stoping taking place in high stress ground rather than the expected yielded rock mass conditions. This can result in higher levels of seismicity than expected and incorrect ground support recommendations in areas where workers could be exposed. Seismic monitoring is the mines’ premier tool to track the actual rock mass response to mining. Yielded zones have a reasonably clear seismic signature/pattern over time that is visible in the collected data. The ability to accurately locate small magnitude seismic events can help define and delineate different stages of rock mass failure. The vast majority of the seismic data recorded by the microseismic monitoring system is below moment magnitude (Mw) 0.0. Examples are provided showing how small seismic events (Mw < -1.5) can help provide feedback from the rock mass deformation that allows the ground control engineer to gain a better understanding of how the rock mass is responding to mining processes.

Apr 26, 2022

By M. Plesha, A. Zubelewicz, K. Connor, T. B. Belytschko, O&apos

INTRODUCTION The Northwestern University Rigid Block Model (NURBM) has been extended to explicitly model dilatant behavior of rock joints. By incorporating dilatancy, the resistance to shear displacement and slip, which previously was limited to simple Coulomb friction, along edge-to- edge contacts is now provided by contact roughness as well as frictional resistance. The explicit formulation is based on concepts presented by Plesha (1985) in which a plasticity-based phenomenological approach was used to model dilatant joint behavior. This approach has been modified to allow for joint closure upon reversal of shear deformation. After presenting a brief description of the new joint constitutive relationship, model behavior will be illustrated with two examples. The first involves a simple two-block model to simulate a cyclic direct shear test of a single joint. The second example involves simulation of a large cavern in a moderately jointed rock mass subjected to cyclic excitation. It will be shown that in- creasing joint roughness increases cavern stability by interlocking of adjacent blocks. JOINT CONSTITUTIVE RELATIONSHIP Within NURBM adjacent two-dimensional rigid blocks interact along their common edge-to-edge contact with contact forces transmitted at the contact midpoint. The contact constitutive relationship was originally restricted to a Coulomb friction yield criterion with no tensile stresses permitted along the smooth contact surfaces. The local coordinate system for any contact is described as shown in Figure 1 with the elastic stresses and displacements related by the linear- elastic constitutive law: [ ] The contact microstructure has been modeled as Patton-type saw-tooth surfaces such as shown in Figure 2. Sliding can be active along the right asperity with angle aR and length lR or along the left asperity with angle a and length lL. In general, aR and aL will difkr as will lR and lL but for purposes of this paper it has been assumed that these angles and lengths are identical. By a simple transformation, the microscopic asperity stresses can be expressed in terms of the macroscopic contact stresses [ ] If Coulomb friction is assumed along the asperity surfaces, the plastic slip condition will be [ ] where µ is the friction coefficient with the convention that compressive stress is negative. When F is negative, the behavior is elastic and when F is zero, slip can ensure. When plastic slipping does occur, it is assumed that the relative tangential and normal displacements between the two contacting rigid blocks are additively composed of recoverable elastic parts and nonrecoverale plastic parts, [ ]

Jan 1, 1986

By A. Ersoy, M. D. Waller

Because of a shortage of diamonds for the drilling industry in the 1970s, efforts were made to develop synthetic diamonds. The resulting development of polycrystalline diamond compacts (PDC) enabled the evolution of a new generation of rotary-drilling drag bits. Bits based on PDC cutters were very successful in drilling soft, abrasive and medium strength rocks. However, because of their thermal instability, they failed rapidly when drilling hard rocks. Thermally stable diamond (TSD) was then developed for hard-rock drilling. The rate of penetration (ROP) and bit life significantly influence the drilling cost. The ROP and the life of the TSD bits or cutters are affected by a number of complex factors. However, in laboratory testing and infield operations, the principal factors that need to be considered in predicting drilling and wear rates are bit design characteristics, bit operating parameters and the characteristics of the penetrated rock. These characteristics are comprehensively reviewed for both laboratory and field drilling results using TSD bits. Optimization of these parameters increases the drilling rate, prolongs the bit life and reduces the drilling cost.

Jan 1, 1998

By Armando M. Lago

This paper reviews the postwar changes that have occurred in the expenditure and financial behavior of mineral corporations, and in particular the relative importance of factors that affect the industry&apos;s supply and demand for funds. This quantitative statistical investigation of factors affecting sources and uses of funds of mineral corporations is achieved through the formulation of a simultaneous equation model of sources and uses of funds. After a description of the study sample in Section I, the postwar cyclical and secular financial behavior of mineral corporations is analyzed in detail in Section II. Section III formulates an econometric model of mineral corporate financial behavior which is then estimated and reviewed in Section IV. Section V discusses policy issues regarding depletion allowances, investment tax credits, and industrial organization of the mineral sector. Finally, the optimum financial structure of mining firms is discussed in Section VI, which analyzes the effect of corporate financial structure on the mining corporations&apos; average Cost of capital.

Jan 1, 1969

By D. C. Torre

The need for a successful, efficient and reliable continuous face- haulage .system for use with continuous mining machines has been apparent since the first continuous miners were introduced some twenty-five years ago. Some thought and effort were given to the development of continuous haulage over the last twenty years or so. Some initial success was experienced in this area by The Long Company with a system involving bridge conveyors, as far back as the late 1950&apos;s. However, extensive success with continuous face haulage for use behind modern, high product¬ion, continuous miners escaped the industry until the early 1970&apos;s. It is true that continuous face haulage was successfully introduced into low seam mining, generally in seams ranging under 1.07 m (42") in height, in the 1960&apos;s, for use with relatively low production, low seam, auger type continuous miners. However, those available systems which have achieved success in the lower seams, where relatively low capacities are adequate, are not directly applicable for use with today&apos;s medium and high seam miners with their standard production capacity of 9.1 - 10.9 mtpm (10 - 12 tpm). Since the late 1960&apos;s and early 1970&apos;s, several machinery manufacturers, and at least two major coal companies, have devoted considerable expense, time and effort to developing truly successful continuous face haulage for use in coal seams ranging upward from 1.2 m (4&apos;) in height. Continuous face haulage equipment has been developed in the last five years employing a variety of transportation media and design concepts, including some which have received wide publicity such as those employing hydraulics, and pleated belt. The truth of the matter is, however, that the only universally applicable and widely successful continuous face haulage equipment introduced to date has been that employing a combination of bridge conveyors and mobile, self-propelled, bridge carriers, connected at the inby end to the tail of the high production continuous miner, and at the outby end attached directly to the panel conveyor (Fig. 1 & Fig. 2) or, on occasion, to a short, self-contained conveyor which, in turn, empties directly onto the panel conveyor. While these bridge type continuous haulage systems are available with both belt and chain conveyor components, the most widely applied and successful system in seams above 1.07 m (42") in height is that offered by The Long-Airdox Company. The standard unit for use with 9.1 - 10.9 mtpm (10 - 12 tpm) continuous miners is normally equipped with belt. However, they have built, and successfully applied, versions of this system employing chain conveyors, (Fig. 3), mainly for operators having extremely difficult mining conditions, including excessive amounts of water

Jan 1, 1977