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By Jamal Rostami

With the recent technological advances; Tunnel Boring Machines (TBM) have become a viable economic alternative to drill and blast in hard rock excavation. These machines are finding increasing use in underground mine development due to the potential benefits they offer. The successful application of TBM technology to mining depends on the selection of the most suitable equipment and cutting tools for the rock and ground conditions to be encountered. In addition to required testing and geotechnical investigations, cutterhead design optimization is an integral part of the machine selection to assure a successful application of the machines in a specific project. This paper presents and discusses the back-ground of TBM application in a mining environment and the lessons learned, the process of conducting site investigation and the necessary laboratory testing together with machine selection and performance estimation for TBM use III mine development.

Jan 1, 1997

By J. L. Workman

The stability of spoil and highwall slopes are important to successful strip mining. This paper discusses soil parameters of importance in stability analysis. The common mechanisms of spoil failure are discussed. Modes of highwall failure are also outlined. The importance of water to stability is presented. Dragline mining methods most suited to creating spoil and highwall profiles that address stability needs are provided. It is concluded that groundwater and surface runoff must be well controlled, and that it will often be necessary to modify spoil and highwall configurations to achieve a stable overall slope angle.

Jan 1, 1996

By T. A. E. Breen

A new high intensity grinding technology has been developed and is now being commercialized by Hicom International of Australia. This has led to the development and manufacture of Hicom 120 grinding mills which are now being sold internationally for a range of mining industry applications. The first commercial process application has been for the diamond sector, where the ability of the mill to remove host kimberlite rock by attritioning, with no damage to the thereby liberated diamonds, is seen to be a most beneficial attribute. There are, however, many other mineral industry applications for this mill.

Jan 1, 1999

By B. Denby

Recent advances in Virtual Reality (VR) technology have resulted in major opportunities to improve safety training of mine workers, supervisors and managers. The hazardous nature of both surface and underground environments limits the type of safety training that can be carried out. VR offers the potential to expose mine personnel to hazardous situations in a safe, highly visual and interactive manner. Customized simulations of mine layouts, dynamic operations and mine environment can be set up and mine personnel can be allowed to move around the virtual mine, making decisions at key points. The consequences of both correct and incorrect decisions can be immediately fed back to the trainee.

Jan 1, 1998

By A. G. Nitschke, I. Ossenbühl, W. Dolsak, J. Dase

"The authors have successfully developed an innovative rehabilitation concept, which meets actual technical standards and minimizes the impact on the operation of a key conveyor tunnel. A flexible and stepwise ground support rehabilitation program utilized a combination of yielding steel arch sections with grout filled fabric hoses, which represents a fast and safe backfilling support system. The system was developed and successfully installed at Bingham Canyon Mine, the world’s largest open pit copper mine, which also includes numerous underground structures. The mine is owned and operated by Kennecott Utah Copper (RTKC) a fully owned subsidiary of Rio Tinto. The system was implemented at a conveyance tunnel, which was originally designed as a railroad tunnel, but was re-commissioned and is currently used as a conveyor belt tunnel to transport ore from the open pit to processing facilities outside the mine. The existing structure showed signs of overstressing due to movements in the surrounding rock mass, which had to be addressed with a yielding support system in order to provide safe access into the tunnel. However, the impact on the conveyor belt operation due to support installation was to be limited to the bare minimum. Keywords: rehabilitation, yielding support, mining, stabilization, and grout filled fabric hose.PROJECT DESCRIPTION The C6 Tunnel is considered a critical asset for the operation of the Bingham Canyon mine as it comprises the standard route for transporting crushed ore from the open pit to the processing facility on a continuous conveyor belt system. The C6 tunnel was originally built as a railroad tunnel. Construction was completed in 1959. The original horse-shoe shaped cross section of the tunnel was 18 feet wide and 24-1/2 feet high [8]. The tunnel length was approximately 18,000 feet. During its service life, the railroad tunnel has undergone several transformations. The length of the tunnel towards the Bingham Canyon pit was shortened over the course of mining operations and the railroad tracks in the tunnel were removed. The invert was partially backfilled (3 feet typical) and a 72 inches belt conveyor was installed. Currently, the C6 Tunnel is nominally only approximately 21 feet high and 15,000 feet long. Ground Support and Rehabilitation Measures prior to 2014 The C6 tunnel was mined mainly with conventional mining methods, using drill & blast as the typical one. However, the weak geological sections, such as those encountered in the Fortuna fault, were excavated by shaping and chipping with machine-held moils [8]."

Jan 1, 2015

By M. Ploc

The grinding process in the mining industry has seen an increased throughput to accomplish greater production and lower overall production costs. Up to 8-9 MW, a single motor can be connected to drive the mill through a pinion and ring gear. Above the single-pinion drive limit, the dual pinion approach was developed to allow two drive motors to be used. This allowed the mill input power to be increased to 16-18 MW. Recent developments in clutch and ring gear materials and design are allowing greater input powers of up to 20 MW to be considered. To avoid damage to the ring gear by uneven load sharing between the two motors, it is critical that the two machines accurately share load. This is typically achieved by either drives capable of controlling load torque between motors or using a GE proprietary control system that accurately controls the load torque from each motor.

Jan 1, 2010

By Gregory M. Rumbaugh, Todd Kostecki, Christopher Mark

From 2015 to 2021, five massive pillar collapses occurred at four underground stone mines in the eastern United States. These events resulted in powerful airblasts that damaged mine infrastructure and mobile equipment, seriously injured miners, and disrupted underground travelways. Each of these pillar collapses propagated through the overlying strata, causing a subsidence basin on the surface. Pillar collapses are particularly hazardous for miners because they can occur with little warning and can affect miners that are far from the pillar collapse area. The following case studies show that certain factors may increase the likelihood of a pillar collapse. For instance, each of the five events involved the collapse of at least twelve benched pillars with width-to-height ratios (w/h) of 0.8 or less. This study elaborates on

By R. G. Hite, R. G. Rowe

The Crandon deposit, located in northern Wisconsin, is a 65.8 million ton Precambrian volcanogenic massive sulfide deposit which averages 1.4% copper and 5.8% zinc. The deposit is classic in origin, mineralogy and shape. Comprehensive understanding of the geology of both the deposit and the project site has been fundamental to the credibility and success of project feasibility studies and regulatory permit preparation. Mine development in the current investment and regulatory climate requires thorough and timely application of geologic tools. From 1977-1980 geologic activities were focused on the typical objectives of orebody delineation, reserve estimation, depositional geology, and general site investigation. Results from these programs formed the basis for the project's preliminary feasibility studies. Since 1981 further geologic investigations have provided criteria for project design, site management, and operational permitting. Activities have included mine entry position definition drilling, devel reserve modeling, large diameter (150 mm) core drilling for mill pilot samples, rock mechanics testing, and evaluation of the glacial overburden/deposit geohydrology for computer simulation.

Jan 1, 1992

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

4.1 THE GENERAL COMPOSITION AND PROPERTIES OF CLAY-BASED GROUTS Clay-based grouts are visco-plastic systems; they are made by adding structure-forming reagents to a clay mineral mortar. Small amounts of cement and various chemical ad¬ditives constitute the reagents. If necessary, fillers also are added. The distinguishing feature of clay-based grouts is that throughout their entire stabilization period they do not form a crystallized structure as does cement grout. The structure of stabilized clay-based grouts does not deteriorate during minor rock movement or upon the initiation of blast¬ing during the construction of a shaft or a drift. Because they can possess good rheological properties during the ini¬tial structure-forming period relative to cement grouts, clay¬based grouts are not easily eroded from large fractures and karstic cavities by flowing ground water. In addition, their finely dispersed clay particles facilitate a greater fracture penetration capacity, especially in dual porosity rocks. Numerous investigations of the structural-mechanical and rheological properties of clay-based grouts by STG have demonstrated that the most effective grouts for pre¬venting the inflow of ground water into underground work¬ings are grouts that have a cement content of 8 to 10% (90 to 120 kg/m3) of the clay grout by mass. The grout should have a density of 1.18 to 1.30 T/m3. Various additional substances and chemicals can be added as additional fillers and structure-forming reagents. In pure ground water at temperatures above freezing, sodium silicate in the amount of 0.8 to 1 % by mass normally is the only structure-forming reagent that is necessary if the proper clay is selected and if it is available. The production process for making clay-based grout is divided into two stages: 1) the production of an initial clay mortar with specified properties and 2) the production of a clay-cement additive grout mixture using the initial grout along with the structure-forming reagents, including the ce¬ment. The properties of clay-based grouts depend on the phys¬ical-mechanical properties of the initial clay mineral, the properties of the cement and the properties of the chemical reagents that are added. 4.1.1 TECHNOLOGICAL PROPERTIES OF CLAY-BASED GROUTS The dynamic shear stress' To, the viscosity 11, the static shear stress2 0, the maximum shear stress of the structured that the rheological and structural-mechanical prop¬erties of clay-based grouts must fall within the following limits: the dynamic shear stress To = 50 to 200 Pa; the viscosity -9 = 0.02 to 0.07 Pa sec; the static shear stress 0 = 150 to 600 Pa; the plasticity strength Pm of the structure one minute after preparation (according to P.A. Rebinder's method) equals 150 to 500 Pa. The plasticity strength 10 days after preparation is ? 0.15 MPa. The general relationship of the change in the structural strength of clay-based grouts relative to stabilization time is shown in Fig. 32. The figure shows that the structure-form¬ing process for clay-based grouts is characterized by three stages. These stages correspond to the time periods required for conducting the principal operations during the injection of the grout into a fractured aquifer as described below. Stage I, corresponding to the time period T1, reflects the small development of structural strength during early stabi¬lization. During this time, the structural strength must not preclude the capability to pump the grout. The time Tl must correspond to the length of pumping time from the moment water is shut off and mixing occurs until the pump stops forcing the grout through the manifold block and pipeline into the fractured rock as described subsequently herein. Stage II, corresponding to the time interval T2, reflects a sharp but controllable increase in the structural strength of the grout. The length of time period T2 is controlled by the addition of appropriate reagents. Stage III produces the final value of the structural strength. This final strength is used to design the dimensions of the isolation curtain. Consequently, the development of grouts is guided by two principal criteria: 1) the grout must develop the highest possible structural-mechanical properties and 2) it must be able to be pumped by a piston pump prior to final stabili¬zation. Successful grouting depends to a large extent on the correct design of a grout for each specific case. It is impor-

Jan 1, 1993

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982

By Keith E. Dyas, John Nelson, Ronald T. Johnson

GENERAL DESCRIPTION The Bruce mine of Cyprus Mines Corp. is located in Bagdad, AZ. The mining method used is open cut-and-fill. Of the annual production of 81 647 t (90,000 st), approximately 83% is taken from load-haul-dump (LHD) stopes and the balance from slusher stopes. All ore is produced from the area between the 1250 level and the 2300 level. The average travel time from the shaft pocket to the stope is approximately 5 min. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size, Shape, and Dip The Bruce ore body occurs in quartz-sericite schist with Dick rhyolite on the footwall and andesite on the hanging wall. Diabase dikes are found in the hanging wall; there is also a dike coming off the footwall and crosscutting the ore body. All of the rock types are of the Precambrian Yavapai series and have been subjected to regional metamorphism. A composite of the ore body is given in Fig. 1. The deposit is of massive sulfides occurring as a steeply dipping replacement body. On the upper levels the ore is veinlike with widths from 0.6 to 4.6 m (2 to 15 ft), dipping at 1.4 to 1.5 rad (80° to 85°). On the lower levels the ore is dipping from I to 1.2 rad (60° to 70°) with widths from 3 to 16.8 m (10 to 55 ft). The strike length varies between 107 to 183 m (350 to 600 ft). The rhyolite footwall generally has a knife-edge contact with the massive sulfides. The exceptions to this are the upper levels where there is a 1.5 to 3 m (5 to 10 ft) band of silicified sericite schist between the sulfides and the rhyolite. In the southern part of the ore body the hanging wall is tuffaceous andesite and andesite. In this area the contact is generally sharp and easy to follow. However, to the north there is a large chlorite schist zone that crosscuts the bedding and comes in contact with the massive sulfides. This is apparently due to hydrothermal alteration of the andesite. The chlorite schist is highly mineralized with chalcopyrite and pyrite and quite often forms economic pockets of ore. In the massive sulfides the chief ore minerals are sphalerite and chalcopyrite. Pyrite is the predominant sulfide with considerable pyrrhotite throughout. Bright arsenopyrite ouhedrons in fine grain massive sulfides are quite common. Occasionally small amounts of galena are seen, usually near the foot or hanging wall contacts. On rare occasions tennanite is associated with massive arsenopyrite. Minor amounts of quartz, calcite, and un¬replaced remnants of sericite schist occur, but essentially pyrite is the gangue in which the ore minerals occur. The ore values are in excess of 3.5% copper and 12.5% zinc with some silver and rare gold as byproducts. Ground Conditions The massive sulfides are generally self-supporting. One exception is in the 1850 stope where the ore body is 9 to 11 m (30 to 55 ft) wide and 152 m (500 ft) long. There are flat to shallow dipping slips and seams in the ore, creating extremely blocky ground. For support, old 25.4-mm (1-in.) hoist ropes were installed tensioned to 27 t (30 st), and then cement grouted over the entire length in longholes [14 to 15 in (40 to 50 ft) in length) drilled on 3-m (10-ft) centers from the level above. This has tied the formation together very successfully and virtually eliminated the blocky ground condition. Both the hanging wall and footwall are quite shaley in some areas. Reasons for Adopting Trackless Open Cut-and-Fill Methods First, any method other than open cut-and-fill would have caused too much dilution. The use of rubber-tired mining equipment in the pro¬duction stopes requires a footwall ramp. The inclines in ore will be mined out, so this ramp in the footwall will provide access to and from the stopes (Fig. 2). This incline is very expensive, but necessary to convert existing stopes to LHD mining. 'The final cost of ore mined by the LHD machines has not been determined. As of 1972, tons per manshift in the 2150 stope-the only one to complete a full cut-had increased from 7.58 t (8.36 st) to 12.83 t (14.14

Jan 1, 1982

By W. A. Haley

The use of diesel engines in underground hard-rock mines dates back to the late 1940s. For the first several years, they were used only occasionally, being limited to a few metal mines that experimented with crawler¬mounted front-end loaders, tractor-trailer hauling units, a few tractors for drill-compressor mounts, and utility cleanup machines. By the mid-1950s, track loaders had become commonplace in limestone mines and uranium mines on the Colorado Plateau in the United States, as well as in Canada. Use of crawler-mounted tractors as drill and compressor mounts also increased. By the end of the 1950s, rubber-tired loaders and some haulers began to replace the track-type machines and rail-mounted cars that had been in use. About 1960, the rubber-tired machines brought about a new era of underground mining mobility and flexibility, centered on a method commonly known as "trackless mining." Ultimately, many of the underground rail-type systems for loading and hauling were replaced by the trackless mining technique. ECONOMIC CONSIDERATIONS The size and nature of mineral deposits, plus ground control techniques, historically had dictated small open¬ings to the surface from many underground mines. The small mine openings led to the development of special rubber-tired loaders and haulers designed specifically for access through the small openings. However, some mines, particularly those in massive mineral deposits, are able to excavate and maintain very large openings, and some use modified room-and-pillar systems. With the large mine openings, the use of larger, more produc¬tive equipment such as that commonly found in surface mining becomes economical. In fact, productivity gen¬erally increases at a more pronounced rate than machine size increases because many of the larger machines were designed for heavy-duty shot-rock applications in surface mines and construction sites where the handling of blasted rock is common. Table 1 can be used as a very Table 1. General Productivity Comparison for Conventional Machines In Underground Use (Shot-Rock Conditions) 2.3 m3 (3 cu yd) 4.6 m3 (6 cu yd) Loader Loader Expected Surface 230 t/h 540 t/h Production (250 stph) (600 stph) Expected Underground 90 t/h 270 t/h Production (100 stph) (300 stph) Expected Total Efficiency: Surface 40%-60% 50%75% Underground 25%.-40% 30%50% Expected Useful Machine 8000 hr 12,000 hr Life Before Replacement general comparison of the production and efficiency between small and large machines. Combining greater productivity often inherent in larger machines, with reduced downtime resulting from using fully developed machines with fast parts and service backup, some mine operators have been able to reduce material handling costs appreciably while reduc¬ing manpower requirements for operators and main¬tenance men. Large mine openings increase the amount of rock that must be handled in the development work, and they sometimes increase the dilution in stopes or rooms, de¬pending upon the dimensions of the ore zone. Providing adequate space for the unrestricted operation of large surface mining machines could, therefore, lead to more waste segregation and handling costs. It could also cause greater ore dilution that would result in a lower grade of ore being delivered to the processing plant. The tradeoffs between opposing cost factors must be reconciled and balanced to achieve the best overall cost of the crude ore, concentrates, or product. EQUIPMENT MODIFICATIONS Loaders and haulers designed for surface mining are seldom used underground in their standard con¬figurations without some modifications. If done, the modifications generally are made by the equipment dealer and/or the user, and the modifications usually include one or more of the following items: 1) The exhaust stack is lowered, and its direction is changed. Usually, it is repositioned horizontally to the rear, or it is fed into the engine fan to diffuse the exhaust gases. 2) The operator's position is lowered by either lowering the seat or changing the seat to a side mount. 3) The operator controls are adjusted to fit the new operator position. 4) Other components, such as the radiator and loader tower, are lowered. 5) Special bumper guards are mounted at the base of the radiator area. 6) An exhaust conditioner is mounted and con¬nected, using either a catalytic or a water-type condi¬tioner, or both. This usually is controlled by the safety and health regulatory authority having jurisdiction. 7) The positions of other components are rearranged

Jan 1, 1982

By M. R. Yalamanchili, J. D. Miller

Discussion by M.R. Yalamanchili and J.D. Miller The authors, M. S. Celik et al., should be recognized for their efforts to describe the flotation behavior of boron minerals. In the case of borax and other soluble salt minerals, analysis of the flotation chemistry has been difficult because of the high ionic strengths associated with these soluble salt systems. However, considerable progress has been made in this area, and recently a surface charge/collector colloid adsorption model was proposed by Miller and his coworkers to explain the collector adsorption phenomena observed in soluble salt flotation systems (Milleret al, 1992; Yalamanchili et al., 1993; Miller and Yalamanchili, 1994; Yalamanchili and Miller, 1994a: Yalamanchili and Miller, 1994b). In this work, the sign of the surface charge of alkali halides in their saturated brines was established on the basis of nonequilibrium electrophoretic mobility measurements by laser-Doppler electrophoresis (Miller et al., 1992). Generally, these results are what would be expected from the simplified lattice-ionhydration theory. This electrokinetic information coupled with the stability and prevalence of collector colloids in such soluble salt flotation systems indicates that the selective flotation of alkali halides is due to the adsorption of oppositely charged collector colloids by heterocoagulation. Experimental flotation/bubble attachment results for 21 different alkali halides (Yalamanchili et al., 1993; Yalamanchili and Miller, I994b) confirmed that the flotation response of soluble salt minerals with weak electrolyte collectors can best be explained by the adsorption of oppositely charged collector colloids rather than by the adsorption collector ions and/or neutral molecular dipoles as originally suggested by many researchers (Fuerstenau and Fuerstenau, 1957; Schubert, 1967; Roman et al., 1968). In addition, the flotation of certain alkali oxyanions (Pizarro et al., 1993) and double salts such as schoenite and kainite can be explained by the same collector colloid adsorption mechanism (Miller and Yalamanchili, 1994). The borax flotation results reported by Celik et al. need to be examined in terms of the above mentioned surface charge/ collector colloid adsorption model. Unfortunately, the authors seem to be unaware of this recent work that nicely describes soluble salt flotation with weak electrolyte type collectors such as amines and carboxylates. In view of our past work, the flotation characteristics of borax were of particular interest, and, in this regard, the results of dodecyl amine flotation of borax reported by Celik et al. have been examined in further detail in the light of experimental results from our laboratory. In our research, a vacuum flotation technique was used to study the flotation response of borax (Na2B407.10H20), which has a solubility of 39 g/L at 25 °C) with dodecyl amine hydrochloride as collector. These chemicals were purchased from Eastman Kodak and used as received. Saturated solutions of borax at desired pH values were prepared by continuously stirring the salt solutions over a period of about 10 hrs. It should be mentioned that the conditioning time to achieve equilibrium is an important variable and can significantly change the flotation response of some soluble salts (Yalamanchili et al., 1993). Collector was added to the saturated borax solutions containing about one gram of 100x 150 mesh borax particles, and conditioning was done for about 20 minutes prior to flotation. The borax flotation recoveries from saturated brine are presented in Fig. 1 as a function of collector addition at the natural pH of 9.3, as reported both by Celik et al. and as measured in our laboratory. In addition, the region of precipitation for the dodecyl amine hydroborate is included in Fig. 1. It can be seen in Fig. 1 that the flotation response curves are separated by about one order of magnitude in R12NH3CI collector addition. The flotation results of Celik et al. show that the maximum borax recoveries can be obtained below the solubility limit of the dodecyl amine hydroborate collector. However, in our experiments borax flotation seems to occur only after the precipitation of the dodecyl amine hydroborate collector as might be expected from the collector colloid adsorption model (Yalamanchili et al., 1993) if borax were negatively charged. Further analysis by nonequilibrium and equilibrium electrophoretic mobility measurements for borax indicates that borax is negatively charged at the natural pH of 9.3, as discussed below. The reliability of the nonequilibrium electrophoretic measurements has been demonstrated previously for alkali halides and alkali oxyanions (Miller et al., 1992; Miller and Yalamanchili, 1994). The equilibrium and nonequlibrium electrophoretic measurements for borax were found to be consistent and are presented in Table 1. These results provide clear evidence that borax carries a negative surface charge in its saturated brine (pH 9.3), and the sign of the surface charge of borax reverses and becomes positive if the pH is reduced to 8.6. The equilibrium between borax and its saturated brine can be described by the following reaction: [2Na2B407.1OH2O-4Na++B407=+HB4O7 +OH+19H20] It appears that the oxyanions of the borax lattice provide

Jan 1, 1995

By Rudolf E. Greuer

Modeling a mine ventilation system is a formidable mathematical problem requiring the solution of twice as many equations as there are airways with half of these being quadratic equations. The addition of a mine fire to the problem adds many more complexities. The U.S. Bureau of Mines and Michigan Technological University attacked this problem in stages. During the 1970's and 80's, a series of computer simulation programs were developed which offered progressively increasing levels of modeling capability and accuracy. The current version of this software, MFIRE 2.0, was released in 1993. This programs provides for full dynamic state mass flow-based modeling of a mine ventilation system, and includes natural ventilation, mechanical ventilation, air reversals and recirculation, methane and fire contaminant spread, and moisture condensation and evaporation in the mass flow and heat exchange calculations. This program is a useful tool for routine ventilation planning and can significantly improve mine safety by enabling detailed quantitative analysis for mine emergency preplanning. This paper reviews the evolution of the MFIRE program from its infancy as a simple digital network simulator replacing analog computers to the present version.

Jan 1, 1994

By Hans Svensson

INTRODUCTION The Alimak method of raising (Fig. 1) was intro¬duced in 1957. Because of its flexibility, economy, and speed, it soon became the most used system of raising in the world. With the same unit, raises of various areas, inclinations, and lengths can be driven. Air-motor¬-driven raise climbers normally are used in mines with short raises (Fig. 2). For longer raises electric or diesel-¬hydraulic drives are used. The Alimak raise climber is designed entirely in ac¬cordance with the instructions of the Swedish Board of Industrial Safety (Arbetarskyddsstyrelsen). There are good margins of safety concerning the breaking strength of the material and the raise climber comprises safety devices and features that practically eliminate the risks of accidents. The drive gear of air-driven climbers is equipped with an air-operated brake which is automatically activated when air for the motors is shut off. There is a safety device which activates automatically at overspeed and a speed-regulated brake for descent by gravity, by which the raise climber can be taken down to the bottom of the raise in case of cutoff of the air supply. The raise climber climbs along a pin rack welded to a guide rail which also comprises pipes for air and water. The guide. rail can be extended by using 1- or 2-m (3.2¬or 6.4-ft) sections as the driving progresses. Each guide rail section is bolted to the rock wall by special expan¬sion bolts of Alimak design. Men ride up to the face in the cage, traveling com¬fortably and easily while material is transported on the platform. There is no dangerous climbing up ladders. The men are well protected during ascent and arrive in good condition at the face. The greatest advantages of raise climber systems are flexibility and versatility. The equipment provides a comfortable working position in vertical as well as in¬clined raises of any area and length. The direction may

Jan 1, 1982

By Eleanor Sonley

In October 2000, temporary seismographic stations were installed above the Trail Mountain Mine, Emery County, Utah, USA to supplement a regional network. Over seven months, 1826 mining-induced seismic events were recorded with a magnitude range of M 0 to M 2.2 (Arabasz et al., 2002, 2005). Routine event locations for this dataset cluster around the longwall mining activity during the recording interval. However, applying a double difference relocation technique to the routine locations significantly improves the correlation between mining activity and event location. The resulting relocations depend greatly on station distribution and the number of arrival time picks. We use this dense temporary array to assess station distribution requirements for applying the double difference technique at other mines by systematically removing data from available stations, thus simulating sparser networks. The insight gained using the Trail Mountain dataset is used to develop a metric for assessing the usefulness of above-mine networks and to develop a tool for expanding seismic network coverage. To verify the appropriateness of these tools for wider application, we examine data from two additional mines located in the western United States. Improving the relative locations of mining-induced seismic events highlights the association of these events with active portions of the mine. Clusters, regions of quiescence and relative timing of events can be used by mine operators to assess hazard, optimize production, and evaluate performance of mine plans. The lessons learned from this study may be used in future network design to improve accuracy of event locations at other mines.

Jan 1, 2009

By R. Reddy Kallu

Prior to July 2007, the only accepted method by MSHA for evaluating the mine seal design and performance is the full-scale explosion tests conducted in the NIOSH?s Lake Lynn Experimental Mine (LLEM) for deeming a seal design suitable for use in U.S. coal mines. Since the new regulations came into effect use of these conventional full-scale explosion tests for evaluating mine seals for the new elevated design standards may not be practicable. Generation of such high explosion pressures in experimental mines is dangerous and may cause permanent damage to the testing facilities. Further, they require significant amount of time, labor and money. Considering these factors, there is a need to develop scientific methods for evaluation of new seal designs and for understanding the dynamic seal response. This paper tries to address the reinforced concrete (RC) seal design using the three dimensional finite element based explicit software ABAQUS. Full-scale explosion test data available from NIOSH?s LLEM on 20 psi reinforced concrete seal is used for validating the numerical models. Further in this paper, based on the results from the validation models, a reinforced concrete seal is designed for 120 psi design standard for typical entry dimensions.

Jan 1, 2009

By Nan Hua

As a high production and high efficient mining method, top coal caving method is now widely used in China. However, as more extra-thick coal seams are mined, it becomes more and more difficult to make the top coal fully broken up, resulting in more and more top coals flowing into the gob, rather than onto the rear AFC at the working face as designed. This research is aimed to investigate the failure mechanisms of the extra-thick top coal caving mining method.

Jan 1, 2009

By Q. Feng

Study has been carried out to apply fly ash as a high strength, water-resistant precast construction material through geopolymerization. Experiment results show that the working conditions such as water content, the concentration of NaOH, curing temperature and curing time, significantly affect the mechanical property of geopolymer matrix. Through optimization, an above 100 MPa compressive strength has been achieved with the geopolymerization products. Water soaking tests show that the geopolymerization product has a very high water resistance without losing noticeable compressive strength even after a one month soaking time. To elucidate the geopolymerization mechanism, microscopic techniques such as SEM/EDS and XRD are also applied to investigate the microstructure, the elemental and phase composition of geopolymerization products. The findings of the present work provide a novel method to apply fly ash as a high strength water-resistant precast construction material. INTRODUCTION Geopolymerization is a process of obtaining a polymeric structure from aluminosilicate by dissolving aluminosilicate sources in a strong alkali (NaOH) solution and condensation of free [SiO4] and [AlO4] tetrahedral units (Davidovits, 1985 and 1988; Duxson et al, 2005; Xu et al, 2000, 2002). After being cured at an elevated temperature for a specific time, the geopolymerization product shows a greatly increased compressive strength. Therefore, the product can be used as construction material such as bricks and road pavement. It provides a very promising method to apply aluminosilicate sources such as fly ash as a construction material for a sustainable development. Fly ash is a main industrial waste which needs to be carefully impounded because of its potential impact on environment. Many efforts have been tried to utilize fly ash as construction materials substituting Portland cement because of the following reasons. First, fly ash is a silica/alumina rich aluminosilicate source, which makes the specific material an ideal source for the geopolymerization process. Secondly, the particle size of fly ash is usually very small (< 50 microns) and it saves a lot of energy consumed during the production process of cement for grinding. Thirdly, fly ash is an ideal source for geopolymerization because it is reactive to alkali even at a mild temperature. In the past twenty years, many studies have been carried out on the geopolymerization of fly ash to a make construction material. However, there are some disadvantages associated with the conventional methods of treating fly ash (Palomo et al, 1999; Perera et al, 2007; Temuujin, et al, 2009; Hanjitsuwan et al, 2014). First, in most of the proposed geopolymerization methods, a salient amount of sodium silicate has to be added. It therefore makes it complex whether the achieved mechanical strength of geopolymerization products is due to a real geopolymerization process between fly ash and alkali or just largely the solidification of metal silicate in air at an elevated temperature. Secondly, the authors have noticed that when the geopolymerization product is soaked in water, the solution pH increases sharply and the products will lose its compressive strength gradually in water. Thirdly, the compressive strength of the geopolymerization product as reported is usually less than 50 MPa.

Jan 1, 2019

By Andre C. Zingano

It is a very difficult task to estimate the immediate roof and support behavior in a room-and-pillar mining section in the initial phase of the mining project. There is little information available at this phase, and it is restricted to geomechanic characterization of the rock layers from cores samples, and the in-situ instrumentation of experimental entries and intersections. Even being a pilot excavation for roof support test a good indicator for the behavior of the immediate roof, its behavior for a full excavation mining section is different from a single intersection of two entries, mainly because the stress redistribution caused by the multiple entries excavation and face advance for a full excavation section. The objective of this paper was to calibrate a 3-D numerical model of an experimental intersection, based on the in-situ instrumentation of an experimental intersection excavation, and the geomechanic description of the rock layers. The numerical model was built with the same specifications as the excavation test, including the entry geometry and roof support. The numerical model calibration was based on wire extensometers that were installed at different positions in the roof of the pilot excavation, and on the geomechanic characterization of the rock layers that form the roof, coal seam, and floor. Therefore, it is possible to estimate whether the roof support specification adopted in the test phase will be good enough for the development of the entire mining section.

Jan 1, 2009