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By G. W. Riley, D. E. Pickett

HINDERED SETTLING CONCENTRATION In the free settling of mineral particles in a liquid, the falling particles are at a distance from each other so that no particle is affected by its neighbor. In hindered settling, the concentration of particles is sufficiently high so that each particle is affected by its proximity to other particles in the suspension. Richards and Locke86 have described the hindered settling phenomenon as the condition “.. where particles of mixed sizes, shapes and densities in a crowded mass, yet free to move along themselves, are sorted in a rising current of water, the velocity of which is much less than the free-falling velocity of the particles but yet fast enough so the particles are in motion." This is the condition normally encountered in mineral con¬centration processes. The well known Newton equation for free settling of coarse (ap proximately +10-mesh, 1/16-in. or -- 2000-µm) spherical particles is:87 where v," is free settling velocity, cm/sec; p' is density of the fluid; p is density of the particle, g/cm3; d is particle diameter, cm; g is acceleration due to gravity, cm/sec2; and Q is coefficient of resistance, dimensionless, ~0.4. For the settling of fine spheres in water (approximately 150 mesh or 100 µm) the equation of Stokes pertains:87 where µ is viscosity of the fluid in poises and the other symbols have the same meaning as those in Eq. 1. For particles whose size lies between about 10 mesh (--2000 µm) and 150 mesh (100 µm), their settling velocity can be determined from experimental data. These data are available in convenient form in the text by Taggart88 based on the original work of Richards. Alternatively, a Reynolds number-coefficient of resistance plot may be used to determine the settling rate of such particles.87 The settling rate of spherical particles under hindered settling conditions can also be calculated from Eqs. I and 2 by replacing p', the density of the fluid, by p" the apparent density of the suspen¬sion. The concentration of particles in the fluid thus imparts an appar¬ent density to the composite fluid or suspension greater than that of the liquid alone, resulting in a buoyant effect on the larger particles. Particle shape affects the settling rate of both coarse and fine particles. The general effect is to reduce their settling velocities and the effect is greater for coarse particles and for those settling under hindered settling conditions than for fine particles or free settling ones. For two particles of differing densities but settling at the same velocity under Newtonian conditions, the ratio of their diameters from Eq. 1, called the free settling ratio is: where L signifies the lighter particle and H, the heavier particle. Under Stokesian conditions the exponent would be 0.5. For hindered settling conditions the fluid density p' is replaced by the apparent density of the suspension, p", to obtain a generalized equation for the hindered settling ratio: assuming both particles settle in approximately the same regime. The free settling ratio as given by Eq. 3 has been called by Taggart88 the "concentration criterion" and is used to predict the effectiveness of any gravity concentration process (see Introduction to this section). Based on Eqs. 3 and 4, if two particles of densities pH and p,, settle at the same velocity, the diameter of the lighter particle will be larger than that of the heavier particle. For example, in the case of galena (pH = 7.5) and quartz (pL = 2.65) settling in water (p =1.0) under free settling, Newtonian conditions 3.9. Thus, a quartz particle nearly four times as large as a galena particle will settle at the same velocity. Any quartz particle just slightly less than four times the diameter of the largest galena particle may be separated from it. Under hindered settling, Newtonian condi¬tions in a suspension where p" = 1.65, dL/dH = 5.85 or any quartz particle just slightly less than about six times the largest galena particle may be separated from it. Reference to Eq. 4 indicates that a superior separation between two minerals of differing densities is favored by: (1) coarse particles settling under Newtonian conditions, (2) a large difference in (pa - PL), and (3) separation under hindered settling conditions where p" is high. Of course, there are practical limits to increasing p" excessively because at very high percent solids suspen¬sion fluidity would be lost and the hindered settling separation process defeated. Examples of hindered settling separators are the Dorrco-Fahren¬wald sizer,89 the Rheolaveur box 90 the Spitzkasten,89 and the Willoughby washer.91 These devices make a mineral separation on the basis of both specific gravity and size and all of them are essentially obsolete except for the Dorrco-Fahrenwald sizer and similar devices which still find application for the removal of coarse particles from a much finer particle assemblage and for preconcentration ahead of shaking tables. However, nearly all gravity concentration processes (jigs, tables, flowing film concentrators, heavy media separators) and many sizing devices (sizing classifiers, clarifiers, thickeners, hydrosepa¬rators) make use of the hindered settling phenomenon during the separation of particles. JIGGING Introduction In jigging, a mixture of ore particles, supported on a perforated plate or screen in a layer or "bed" with a depth many times the thickness of the largest particle, is subjected to an alternating rising and falling (pulsating) flow of fluid with the objective of causing all

Jan 1, 1985

By S. J. Bluhm

INTRODUCTION Lesotho is a mountainous area of southern Africa from which water is to be exported in an extensive tunnel system, to industrial regions inland. The related tunnelling project has involved a num- ber of drives using tunnel boring machines [TBMs] to excavate about 100 km of 5 m diameter water tunnels [von Glehn and Bluhm, 1995). This paper describes the ventilation and cooling of some of the tunnel drives from both the operational and design points-of-view with a particular focus on heat generation. There were many common features in all of the drives but this paper is focused mainly on the Hlotse drive which was 18,4 km long. The drives were ventilated using forced ventilation systems to provide appropriate air flow throughout the tunnels and face zones. In addition, the Hlotse drive required refrigeration equip- ment which provided chilled water to the tunnel. While the sec- ondary ventilation systems play an important role in gas and dust handling, the paper concentrates on the primary ventilation and cooling issues. The ventilation of these tunnels was an exacting exercise be- cause: • Rock temperatures and geothermal heat flow were high. • TBMs with relatively high power ratings were used. • Diesel locomotives were used. • Drives were relatively long. • High altitude meant a low air density. An important feature was the simulation and monitoring of the ventilation and heat flow components and the project was characterised by analysis, monitoring and ongoing tactical decision-making throughout the progress. The thermodynamics of the systems were complex because there were many interactive effects and analyses were carried out using special computer pro- grams. The monitoring confirmed the accuracy of the models, and in this manner it was possible to confidently ensure healthy and safe working conditions and still minimise costs. Local ambient climate conditions range from temperatures higher than 35 "C in summer to below -10 OC in winter. Based on available statistical data and the thermal storage/damping effects in the system, design summer ambient conditions were taken as 15 OC/25 "C wet-bulb/dry-bulb. The barometric pressure was 80 kPa and due to the altitude, the ambient air density was only 0,9 kg/m3. The local Authority specified a maximum in-tunnel wet- bulb temperature [at any point] of 32,O OC and a mean wet-bulb temperature [from all locations] of 27,5 OC maximum. The maxi- mum height of ground cover above the tunnel was 1 200 m and the maximum virgin rock temperature was 41 OC; see Figure 1. Diesel dilution criteria specified by the local Authority was a minimum of 0.1 m3/s per rated kW of diesel engine. Other requirements related to gases such as CO, CO2, NOx and CH4 [and the need for intrinsically safe equipment] but these are not of direct relevance to this paper. The actual average face advance was about 30 m/d with good days achieving 60 m/d and good months achieving 1 000 m [23 working days]. The original design tunnelling rate was 50 m/d. DESCRIPTION OF HLOTSE DRIVE VENTILATION AND COOLING SYSTEM The ventilation requirements in the tunnels were dictated by heat and diesel dilution needs. The best ventilation and cooling policy is generally a balance between using increased quantities of fresh air or refrigeration [or both]. In this particular scenario it turned out that, since the diesel emission criteria required large quantities of air, the refrigeration needs were modest. The drive was ventilated using a ducted, forced ventilation system from fans located at the portal. The maximum ventilation requirement was 51 m3/s when the drive was at 18.4 km. From a heat flow point of view, the worst scenario was a heat load of 3.5 MW when the drive was at 7 km. This was cooled by the ventilation air and a supply of chilled water to the tunnel. Refrigeration and chilled water system In the design phase, a detailed comparison was carried out between two general alternatives for providing refrigeration. First, was a system in which refrigeration sets and air coolers are installed on the TBM train; the refrigeration sets are cooled by condenser water piped to and from cooling towers at the portal. Second, was a system in which refrigeration water chillers are in- stalled at the portal and chilled water is piped into the tunnel. The detailed comparison indicated that the capital and running costs of the second system were at least 60 % lower than the in-tunnel plant. There were also many obvious practical benefits for favouring the portal system. The refrigeration plant supplied 23 11s of cold water at a temperature of 10 OC. After providing the cooling effect in the drive, the water returned to the portal where it was initially cooled in open-circuit evaporative pre-cooling towers, chilled in the refrigeration plant and then returned to the tunnel. The cold water flowed into the tunnel in an insulated supply pipe and returned in an uninsulated pipe; the water was simply circulated to the end of the pipe and returned. The cooling effect in the tunnel was achieved entirely through heat transfer from the pipe [long linear heat exchanger] and no air coils or other heat exchangers were required. The cooling requirements were satisfied by the heat transfer to the returns chilled water steel pipe [200 mm]. The pipes were eventually installed to a maximum distance of 10,8 km in what was considered a very practical and cost effective solution.

Jan 1, 1997

By H. Wagner

INTRODUCTION One of the most important tasks facing the opera¬tors of deep-level gold mines is supporting the roof strata in the immediate vicinity of the stope faces. Some of the deeper gold mines operate at depths ex¬ceeding 3.0 km; the extreme mining depths and the tabular nature of the gold-bearing reefs result in high stress concentrations at the abutments of the stope faces. These, in turn, lead to extensive fracturing of even the hardest quartzites. Although most of these fractures develop in a stable manner, there have been many instances where rock failure occurred suddenly, resulting in a considerable amount of damage to the underground workings. Such violent rock failures commonly are known as "rock bursts," and they are typical of many hard-rock mines that are operating at great depths. To safeguard mining personnel and equipment against the effects of stable and unstable rock failures, adequate support must be provided to the fractured roof as quickly and as effec¬tively as possible. Traditionally, timber props and timber packs have been utilized to provide roof support in gold mine stopes up to 1.5 m wide. Both methods of support suffer from the fundamental disadvantage of generating sup¬port only in reaction to movement of the strata. Since unrestrained strata movement often results in a break¬down of the frictional forces that preserve the stability of the intensely fractured roof strata, it is essential that stope supports have rapid load-bearing characteristics. An equally important requirement is that the roof sup¬ports be able to accommodate displacements of the rock mass. A third requirement of roof supports in gold mine stopes is that they must be able to accommodate the high rates of stope convergence typical of rock bursts. The hydraulic prop is the obvious choice for sup¬porting the stope faces in deep-level hard-rock mines. Since hydraulic props can be installed with a positive setting load, they can provide roof support that is independent of strata movement. By equipping the hydraulic props with yield valves, they meet the require¬ment of being able to accommodate rock-mass displace¬ments. However, the requirement of being able to accommodate high rates of stope convergence without overloading or damaging the rock strata was rather difficult to meet. The first introduction of hydraulic roof supports into gold mine stopes showed that conventional hy¬draulic props, such as those used in coal mines, were completely unsuitable for a number of reasons, in¬cluding: 1) Since drilling and blasting are used to break the ore in gold mine stopes, supports installed close enough to be effective are exposed to the vigor of the blast. Conventional coal-mining props suffered extensive blast damage, despite being protected by a blast barricade. 2) The environmental conditions in gold mine stopes favor corrosion. The high quartz content and the abrasiveness of the ore further aggravate the problem by damaging the protective coatings on the hydraulic surfaces. Simply changing the design of gold mine props to have the piston project from the bottom rather than from the top resulted in dramatic improvements in the service life of the props. The improvement was attributable to external water in the stope washing the fine quartz grit away from the hydraulic surfaces, rather than concentrating the particles between the cylinder and piston of the prop. 3) The most obvious deficiency of conventional coal mining props was their inability to yield rapidly enough to accommodate the high convergence rates of rock bursts without overloading the supports. As a result of the first disappointing results with hydraulic supports in narrow hard-rock stopes mined by drilling and blasting, a comprehensive research and development program was undertaken. The hydraulic props resulting from several years of research and exten¬sive field trials are described herein. To date (1978), more than 130,000 of these props have been manufac¬tured and used successfully in a number of mines and under a variety of geological conditions. Although many of the detailed specifications are directed spe¬cifically to gold mining conditions, there are a number of points that are of general importance in the design of underground support systems for tabular excavations in hard-rock mines of any type. SPECIFICATIONS FOR RAPID-YIELDING PROPS Yield Requirements The yield requirements for hydraulic props in deep¬level hard-rock gold mines are summarized in the fol¬lowing paragraphs. Slow Yield Load: The design of the yield valve must allow the yield load to be preset to between 300 and 400 kN. This flexibility is necessary to account for the differences in the strength properties of the roof and floor strata. Drop in Load Under Slow Yield: The drop in sup¬port load during a slow yield should be as small as possible, but it should not exceed 5% of the slow-yield load. Rapid-Yield Load: The load exerted by the prop during periods of rapid stope convergence should be as close as possible to the slow-yield load. The rapid-yield load should not exceed the slow-yield load by more than 15% at yield rates of 1.0 m/s. The limitation of the upper valve of the yield load is important; it deter¬mines the hydraulic performance of the yield valve. The hydraulic props that satisfy this criterion have performed well, even under severe rock bursts. Load Changes Under Rapid Load: On some of the earlier rapid-yielding props, conditions of dynamic in¬stability were observed; these conditions resulted in a complete breakdown of support resistance. It is essen¬tial that the force oscillations during rapid yielding be as small as possible, not exceeding -*5% of the average rapid-yield load at a yield rate of 1.0 m/s. Longitudinal Resilience: During rock bursts, vibra¬tions of the roof and floor sometimes are observed. To

Jan 1, 1982

By A. L. Wilder, S. N. Dixon

Introduction Coeur d'Alene Mines Corp.'s largest precious metals property is located in the historic Rochester Mining District 40 km (25 miles) northeast of Lovelock, NV. The property encountered cold weather operational problems soon after its fall start-up in 1986 due to its elevation of over 1830 m (6000 ft). The problem of ice buildup on the heaps because of sprayed solution application was faced immediately. It was felt that allowing ice to build up all winter long until a spring thaw was impractical due to the large area under leach. Further, the operating cost and delivery schedule for a solution heating system was unacceptable. The development and installation of a leach solution distribution system using drip emitters made efficient, cost-effective winter operation possible. Other benefits of this system have also been observed and are discussed here. General process description 15,422 kt/day (17,000 stpd) of - 1.27-cm (-1 /2-in) crushed ore from the three-stage crushing plant are delivered to the leach pad using 77.1 t (85 st) rear dump haul trucks. The ore is drifted into place with a D-9 bulldozer. Leach panels are contiguous and are approximately 8861 m'(90,000 square ft) in area built in 6-m (20-ft) lifts. New panels are built on top of older areas to a final height of 61 m (200 ft). Each panel is ripped and cross-ripped prior to leaching. Barren solution is distributed to the heap using drip emitters at rates of 0.02 to 0.41 L/min/m2 (0.0005 to 0.01 gpm per sq ft), depending on the age of the panels. The pH of the leach solution is 10.7 with a cyanide concentration of 0.75 kg/t (1.5 lb per st). Approximately 50% of the silver and 80% of the gold are finally recovered. Pregnant solution percolates though the heap and flows by gravity into one of two 9.46 ML (2.5 million gal) pregnant solution ponds. The solution is then pumped to a conventional Merrill-Crowe process plant. Clarification takes place in three 9464 L/min (2,500 gpm) capacity filters. The solution is then pumped to a packed vacuum deareation tower for the removal of dissolved oxygen. Typical deareated solution contains 0.7 parts per million dissolved oxygen. Precipitation of gold and silver is accomplished by adding a zinc dust slurry to the deareated solution at the suction of the filter press feed pump. Precipitated gold and silver are recovered in three recessed plate and frame filter presses. Barren solution is discharged into a 11.7 ML (3.1 million gal) pond where cyanide makeup occurs. This solution is pumped back to the heap for further leaching. The precipitate filter cake, containing approximately 75% dore (Ag + Au), is then fluxed with anhydrous borax, soda ash, sodium nitrate and fluorspar to yield a neutral, bisilicate slag. The fluxed precipitate is then charged into a propane-fired melting furnace and heated to 1150° C (2100° F) for 3 1/2 hours. Slag and dore bullion are poured into conical cast iron pots yielding buttons of 800 to 1000 troy oz. The dore typically contains 98.5% silver and 1 % gold. Slag is crushed and tabled to recover the trapped dore blebs and beads. Concentrate from the table is returned to the furnace. Table tails are sent to the crushing circuit and out to the leach pad. Solution application The area kept under leach at Rochester is approximately 130 000 m2 (1.4 million sq ft). Barren solution is delivered to the pad at 21.2 kL/min (5600 gpm) for a resultant application rate of 0.16 L/min/m2 (0.004 gpm per sq ft). A traditional solution sprinkling system using No. 12 Senninger Wobblers with individual pressure regulators was installed at the onset of leaching activities. The Wobblers were placed at 9.1-m (30¬ft) staggered centers and were fed off of a gridwork of Yellowmine plastic piping. Solution flow rates were moni¬tored to each panel. The onset of cold weather with an average nighttime temperature of -12° C (10° F) made it apparent that continual operation would not be possible with the sprinklers. A significant amount of ice was built up on top of the heap, making maintenance and pipe removal dangerous, if not impossible. Leach solution application was restricted to daylight hours to inhibit ice formation. Process plant flow rates were reduced to maintain steady-state operating conditions. However, as daylight temperatures dropped below freezing, ice continued to accumulate due to the sprays. Besides the obvious operating hazards brought on by the growing icefield, there was also the potential environmental hazard associated with an early thaw melting the ice too rapidly for the solution containment facilities. One other option for preventing ice formation was heating of the barren solution prior to spraying. Initial plant design allowed for expansion of the propane storage and distribution system as well as modification of the barren piping for a solution heater. This option was not exercised because the operating costs for an adequate system would have been prohibitive, and timely delivery of a system was not available. An investigation was conducted on the various drip irriga¬tion products available, since subsurface solution applicators would eliminate ice formation altogether. Systems utilizing external flow emitters were ruled out because of their ten¬dency to clog when buried. Emitter systems using perforated tubing were also eliminated from consideration due to their inability to adequately control flow over required lengths of tubing. An in-line emitter system was finally selected which demonstrated clog resistance and adequate flow control, enabling direct burial.

Jan 1, 1990

By M. Bleeck

Calcium carbonate (CaCO3) is one of the most ubiquitous and versatile minerals found in the earth's crust. Its availability, attractive physical properties and relatively low processing cost make CaCO3 the most widely used filler material today. It is mined in three different forms - chalk, limestone and marble. Each physical form of CaCO3 has different qualities due to differences in postdepositional geology. But the chemical composition remains the same, with CaCO3 an inert component of the finished product. In the past, the paper industry largely left CaCO3 by the wayside, as it cannot withstand the acid-based papermaking process. But conversion to an alkaline system by many US mills changed this picture. Carbonate suppliers have put time and effort into research and development, demolishing barriers and creating new possibilities for what is a simple, natural product. By controlling particle size, size distribution and particle charge, the industry uses ground calcium carbonate (GCC) as a performance enhancer and as an extender for more expensive ingredients. It is estimated that the United States uses 3.6 to 4.1 Mt/a (4 to 4.5 million stpy) of CaCO3. Consolidations and mergers are taking place in the industry. Of the 12 major GCC producers in operation nine years ago, seven are left. Mineral Technology (US) is the dominant precipitated calcium carbonate (PCC) producer with more than 50 satellite plants worldwide. Other producers include Georgia Marble (French); Franklin Industries (US); OMYA (Swiss); J.M. Huber (US); ECC (English); and Filler Products (US). Global Stone PenRoc (Canada) is the only newcomer. In addition to this group, there are three small producers left in North America, each with a capacity of less than 100 kt/ a (.110,000 stpy). The trade organization operating as the Pulverized Limestone Division of the National Stone Association renamed itself the Pulverized Mineral Division, to increase its membership pool. The US paper industry is a predominant GCC con¬sumer, using approximately 800 kt/a (882,000 stpy) at an approximate cost of $130/t ($1.43/st). European paper mills pioneered alkaline papermaking. In the early 1960s, they began using GCC as filler and soon thereafter added GCC to their coating formulations. A decade later, the North American paper industry followed suit. The conversion from acid to alkaline paper production benefits the economic and performance aspects of the industry. Less pulp is needed, paper machine maintenance and effluent treatment costs are reduced, and sheet strength, opacity and brightness are increased. Perhaps most important to the reader, the sheet is desensitized to ultraviolet light, extending the paper's archival ability. CaCO3 can provide the papermaker with additional control of his paper. For example, PCC has long supplied the tobacco industry with a means to slow down the burning rate of cigarettes. Due to enhanced performance with regard to bulk and opacity, filler PCC use has risen to 1,500 kt/a (1,650 stpy) in the United States, at an approximate price of $130/t ($143/st). The majority is produced onsite at the paper mill, using "satellite plants." This concept reduces freight cost because only quicklime (CaO) is shipped to the mill, not CaCO3 slurry. The future of CaCO3 is encouraging. The amount of natural ground CaCO3 used is expected to double by the year 2005 to approximately 8 Mt (8.8 million st) worldwide. Acid papermaking practices will feel an increasing pressure to convert to an alkaline process as larger volumes of GCC containing paper enter the recycling market. CaCO3 reserves are plentiful. They will supply the ever growing demand for increasingly sophisticated paper. The plastics industry is supplied with almost 900 kt/ a (990,000 stpy) GCC at an annual growth rate of 4% to 5%. The price of a functional, inorganic filler, surface modified for the plastics industry, has an average selling price of $220/t ($243/st). GCC represents the most common filler, creating a product with higher gloss, better dielectric properties, impact resistance, weatherability and shrinkage control. CaCO3-filled plastics surround us - auto hubcaps and dashboards, shower enclosures, floor tiles, wire coatings, microwave dishes and Tupperware. The caulking and sealant industry is an enormous GCC user, with annual consumption requiring 1.13 Mt (1.2 million st) at about $44/t ($48.50/ st). Caulking and sealant may be highly filled with GCC yet undergo no adverse flow effects, with a narrow particle-size-distribution filler decreasing the binder demand. The CaCO3 industry, as well as the carpet industry, are more or less tied to the growth rates of the construction industry. It is estimated that the carpet industry uses some 680 kt/a (750,000 stpy) of GCC at about $25/t ($27.50/st). The paint industry uses approximately 300 kt/a (331,000 stpy) at a 1 % to 1.5% annual growth rate. Here, too, GCC is the dominant filler. It is used to enhance flow characteristics and color uniformity. It also extends costly titanium dioxide, creates sheen and controls roughness, hardness and tack. More CaCO3 should be used in the future as industry shifts from solvent-free or water-based formulations that can accommodate higher GCC volumes. CaCO3 is not imported or exported in any great quantity. Most areas have reserves of their own and the selling price is relatively low. Even so, quality varies from deposit to deposit. As our needs become morespecific, it remains a challenge to provide varying industries with a fitting product.

Jan 1, 1998

By K. Mohamed, G. Rashed

A conventional mechanical anchor bolt consists of a smooth bar threaded at the anchor end, with a conical wedge, and a mechanical shell anchor. When the smooth bar is torqued, the conical wedge pushes the shell anchor against the borehole wall. Coal mines use mechanical bolts in addition to other types of bolts to control the deformation and stabilize the yielded coal ribs. Limited research has been conducted to understand how a mechanical anchor performs in a coal rib. Researchers from the National Institute for Occupational Safety and Health (NIOSH) conducted the study described in this paper to define the loading characteristics of mechanical bolts (stiffness and capacity) installed in coal ribs at five underground coal mines. Standard pull-out tests were performed in this study to define the loading characteristics of a mechanical rib bolt. Based on the strength of the coal seam and the conditions of the coal mines, the bolts were installed at different amounts of torque. A typical tri-linear load-deformation response was obtained from these tests. It was found that the anchorage capacity depends on the coal strength, and the anchorage stiffness depends on the installation torque. The outcome of this research provides essential data for rib support design. INTRODUCTION Rib falls are a serious hazard in underground coal mines. In the last 10 years, rib failures have resulted in 17 fatalities, representing 52% of the ground-fall fatalities in underground coal mines in the United States (MSHA, 2018). Although resin bolts has been used increasingly in the last few years to control coal ribs, mechanical bolting (mainly expansion-shell bolts) is still used in many underground coal mines. Recently, NIOSH researchers surveyed 90 sites in 14 underground coal mines in the Eastern United States and found that 79% of the mines determine their rib support requirement based on the depth, mining height, and the loading conditions, of which 35% use the conventional mechanical bolts. Conventional mechanical bolts are preferable for rib support because they do not require the miners to be too close to the rib during rib bolting. Hence, using the conventional mechanical rib bolts can reduce the risk of injuries and fatalities during the installation process. The mechanical bolt or the point-anchored bolt is the oldest type of rock bolt that is still in use today. The point-anchored bolt is a tensioned bolt anchored down the hole with a mechanical expansion shell or a grouted medium (Peng and Tang, 1984). Figure 1 shows the components of the mechanical anchor bolt. The mechanical anchor bolt consists of a plug, a shell anchor, a headed smooth bar with the top end threaded, a bearing plate, a hardened washer, and sometimes a friction-reducing washer. The shell anchor can be either standard or bail type. The shell of the standard type is held in place by a nut, while the bail anchor uses a strap to hold the shell in place (Peng, 2008). The diameter of the smooth bar is generally 5/8 in for the #5 bolt or 6/8 in for the #6 bolt. The diameter of the hole that the bolt is inserted into must be carefully controlled since an oversized hole can result in poor anchorage (Moss, 1971). The anchorage is obtained by applying torque to the bolt head, which in turn pulls the wedge down in the shell and expands the serrated leaves against the sides of the hole-wall. In the resin-grouted anchor, the resin replaces the mechanical expansion shell at the end of the hole. The anchorage is achieved by bonding between the resin, bolt, and the hole-wall.

Jan 1, 2019

By M. Ge, C. Mborah

"The acoustic emission/microseismic monitoring (AE/MS) technique has widely been used in the mining environment for the monitoring of rock mass stability. A successful implementation of the method is hinged on being able to accurately determine the P- or S-wave arrival which is then used in the determination of the source of an event. Determination of the arrival phase is usually performed visually by a human expert. However, the mining environment is characterized by high noise levels and different noise types of varying intensities. Therefore, performing manual phase picking on time series data acquired from the AE/MS monitoring system can sometimes be tedious and time consuming if not impossible. In this paper, a proposal for improving manual P-wave phase arrival picks is discussed. The method involves the filtering of the AE/MS data using a stationary discrete wavelet transform (SDWT) method. After filtering, appropriate scales/levels of the multi-level decomposition process are then selected for further processing. The first step in the picking process after filtering, requires the computation of the signal power and the root mean square (RMS) using the signal amplitudes at each scale. The preliminary P-phase arrival time is picked using the results of the calculated power and RMS values respectively. The calculated RMS values are then superimposed on the calculated power values to determine the final arrival picks. If the arrival time picks obtained by both methods coincide or are within 5 to 10 sample points, the pick is considered reliable. The accuracy of the method was verified using AE/MS data from two datasets obtained from two separate underground mines. The results of the study showed that the picking method is reliable and robust. Based on the data studied, it can be concluded that the method will help reduce the time usually spent in processing AE/MS data from mines. INTRODUCTION A successful identification and location of the source of acoustic emission/microseismic (AE/MS) events is mainly dependent on the onset time of the P-wave phase. In many seismic networks, initial picks are performed by picking algorithms followed by manual picks by human experts. If the waveform signal-to-noise ratio (SNR) is low, the P-wave onset time picks by the human expert is generally reliable and accurate [1]. However, for waveforms with high SNR, picking of the P-wave onset could be a major challenge. Also, due to the volume of data acquired daily, manual processing could be stressful and time consuming [2]. These challenges continue to provide motivation for alternative means of providing reliable and efficient ways of improving manual phase picking. The AE/MS technique has been an intergral part of safety monitoring in mines [3] due to its capacity to identify zones of instability in a monitoring area. To obtain an accurate source of an AE/MS event however depends on the quality of the data used for P-wave phase picking [4, 5]. AE/MS data from mines are generally contaminated with excessive background noise generated by the different mining operations. The AE/MS signals of interest are mostly buried in background noise which makes it difficult to pick the P-wave phase arrivals [4]. Processing AE/MS data acquired in noisy environment remains a major challenge due to the weak nature of AE/MS signals [6]. As noted by [4] most of the data recorded in the mine environment are mainly due to events not related to AE/MS events of interest. The complexity of data in this type of environment makes the detection and picking of P-wave arrival by human experts a very difficult task."

Jan 1, 2017

By J. C. Agarwal

Major scientific discoveries cannot be programmed. New or frontier technology, however, can and should be planned. It is the vital force underlying the business strategy of a vigorous and profitable mining and metals company. Generally, a radical departure from existing technology is required to 1) Enhance the competitive value of ore reserves, i.e., to turn low grade or marginal reserves into competitive reserves. 2) To increase yield and decrease costs. 3) To decrease investment. 4) To increase mine life. 5) To improve product quality to meet market demands. The very definition of frontier technology implies that it is so different that the old plant, if already in use, will not be directly suitable to use the new technology. Therefore, a priori, we know that the commercialization of frontier technology would very likely involve new investment. The recommended technique of economic and technical analyses, I propose to call "Success Analysis.'' Simply stated, it forces us, by step- by-step procedure, to ask "what is needed to succeed in this particular project." Therefore, the first step is to identify what is meant by "success." The definition of "success" requires an agreement with top management or sponsors of the project as to what constitutes "success" in their minds as contrast to the technical people perception of "success." In general, top management would define success as meeting: 1) The technical objectives, i.e., technology being developed would produce the desired product at the acceptable quality. 2) Financial objectives, i.e., when commercialized, the project's capital cost, operating costs and return on investment are better than the projects they would normally fund. The return has to be better because new technology is inherently riskier. 3) Marketing objectives in that the project will produce product of acceptable quality in timely manner, i.e., scale-up to commercial plant would not cause undue delays in start-up and that large fix-up expenditure would not be required to protect the projected return on the investment. Many projects are started in the laboratory to develop "Frontier Technology" in mineral processing. Most of these efforts are never commercialized. There are many causes of these failures. The major causes are listed below: 1) Inadequate communication with the sponsors of the project so that project objectives are defined and agreed to by all. 2) Ill-defined technical problems. 3) Incomplete development of technology. This has two major components: a. Incorrect scale-up and engineering. b. Inadequate start-up and maintenance planning. 4) Incomplete and over-optimistic process economics. 5) Lack of understanding of the future behavior of the marketplace. 6) Failure to complete R&D effort by the time it is needed. 7) Incorrect projections of human resources required; that is, technical, managerial and operating. 8) Lack of realistic forecast of the expenditures required for commercialization of the R&D effort. These pitfalls can be avoided by comprehensive planning and the use of success analysis technique. The major elements in "Success Analysis" are: 1) Identification of the "need" for new technology. 2) Unbiased and complete scale-up engineering. 3) Acceptable economic feasibility with enough margin of error in the sale prices and operating costs to justify the inherent risks in frontier technology. 4) Timing for availability of investment capital and markets. Many failures in the development of frontier technology are preordained because the initial selection of the research project has been without adequate attention to the needs of the industry. Many very innovative hydrometallurgical processes for recovery of copper from concentrates were developed in the past twenty years. They have not adapted except to a very minor extent because they did not satisfy the success analysis criteria. The copper industry in North America needed a process to compete against high grade ores and to avoid

Jan 1, 1985

By Geoff W. Seward, Charles J. Maes

INTRODUCTION The work of the metallurgical engineer engaged in developing a leach, solvent extraction electro- winning process usually begins in the laboratory and generally proceeds via pilot plant testing before finally resulting in a commercial operation. The path leading to the development of the eventual process will involve a test program on leaching in order to derive information on the amenability of the copper-bearing material to acid dissolution, under various conditions. Commonly in North America the leach process of choice is either dump or heap leaching but it may also be in-situ leaching or the treatment of mine waste- waters and bleed streams. In the past, agitation and vat leaching have also been included in the list of options. Currently the economics of copper production dictate that only material of mineralogy and grade, capable of being leached without prior milling, will lead to a viable process. This is not necessarily true in other parts of the world where grades are higher and milling costs lower and therefore this paper con- fronts all the techniques available to the metallurgical engineer in terms of the evaluation of the solvent extraction process. Understanding the objectives of a laboratory and pilot plant evaluation of a solvent extraction route is important and these are described in some detail in this paper. Of at least equal importance is understanding limitations of such a test program and this aspect of process development is also discussed with reference to the kind of equipment and experimental practices that have been shown to produce results consistent with the objectives. WHERE TO BEGIN? At the inception of a project much information is developed on the variables of the leach process and indeed in some instances the type of leach that best suits the ore to be treated is the subject of intense investigation (Wadsworth 1983). The decision to treat the leach solution by solvent extraction focuses attention on the selection of the extractant most suited to the nature of the chemical composition of the solution. A laboratory screening program will probably need to be arranged in order to assist in narrowing the options and the following is suggested for this initial screening. Selection of a Representative Leach Solution This is preferably obtained from leaching a representative sample of the ore under conditions likely to pertain in the commercial leach circuit. As the conceptualized circuit flow sheet will undoubtedly involve the recycle of aqueous solutions via the leach, then it is desirable to know the nature of and the likely level of the impurity ions in the solution as well as the concentration of copper. The other most important parameter that is required is the free acid (Anon, Acorga Analytical Method) content of the solution. As a rule of thumb, it is always true that recovery across SX is maximized as the pH of the leach solution rises. Whilst a high pH may be desirable for SX recovery, it is not always desirable for the pH (or more correctly, the free acid) of the leach to rise too high. The working range, therefore, tends to be between 1.5-2.5 pH. At an early stage in process development it is likely that a number of ill-defined conditions will exist, making it difficult to decide on typical leach conditions in which case the screening tests should be carried out using a synthetic leach solution prepared from laboratory chemicals. In doing this the copper and free content should be carefully controlled along with the appropriate addition of major impurity ions as their sulphate salts. Preparation of Suitable Organic Solutions Manufacturers information sheets on the reagents to be tested are required at this stage The information necessary to make an informed judgment on reagent performance is: Copper Uptake Value. This will indicate the amount of copper that can be extracted by a reagent as a function of its operating concentration in a

Jan 1, 1986

By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation's energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988

By Earl G. Hoover, V. S. Znamensky

For most geologists and miners the term asbestos refers to a group of highly fibrous silicate minerals that readily separate into long, thin, strong fibers of sufficient flexibility to be woven; these fibers are heat resistant and chemically inert, and possess a high electric insulation and therefore are suitable for uses where incombustible, nonconductive or chemically resistant material is required. The US Mining Safety and Health Administration (MSHA) specifically refers to the naturally occurring minerals chrysotile, amosite, crocidolite, termolite asbestos, anthophyllite asbestos, and actinolite asbestos as asbestos if the individual crystal fragments have the following dimensions: length-greater than 5 µm; maximum diameter-less than 5 µm; and a length to diameter ratio of 3 or greater. Since asbestos has been identified as one of the ten most carcinogenic substances (a cancer-causing agent), it has become, for many government regulatory agencies, a major issue concerning banning emissions to the environment. Very few recent issues have caused as much concern for the crushed stone industry as asbestos. Complicating the problem is the fact that several federal and state bureaus have greatly compounded the confusion because each has become involved in the asbestos dilemma. Simple solutions to a very complex problem have been proposed by these various governmental groups. In addition to governmental bungling is the fact that public interest groups have been able to use federal and state laws that have been enacted with a strong environmental bias to foment even greater complexities. Such is now the state of the asbestos problem. No con- firmed health risk has been associated with low-level exposure, namely crushed stone in an unbound form; a methodology for sampling and analyzing for asbestos has not been developed; a precise definition of what is asbestos does not exist, nor does a concensus among the various government groups as to what course of action to pursue. Pure asbestos as it is mined for commercial purposes does not exist in the properties mined by the crushed stone industry. Where the six asbestiform minerals do appear (to date we've been referring mainly to chrysotile in serpentinite formations in Maryland) the amount of the asbestiform mineral present is generally quite small. They generally appear in scattered, miniscule amounts. In the case of Rockville Crushed Stone in Montgomery county, Mary- land, the chrysotile is less than 1/2 of 1% of the rock deposit. No health risk has been identified with such minute quantities of asbestos minerals. MSHA standards for the occupational health and safety of quarry workers are being met; and there is every indication that normal background levels of asbestos, throughout the surrounding countryside, are as high or higher than levels measured on roads where the rock is used in pavement or base. The problem and the controversy lie with improper terminology on the part of so-called public interest environmental groups, aided and abetted by some governmental agencies who, to be charitable, we can best describe as uninformed. The news media frequently has fanned the issue, unintentionally we'd like to think, by quoting these misinformed authorities and not bothering to question the facts or adequately report balancing information handed to them. It is much better copy to talk about an asbestos cancer scare, than to deal with rather mundane measurements and descriptions that are hard to understand. At the heart of the issue is the improper use of the word asbestos. Instead of being confined to the geological description of the six asbestiform minerals-chrysotile, amosite, crocidolite, termolite asbestos, anthophyllite asbestos, and actinolite asbestos-the word asbestos is also wrongly used, but too-frequently to identify many other minerals that happen to have elon-

Jan 1, 1986

By E. Steel, V. S. Znamensky, A. Wylie

The asbestiform habit is most commonly developed in certain amphiboles and chrysotile, but other minerals also may crystallize with this unusual habit. The habit may be characterized by (1) a fibril structure, single or twinned crystals of very small widths (generally less than 0.5 pm), which have grown with a common fiber axis direction, but are disoriented in the other crystallographic directions; (2) anomalous optical properties, primarily parallel extinction; (3) unusual tensile strength; (4) high aspect ratio; and (51 flexibility. In addition, there is evidence to indicate that some amphibole asbestos may have unusual surface properties. ASBESTOS AND THE ASBESTIFORM HABIT Asbestos is defined as a group of highly fibrous silicate minerals that readily separate into long, thin, strong fibers of sufficient flexibility to be woven, are heat resistant and chemically inert, and possess a high electric insulation, and therefore are suitable for uses where in- combustible, nonconducting, or chemically resistant material is required (Gary, et al., 1974). The most common minerals that may occur with the asbestiform habit are chrysotile, grunerite (amosite], riebeckite (crocidolite), actinolite, anthophyllite, and tremolite; although the development of this habit among these minerals is rare. Other minerals, most notably other amphiboles, can occur in this habit, but no others have been mined commercially as asbestos. All asbestos is confined to metamorphic rocks, even though other habits of the amphiboles are common in igneous rocks. Slip fiber veins are the most common deposits (South African amosite, Canadian chrysotile, etc.), but mass fiber deposits also may occur such as California chrysotile or mountain leather. Field relations and experimental data support the hypothesis that metasomatism is the dominant process in the formation of asbestos fibers, amphibole as well as serpentine. The crystal habit of a mineral is the shape or form a crystal or aggregate of crystals take on during crystallization. The asbestiform habit has a number of characteristics that differentiate it from other habits. Chief among these is the fibril structure.Afibri1 is a single or twinned crystal with a very small width, generally less than 0.5 pm, and an extremely high aspect ratio; bundles of fibrils may have lengths reaching into the cm. Fibrils share a common crystal growth direction along the long axis of the fiber, but appear to be disoriented with respect to one another in the other crystallographic directions. The structure of the individual fibrils, and the organization of fibrils within a fiber may differ among the various types of asbestos. The fibril structure of asbestos is probably a factor in controlling a number of secondary properties that include high tensile strength, flexibility sufficient for weaving, and anomalous optical properties. The high tensile strength of asbestos is quite remarkable, exceeding that of the ordinary varieties by a thousand fold (Zoltai, 1978). In part, this may be attributed to the lack of defect on the fibril surfaces (Zoltai, 19781, but some contribution also must come from the existance of bundles and the nature of the ordering of and forces among the fibrils. The flexibility is not only enhanced by the fact that the fibrils may slip past one another, but by their small widths and extreme aspect ratios. However, structural defects parallel to the fiber axis also may contribute to this property. Among the common commercial asbestos types, the greatest flexibility is found in chrysotile and the least in grunerite asbestos (amosite), varying inversely with fibril thick- ness. In addition, at least for chrysotile, composition also may affect the flexibility of the fibers. Finally, most asbestos minerals are mono- clinic, but optically display parallel extinction (Heinrich, 1965, and Wylie, 1978). This also must be due, in part, to the ordering or lack of ordering of the fibrils. The fibril structure is probably a random

Jan 1, 1986

By Jasinder S. Jaspal, Bradley J. Bourquin

Introduction Operation is underway at the Dutch Creek No. 1 mine of the first US longwall panel to be mined under the cooperative agreement between the US Bu¬reau of Mines and Mid-Continent Resources. The 244-m (800-ft) wide longwall face is part of a multilift longwall demonstration. A 3-m (10-ft) top slice is being extracted from the 8.5-m (28-ft) coal seam. This occurs in an area of the property where the Coal Basin A and Coal Basin B beds have come together to form one seam. This face, longest ever operated in the US, is being worked on an advancing longwall system. During the first months of its operation, this longwall face has been very productive and safe. Mid-Continent Resources operates two coal mines in the Coal Basin, located in the White River National Forest near Carbondale, CO. The Colorado Fuel and Iron Co. originally opened this property in 1895, and conducted a small mining operation for 10 years on the north rim of the basin and then abruptly shut down. Mid-Continent began operations in 1956, after the Coal Basin fields had lain idle for about 60 years. Operations commenced at the outcrop of the Coal Basin B bed situated at the 3-km (10,000-ft) elevation. At this time, coal was produced from Coal Basin A and Coal Basin B seams in the area. Since 1981, operations have been centralized at two mines, the Dutch Creek No. 1 and the Dutch Creek No. 2, both located in the center of the basin. Since 1981, Mid-Continent has shifted its emphasis to longwall mining as the primary method of production, and each of these mines is equipped with state-of-the-art longwalls using shields as the method of roof support. A 4-km (13,000-ft) long twin-bore rock tunnel is being driven at a lower elevation of 2.4 km (8,000 ft) to intersect the coal seams at depth, as they dip 12° in a westerly direction. This tunnel will enable the company to reduce costs by replacing trucks with conveyor belt haulage (Fig. 1). Mid-Continent began longwall operations in May 1976 with the US's first modern, mechanized advancing longwall face, using chocks. In August 1978, a second advancing longwall face was started. Results from two faces showed that the mining system could be used effectively to manage the difficult conditions encountered in both the Dutch Creek No. 1 and the L. S. Wood No. 3 mines. Unfortunately, low productivity and high costs plagued the company's advancing faces. The problems resulted principally from major equipment failures, poorly designed chocks and face conveyors, and the lack of an efficient, cost effective system for building packwalls. Flushing of rock between chocks and roof control were also major problems. In Sept. 1979, Mid-Continent and the US Bureau of Mines entered into a cooperative agreement to demonstrate the feasibility of mining thick coal seams using the multilift longwall mining method. Two lifts are to be used to mine the 8.5-m (28-ft) thick coal seam in Dutch Creek No. 1 mine. The Bureau provided mining equipment for 180 m (600 ft) of longwall face. This included the shields, face conveyor and drive units, shearing machines, panel belt, and stageloader. As a result of previous experience in advancing longwall mining, Mid-Continent purchased an additional 60 m (200 ft) of face, including roof supports and an additional face conveyor. Design The Dutch Creek No. 1 is a slope mine located on the flank of a plunging anticlinal structure that

Jan 1, 1984

By Cong Shi, Cuiping Li, Enji Sun

"Tailings dam is a subsidiary facility of mining production. In order to decrease tailings dam accidents caused by the failure of flood drainage system, this paper proposed a framework of potential failure detection system of tailings dam flood drainage facility based on three dimensional panorama technology. A collection device of three dimensional panorama is designed to capture the onsite images of flood drainage system of tailings dam. The spatial information of flood drainage facility of tailings dam is recovered through projecting the image information to spherical surface by using three dimensional panorama technology. Two dimensional images are transformed into three dimensional images in three stages of panorama images process, which are pretreatment, registration and splicing. Combined with the framework of HTML5, panorama images are embedded into webpage. These web pages have the functions of user interface that the images could be displayed, moved and zoomed. With the function of interaction with panoramic display, 3D panorama scenes can be changed and scaled. Browsing the webpage is helpful to find potential risks such as possible holes, fractures and leakages of flood drainage system of tailings dam. The realization of this system can detect the potential safety risks of flood drainage system of tailings dam. The application of this system would improve the safety and provide the ability of accident pre-warning. It is useful to maintain the healthy and sustainable operation of tailings dam.Introduction It is reported by Xie et al (2009) that tailings dam are necessary facilities for mine production and places which constructed by damming troughs or encirclement used to store tailings that eliminated after ore selection in Metal and non-metallic mines. It is reported by Mei (2012) that tailings dam is a dangerous source of high-energy man-made debris flow, the collapse of the tailings dam takes devastating damage, which seriously threatens the safety of life and property of the enterprises and the downstream residents. It is reported by Dong (2011) that the collapse of the tailings dam causes secondary disasters such as environmental pollution at the same time. It’s a serious impact on regional economic development and social stability. The phenomenon of collapse of destruction is general excessively of tailings dams in China. It is reported by Zhang et al (2011) that most of these phenomena are caused by flood over the top of dams. The main reason is the lack of capacity or the failure of the drainage system. There are also accidents caused by the failure of drainage system in foreign countries in recent years. For example, November 5th 2015, the tailing dams collapsed because of the insufficient drainage capacity of the Germano mining area in Germano Mariana area in central Minas Gerais state in Brazil. After a small post-earthquake, the liquefied tailings fill sand submerge downstream cities and towns. It finally destroyed 158 houses, killed at least 17 people, polluted 663km river and destroyed 15 square kilometres of land along the river. It is reported by Rico et al (2008) that it’s necessary to continuous monitor and control the tailings drainage system in order to reduce the occurrence of tailings accidents and guarantee the sustainable production of the mine. The traditional artificial detection methods of flood drainage system of tailings dam exist many problems such as heavy workload, low efficiency, and susceptible to meteorological conditions. And it is reported by Luo (2013) that the tailings dam wastewater is corrosive which result a certain disease in flood discharge facilities such as tailings drainage culvert, etc. In order to realize the effective monitoring, management and maintenance of flood drainage system of tailings dam, advanced technologies need to be used to acquire the information of collection and management of flood drainage facilities of tailings dam."

Jan 1, 2017

By David M. Stearns

Introduction Millions of tons of coal are shipped by rail each year in the US. Loading those trains efficiently is a topic being studied by coal producers and railroads. Alternatives range from volumetric loadouts to systems that weigh the product, using either belt conveyor scales, track scales, or weighbins. Each system has its proper application and each has pros and cons, depending on the volume, nature, and layout of the specific loading facility. In recent years, a batch weighing system has been developed that offers significant benefits to higher volume operations shipping by a unit train. The Unitrain Loadout System (ULS) features a precision electronic batch weighing system married to a mechanical loadout system by a microprocessor-based digital control system. A batch weighing loadout could be designed using analog electronics or even a mechanical weighbin. But it is the adaption and use of a microprocessor that provides the primary benefits of a unit train loading facility. Those benefits are accurate weight determination, optimum car use, overload prevention, speed in loading, optimum labor use, and documentation of the loading process. Unitrain History The first ULS was erected for iron ore applications in Canada during the 1960s. The first domestic system on coal was installed at Pittsburg and Midway Coal Co.'s McKinley mine at Gallup, NM, in 1978. At present, there are more than 20 Unitrain Loadout Systems in operation worldwide. They are loading minerals, concentrates, and coal. Eleven systems are presently in service or under construction in the US. In the last several years, a number of comparable systems have also been installed. As a result, the concept has been well proven and is now commercially available from several sources. Unitrain Operation The basic operation of a ULS involves loading a unit train with individually weighed batches of coal while it moves slowly under a loadout tower. Typical coal systems are designed to load at rates of 2.7 to 6.3 kt/h (3000 to 7000 stph). The contents of each car can be accurately weighed. This meets the requirements of the National Bureau of Standards Handbook No. 44 for static weighing. Also, cars can be loaded within very close tolerances. This optimizes railcar usage and prevents overloads. Principal components of the loadout system are the main feed conveyor, the surge bin, surge bin gates, weighbin, weighbin gate, flood loading chute, control room, calibration weights, and hydraulic and electrical systems (Fig. 1). Main feed conveyor: Coal is delivered from a storage area by the main feed conveyor. The conveyor fills the surge bin. It should be capable of conveying coal at the same rate at which it is being loaded. Usually the feed conveyor is equipped with a conveyor scale that can display rate in the control room and control feed to the belt. The surge bin will typically store 227 to 272 t (250 to 300 st) of coal. If the loadout is installed in an over-the-track silo, there is no need for a surge bin. Four double-bladed, hydraulically actuated gates are incorporated into the bottom of the surge bin. These gates open to fill the weighbin. Automatic pre-act points are selected by the control system. They shut these gates as the desired weight is approached. The final amount of coal is added by a single gate that closes. This results in a weighed batch of coal within ± 0.5% of the selected optimum load per railcar. Thus, if loading 91-t (100-st) capacity cars, the system is set to batch up 90.2 t (99.5 st) batches. Each car is expected to be loaded between 90 and 91 t (99 and 100 st). Weighbin: The weighbin is next in the flow scheme. Typically, it is

Jan 3, 1985

By Robert W. Buechley, Charles R. Key, Jonathan M. Samet, Marion V. Morgan, Arnolfo A. Valdivia

INTRODUCTION The Uranium Epidemiology Study at the University of New Mexico was initiated in 1977 with the objective of assessing health effects of uranium mining in the Grants, New Mexico, area. The planned approach was a cohort study of mortality that would capitalize on the availability of health records at the Grants Clinic. During the first year, emphasis was placed on determining the feasibility of the proposed study: Were adequate records available? What was the quality of available information? Could a cohort be established? The first year's work confirmed the study's feasibility (Buechley, 1979). Subsequent efforts have focused on the health and exposure data base and comprehensive follow-up. A retrospective cohort study of early miners is in progress; a prospective cohort study of miners who began working under the current Federal standard is planned. This report will summarize the current status of the Uranium Epidemiology Study. A chronology of the study's activities follows (Table 1). TABLE 1. Uranium Epidemiology Study Timetable 12/77 Official start of study. 12/77-8/78 Abstracted Grants Clinic records of 21,780 persons examined for mining companies from 7/57 through 12/76. Checked company personnel records to establish who had worked underground 1 year or more through 12/76. 7/78- Obtained physical exam and mining history information on qualified miners from Grants Clinic and companies. 9/78- Obtained working level month information from companies and other sources. 12/78- Checked qualified miners list with New Mexico death tapes and New Mexico Tumor Registry files. 9/80- Submitted qualified miners list to Social Security Administration for search of its records. 12/80- Obtained copies or abstracts of early State Health Department and Mine Inspector's reports. 3/81-4/81 Conducted Pilot Respiratory Disease Survey. 3/81- Verified underground status of qualified miners via company personnel records. 3/81- Verified smoking and mining histories of qualified miners via Grants Clinic records. 3/81- Began establishing cohort with first underground experience after 1970. 4/81- Checked qualified miners list with Post Office, Voter Registration Office, Motor Vehicle Division, telephone directories, city directories, and other sources. BACKGROUND Uranium Mining in New Mexico The Grants, New Mexico, uranium belt measures some 110 by 25 miles in Cibola, McKinley, and Sandoval counties (Kelley, 1963). The three chief mining districts are, from east to west, Laguna, Grants, and Gallup. Average annual production was about 5,500 tons, or 45% of United States uranium, from 1966 through 1978. Despite a drop in this figure in 1979, to 40%, New Mexico still leads the nation in uranium reserves and production (Hatchell, 1981). In 1918 carnotite, a yellow uranium vanadate, was identified in the Shiprock area (Chenoweth and Learned, 1980; N.M. Energy and Minerals Department, 1981). It received little attention at the time because of its apparently limited volume. In 1950 an important uranium outcropping was discovered by chance in limestone at Haystack Butte, northwest of Grants. In 1951 deposits were discovered in sandstone at Poison Canyon, north of Grants, and at Jackpile, northeast of Laguna. In 1955 finding of radioactive cuttings at an abandoned oil test site near Ambrosia Lake led to drilling and discovery of the first underground deposits (Melancon, 1963). Ore bodies were also found near Church Rock in the 1950's (Rautman, 1977). A uranium rush followed. At various times during 1952-1962, 60 mines and 5 mills were in operation (Gay, 1963). In the early 1960's the price of uranium fell, and many miners left the area. Subsequent market fluctuations have produced shutdowns and consolidations, development and expansion. Many early uranium miners came to New Mexico with

Jan 1, 1981

By I. S. Parrish

Locating and defining an ore body with minerals or metals that can be recovered at a profit is the objective of minerals exploration. Direct exploration involves examining and sampling rocks either in outcrop, underground exposure, drill core, or cuttings. Indirect exploration infers the measurement of some physical, chemical, or biological factor that reflects the possible presence of the metal or mineral of interest, or a structure that may act to collect, concentrate, or otherwise favor the presence of the desired commodity. Included here are all forms of geophysical, geochemical, or geotech-nical prospecting. In recent years, indirect exploration has given rise to an impressive number of specialists and a variety of techniques. Less impressive is the success rate or productivity of the modern explorationist. Exploration Productivity Losing Ground Judging from the low number of mine openings or mine discoveries in recent years, the exploration industry is not as productive as it once was. Many will take umbrage at this judgment, but a review of the literature will support this statement. In Canada, from 1951 through 1960, there was an average of 5.6 discoveries a year. From 1961 through 1976 there were only four a year. This nearly 30% reduction occurred despite a massive increase in exploration funding. From 1950 to 1955, the Canadian mining industry spent $5 million (1975 dollars) for each minable deposit found. During the period 1970 to 1975, the cost was $25 million for each one found. The industry was spending five times as much and finding 30% less than a generation earlier. This is not indicative of a productive endeavor. The loss in productivity is even more apparent to explorationists who wander the western US, and are awed by the thoroughness and discovery rate of those early prospectors who explored from 1850-1920. The "old-timers" missed very little. They used pans and crude chemical tests in place of plasma or X-ray fluorescent assay techniques. But, they found ore bodies that led to mines. Today, the best prospecting areas are pockmarked by the hand dug pits, trenches, and shafts left by our predecessors. The success or productivity of the old-timers may be attributed to any or all of several factors. Granted, they were looking at virgin ground and at deposits that in many places were exposed. The factor most critical is, however, that they employed direct exploration methods. They walked over areas of interest, traversed swamps and mountains, and followed "colors" in stream gravels to their source. They also chipped rocks, dug pits, trenched along structures, and followed the "leads." Many of today's explorationists, on the other hand, adopt a more sophisticated approach. A geologic model is conjured up so the exploration team "knows" models, but the model should be applied at a later stage in the exploration effort. J. D. Ridge has written an excellent article on the role and timing of models. Geologists should collect and assemble data, facts, and observations and fit these into a model and not conceive a model and seek data to prove their vision. Further, many exploration managers are adept and talented at devising exploration programs. They can program and use modern exploration techniques and tools. They can apply these and their modeling capacity to assemble a glittering array of targets. It would appear that in all too many instances, creating targets has been equated with and is considered being productive. As noted, finding ore bodies is being productive in exploration. Creating targets may or may not lead to ore bodies. As an end result, locating targets is not productive exploration. How then does one foster productivity in exploration? Some suggestions are to increase the use of direct methods, follow the methods and philosophies of old-timers, look at more rocks, do more field work, use foot traverses, and do more drilling. It may be argued that this is what the old-timers have done and we would then be hoping to find

Jan 11, 1983

By James Douglas

For the past several years, inflation has placed increasing pressure on our economic system. The total effect of inflation is complex when combined with other factors which affect the economics of equipment. A mathematical model is described herein which can be programmed on a computer to determine the effect of inflation on economic life and profits after taxes. The computer output results in graphic plots which clearly show the relation of various percentages of inflation to life and profits. Inflation has become a common topic of conversation in the recent past. What is it and how can it be measured? What effect does increased inflation have on equipment policy? Research now in progress at the Construction Institute at Stanford University is beginning to unravel some of the mystery of inflation. As a start, let us answer the first question by defining the word inflation. The dictionary says that inflation is "an increase in the volume of money and credit relative to available goods resulting in a substantial and continuing rise in the general price level. It appears, in fact, that we are in an inflationary spiral which is "a continuous rise in prices that is sustained by the tendency of wage increases and cost increases to act on each other." 3 Inflation behaves just like compound interest. It takes more dollars of future years to buy an equal amount of a given commodity, be it a material or a service. Let us assume an annual inflation rate of 5 percent. Assume that 1Associate Professor of Civil Engineering, Stanford University, Stanford, California. 2Webster's Seventh New Collegiate Dictionary, Springfield, Mass., G. and C. Merriam Co., 1969. 3Ibid.

Jan 1, 1972

By Michael Gay, W. H. Hansmire, G. Rippentrop, V. S. Romero

The new Tren Urbano heavy rail transit project in San Juan, Puerto Rico has one very significant part, the Río Piedras Contract, being constructed underground. This paper will give an overview of the tunneling work that started construction in1997. Design-Build procurement for tunneling is addressed. The new and modern transit system in San Juan will be 17.2 km (10.7 mi)long and have 16 stations in its first segment of construction, which is due to open in 2001. The line connects the city center district of Santurce to the outlying suburban municipality of Bayamón. All construction is being done under design-build contracts as a demonstration project funded by the Commonwealth of Puerto Rico and the Federal Transit Administration (FTA). The Río Piedras Contract is completely underground and involves practically all types of construction in soft ground. One station is built as cut-and-cover at the University of Puerto Rico. One portion has twin tunnels to be driven with an Earth Pressure Balance Machine (EPBM) and lined with one-pass precast concrete segments. Another has four short tunnel drives at a turnout that will be constructed with an initial support of shotcrete and lattice girders (i.e., the New Austrian Tunneling Method of construction). The most complex construction is the Río Piedras Station, which is being built as a stacked drift with a total of 15 individual tunnel drifts. After being sequentially excavated and concreted, they will form the arch of the 19 m (62 ft.) wide by 16 m (53 ft.) high cavern station. This structure is excavated in soil with less than 5 m (16 ft.) of cover to overlying historic buildings.

Jan 1, 1999

By J. C. Irvine

Meramec Mining Company is a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp. which mines and pelletizes iron ore at the Pea Ridge Mine. The Pea Ridge property is located near Sullivan, Missouri, about 70 miles southwest of St. Louis. The orebody was delineated in the mid-1950's by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drilling in 1956 and 1957 confirmed that the orebody was mineable, Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Production commenced in April 1964. The orebody is overlain by about 1300 feet of flat-bedded sediments. It is tabular, about 2600 feet long and up to 600 feet thick. It dips about 80° and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 150 foot intervals. Crosscuts were driven across the orebody on 190 foot centers. Banks of scopes were mined between the crosscuts by a modified shrinkage stoping method. This was done by undercutting a 40 foot wide by 150 foot long block and blasting horizontal "lifts" drilled on 5 foot intervals. A 65 foot still was left between levels which contained the slushing drift, fingers, crown pillar, and adequate thickness for support. By 1971 a ?lattice? of pillars had been left and mining had progressed to the point that an orderly pillar recovery program was necessary.

Jan 1, 1976