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By E. D. Attanasi, J. H. DeYoung

Although fraught with problems of completeness and comparability, statistical measures of mineral exploration in the United States point downward for 1985 and 1986. Exploration expenditures in the US domestic and Canadian companies have dropped steadily from a 1981 peak of about $400 million to about $190 million in 1985, according to a recent survey of exploration statistics by the Society of Economic Geologists (SEG). These totals have been adjusted to account for differences in the estimated coverage of exploration activity by each year's SEG survey. This survey is now supported by the US Bureau of Mines and the US Geological Survey (USGS). The drop in current dollar amounts does not take into account the effects of inflation, which would make the decline even more substantial. Metals Economics Group of Boulder, CO reported in its Mine Development Bimonthly that announced exploration budgets for 1986 indicated an even larger drop for the past year. Worldwide expenditures for companies covered by the report dropped from about $800 million to $750 million from 1984 to 1985 and plummeted to about $600 million in 1986. Exploration activity has been declining since about 1981. This has been due in large part to low prices for base and ferrous metals. Exploration budgets of mining subsidiaries of oil companies have plunged as parent companies try to conserve cash in the face of the sharpest decline in oil prices in nearly 40 years. Mining subsidiaries of petroleum firms that were either liquidated or divested recently include Anaconda (Arco), Getty Mining (Texaco), and Cyprus Minerals (Amoco). Exploration expenditures in Canada increased in 1984 in regions where gold is a traditional target. But there was little change in 1985, based on preliminary statistics from a federal government report (Energy, Mines and Resources Mineral Bulletin MR 211, 1986). Claim staking in Canada dropped 23% from 1983 to 1984 and declined slightly in 1985. Diamond drilling increased in 1984 to the record levels of 1980 and 1981. The ratio of exploration to development expenditures in Canada, however, fell from a 1981 peak of 0.9 to about 0.5 in 1984. The report also indicated that, since 1981, there has been a massive write-off of reserves with high production costs. Reserve levels of all major metals were down since 1981. Precious metals exploration accounts for an increasing share of total mineral exploration. For example, the SEG data show that 90% of 1985 US exploration expenses by domestic mining companies were devoted to base and precious metals, compared with only 51% in 1980. The restructuring of the mining industry has been accompanied by a decrease in response to the last two SEG annual surveys, but the trend towards precious metals was already established in earlier years. Trade journal reports suggest that the reduction of long-term interest rates has encouraged the mining industry to make cost-reducing capital investments in existing mines to maintain competitiveness. This is consistent with the observations that metals explorations in 1986 was characterized not by announcements of "grassroots" discoveries, but by on-property exploration on identified mines or prospects. In an effort to cut costs, gold mining companies have been following up on low cost production opportunities in their prospect inventories. Short-term, low risk, heap leach operations that can produce rapid returns on investment have been pursued. Specific exploration projects are covered in the individual state sections of this Annual Review. Even though economic growth continued steadily through 1986, the domestic mining industry has

Jan 5, 1987

By Frank Florez

INTRODUCTION The Magma mine, located at Superior, AZ, some 100 km (60 miles) east of Phoenix, was discovered in 1875, and was originally mined for silver. The property was acquired by W. B. Thompson who formed the Magma Copper Co. in 1910. Prior to 1965, square-set mining and cut-and-fill methods were used to mine steeply dipping veins at Magma. Ore was moved to the chute by gravity in the square-set stopes and by slusher in the cut-and-fill stopes. The stopes were gobbed with waste from the surface and development headings in the mine. The original square-set method of mining was modi¬fied in order to extract replacement ore bodies in limestones that dipped at 0.52 rad (30°) and varied in thick¬ness from 4.6 to 38 m (15 to 125 ft). Mining of these ore bodies was done by starting at the upper lateral ex¬tremities and retreating to the access openings. The timber square sets and cribbing were used as temporary support, the ground being allowed to cave as the mining retreated. It was found, however, that mining in such a manner was very inefficient; safety hazards increased be¬cause of the blocky and fractured characteristics of the ore and the weakness of the hanging wall. GENERAL DESCRIPTION The undercut-and-fill method of mining is now the primary stoping method used at Magma, and no pillars are left for later extraction. Four ore bodies are pres¬ently being mined. These ore bodies (see Fig. 1) dip conformably at 0.52 rad (30°) to the east with the host limestone and shale sediments, their average strike length is about 122 m (400 ft), and their thickness varies from 9 to 38 m (30 to 125 ft). The ore bodies are separated by varying thicknesses of barren rock. Three of them come together at various elevations primarily in the area of the main feeder fault, but there is an apparent horizontal displacement of about 23 m (75 ft) on the main feeder fault which complicates stope layout. The method as originally applied has been modified in order to mine more efficiently ore bodies that are 9 m (30 ft) or more in thickness. The modification was made to mine all the ore at a given elevation before dropping down for the cut below. The advantage of this is that mining is done on a horizontal plane and there¬fore muck removal can be mechanized. Also multiple faces are made available, thereby improving overall min¬ing efficiency. Another major change made was to carry a 6.7-m (22-ft) face when mining the top sill and a 6.4-m (21-ft) face when mining under fill rather than the former standard 3.35-m (11-ft) high face. This change lengthens the mining cycle and cuts mat prepa¬ration costs in half. Stopes are laid out (see Fig. 2) so that maximum mining panels are available and yet one stope crew will not tie up too much of the ore reserve. One stope might consist of as many as 20 panels, and these panels are divided evenly on opposite sides of the extraction cross¬cut. The stope layout is planned so that the maximum tramming distance to the orepass will not exceed 30.5 m (100 ft) on the crosscut and 38 m (125 ft) in the panel. Stope panel distances are sometimes exceeded when min¬ing to the extremity of the ore body. Tramming distance limits are determined by the efficiency of the com

Jan 1, 1982

By S. D. Thompson, L. Adler

J.R. Sturgul Thompson and Adler (1988) present a new simulator for use in designing conveyor belt systems for underground mines. Their model uses a large Fortran program to do a Monte Carlo simulation. Their approach is similar to that done by other researchers, such as Stanford (1965) and Newhart (1977). All Fortran based simulation models use an imaginary clock that is advanced a time increment. At each advance of time increment the program has to determine what is to take place in the system. Since many things can take place at different places in the system, this approach requires that a large amount of logic is needed at each step of the program. This, then, necessitates a great amount of jumping between different portions of the program. In fact, the Fortran program, CMBCS, takes up to 15 to 20 minutes to run on the mainframe and several hours for the PC. Such execution times are typical of Fortran based simulation studies. There have been other examples of Fortran based simulation studies in mining. Cross and Williamson (1969) presented a model of a working open-pit copper mine. Fosinato (1982) presented a model of the bucketwheel excavators for the large open-pit mines in Southern Victoria in Australia. These excavators discharge into a complex conveyor system, which can be thought of as being similar, but more complicated than the one modeled by CMBCS. All previous simulation studies using Fortran as the source code have several things in common: they all are very lengthy, often being of the order of 10,000 lines of code and take a great deal of CPU time to execute. The program of Fosinato takes more than 4.5 hours to run on a large mainframe. Also not mentioned by Thompson and Adler is the actual time it took to write the Fortran code. This, too, can be considerable, as Fosinato's program took several man years to write and de-bug. In order to change it, as the parameters change or as the mine increases, further large amount of time is needed. An alternate to such Fortran based programs is provided by using a special computer language known as General Purpose Simulation System (GPSS). This language is designed to solve a wide variety of discrete event simulation problems and appears to be particularly suited for mining situations as shown by Sturgul and Yi (1987). Although GPSS was introduced as far back as 1961, it has not been used in mining to any great extent. However, the original form was quite crude and inefficient compared to modern versions. Perhaps with the increasing popularity and availability of these recent versions, this will change. Two such versions are GPSS/H for the mainframe and GPSS/PC for the personal computer. GPSS/H allows for interactive usage, such as allowing input data to be entered via a menu. GPSS/PC has excellent graphical capabilities and one can actually view the simulation in progress. In addition, one can halt the simulation and change values of various parameters. If a particular portion of the simulation is in progress via a concept known as "microwindows." These new versions have greatly increased the potential for using the language for mining situations. Mining applications of GPSS have been given by Sturgul and Yi (1987), Sturgul and Harrison (1987), Sturgul (1987), and Sturgul and Singhal (1988). Schriber (1987) presents a summary of the recent advances in GPSS.

Jan 12, 1988

By D. M. Range

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

Jan 1, 1989

By Peter Joralemon

Introduction The Sierra Nevada Foothill Gold Belt in California extends from Mariposa north for more than 420 km (200 miles). It includes the Grass Valley and Alleghany districts in the zone of Northern Mines and the famous Mother Lode to the south. The gold belt has produced more than 3.1 kt (100 million oz) of gold since its discovery in 1848. This remarkable production, at the present gold price, would be valued at more than $40 billion. The Law of Gambler's Ruin supposes that averages are based on systematic rather than random variations. It emphasizes the vast differences between finite and infinite probability. By applying this Law to this figure, it can be seen that the belt has produced an average of 15.5 t (500,000 oz) of gold per running mile. Many geologists believe, with some justification, that less than half the gold in this belt has been mined. Sixty per cent of the past production has come from lode deposits, the balance from Tertiary and Recent placer gravels. This paper will concern itself with the developing lode mines along the Sierra Nevada Foothill Gold Belt. As a further and final note regarding placer mining along this belt, Mark Twain suggested that placer miners, having lost their senses, had contributed liberally to the population of California's insane asylums. After a 40-year lull with little or no gold production, the lode mines of this famous belt are showing signs of renewed activity. Two open pit mines, the Sonora (Jamestown) and Carson Hill, are now in operation at substantial scales. Two others, the Goldenbell (Pine Tree-Josephine) and the Royal-Mountain King, are being readied for production on equally impressive scales. All are in the southern part of the Mother Lode. Exploration farther north, at Sutter Creek, Drytown, and Georgetown, is showing encouraging results in developing deeper lode deposits. This new activity has been sponsored to a large extent by junior Canadian mining and exploration companies. Most of the established American mining companies have looked the other way, not participating in this important development. One major reason for domestic disinterest in this opportunity is there is little chance for heap leachable open pit ore in this belt, excluding Carson Hill. And heap leaching has become fashionable with American mining companies today. Also, this great old gold belt must have been worked out, for why else is there so little activity here - Catch 22 logic. Another reason is that if a new deposit of size and grade similar to the mines in the past was found, high costs would preclude a profitable operation. And, finally, California environmentalists would not permit mining, even if a rich deposit were found. These may all be pefectly valid reasons not to explore for gold along this belt. But a look at the history of gold mining in California may suggest that the future for exploration and development of the gold resources in the Sierra Nevada Foothills can be greater than the past. It was during the Depression of the 1930s that this gold belt saw its maximum activity, with more than 3000 mines in operation. Some were father and son operations producing only a few tons of ore a day. Other mines employed hundreds of miners to hoist more than 900 t/d (1000 stpd). Some ores were of marginal grade. Others, like the Sixteen to

Jan 7, 1987

By G. J. Jansen

Introduction Coal petrography is the microscopic study and description of coal and other organic substances in terms of the microscopically recognizable organic constituents and of the rank of the material. Petrography studies of coal and other organic substances are useful in themselves, and have a number of practical applications. Some of these include determining coking quality of coal, estimating yields of methane from coal beds, and assisting in coal preparation. Petrography studies can also help in mine planning, in source rock studies, in combustion problems, and in synthetic fuel applications. Most recently, petrography has been found to be useful in studying sediment-hosted ore deposits. Microscopic constituents-macerals Under the microscope, coal is a heterogeneous assemblage of the coalified remains of woody matter and plant remains from the swamp or bog from which the coal was formed. The individually recognizable materials making up the organic mass are termed macerals by petrographers. They are the organic equivalent of the minerals that make up rocks. Coal petrographers can examine coal in two ways. Those interested primarily in the microscopically transparent macerals or in the study of very low rank materials, lignites for examples, like to use transmitted light microscopy where light passes through a microsection about 3 to 6 µm (0.0001 to 0.0002 in.) thick (Fig. 1). The dominant red color of the section in Fig. 1 is inherent, not a stain or produced by a filter. Distinct suggestions of textures in the reddish constituents can be seen. The red material is vitrinite and is thought to be the coalified remains of woody tissues. The yellow, lenticular substance is exinite. It is what remains of the waxy coating on a spore. Also seen are white dots of mineral matter and black opaque material that is better seen and described in polished sections. A polished microsection is prepared by grinding and polishing to a mirror polish a compact of coal and embedding resin or a coherent chunk of coal. The resulting microsection is then examined in a reflecting microscope. High magnification, 200X to 400X, and an oil immersion lens are needed to resolve fully the macerals of the organic matter so prepared. Figure 2 shows two macerals, an even gray vitrinite, unstructured and fusinite, a whitish, more or less cellular constituent. Fusinite is thought to represent more or less completely carbonized cellular material. Coal petrographers appreciate that coal is a heterogeneous material. Figure 3 is a view of another section. The generally structureless gray material is vitrinite. The bean-shaped dark constituents are exinites. The long, narrow, somewhat scalloped dark constituent is cutinite. Cutinite is thought to be derived from leaf coatings. Some coal petrographers would describe the large white, structureless mass as micrinite. Others might prefer the term macrinite. A few fragments of fusinite can also be seen. Figure 4 shows a coalified fungal remain, termed sclerotinite. It is surrounded by vitrinite and crushed fragments of fusinite all floating in a sea of embedding resin. Sometimes the microscope reveals the great heterogeneity of coal even on the microscope scale. Figure 5 shows a section with semi-fusinite, inertodetritinite, vitrinite, and micrinite. The nature of the heterogeneity may be judged by noting that the long dimension of the photomicrograph is about 70 µm (0.0027 in.). Measurement of rank Hoffmann and Jenkner (1932) discovered that measuring the reflectance of the vitrinite constituent of coal in a polished section gave a good measurement of the rank of the coal. In practice, a photomultiplier is used to mea-

Jan 9, 1987

By J. M. W. Mackenzie

Three major developments have dominated the recovery of gold from ores, viz; 1846 Elsner's studies of gold solubility in aerated alkaline cyanide solutions 1887 MacArthur and Forrest's process based on zinc precipitation 1980+ Large scale development of the CIP process The brevity of this list for a metal whose extractive metallurgy has been as widely studied as any metal reflects the fact that the chemistry of the gold recovery process, as opposed to the engineering has seen few significant breakthroughs in over a century of application. to The purpose of this paper is introduce a new chemistry for gold recovery which may, in the future, add significantly to the technology of gold recovery. This new chemistry is based guanidine functionality - (R2R2N) 2 C = N R2 on the In developing the use of the guanidine functionality for the recovery of gold as the aurocyanide anion from leach solutions and pulps, Henkel has, as its objective, the development of an effective gold selective ion exchange resin. As will be seen, the initial investigation of the properties of the guanidine functionality in the solvent form for gold recovery lead to some significant metallurgical developments of their own but, the foremost objective was, and remains, the development of a resin for gold recovery. As a part of this development project, Henkel has entered into association with Mintek, South Africa, and this ongoing association has the potential to develop RIP/RIS technology for the gold industry. More details on Henkel's gold recovery technology and Mintek's role in this programme are given in (1). an The application of ion exchange resins to gold recovery is not new and many reviews have been published on the advantages of this process, and on the current state of the art of gold recovery by ion exchange resins. (2)(3)(4) Widespread application of ion exchange resin based processes for gold recovery has, however, not occurred to date. Fleming and Cromberge (5), described some significant pilot plant RIP trials in South Africa and the partial success of these pilot plant trials held out the tantalising possibility of a breakthrough in gold recovery based on RIP technology. Despite this pioneering work, only two really significant applications of RIP technology for gold recovery exist today. These are the Golden Jubilee Plant in the Eastern Transvaal South Africa and the widespread, although in some respects technically backward, use of ion exchange for gold recovery in the former USSR. There are two main reasons for the current scarcity of RIP and RIS plants used for gold recovery. These are: 1. 2. The perceived problems of engineering RIP plants The lack of a resin having suitable chemistry for both gold loading and gold elution The problems of engineering RIP plants have been largely overcome in the last 25 years. Much of the development in RIP plant design took place in the uranium industry during the rapid expansion of that industry in the mid 1970's. In addition, the development of CIP engineering systems has lead to the development of in-pulp solid extractant based metallurgical plant designs which can, with limited modification, be adopted to RIP. The lack of a resin having suitable chemical properties has almost certainly been the main reason for RIP/RIS not being introduced into the gold industry over the past decade. Attempts to use RIP/RIS have focussed on either conventional we ak base or conventional strong base resins. The weak. base resins have the disadvantage that they

Jan 1, 1993

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 Fred Leighton

INTRODUCTION Dynamic (seismic or microseismic) methods of determining the stability of structures in rock are based on detecting and analyzing the characteristics of seismic energy that has originated from or traveled through the rock mass. This seismic energy can be in the form of naturally occurring rock noise energy resulting from structural adjustments within the rock or can be introduced into the structure by physical means, such as by blasting or impact. In either case, the seismic energy radiating through the rock mass can be detected using standard equipment and can be analyzed by established techniques to reveal a wide variety of information concerning the condition and stability of the rock mass through which the energy has traveled. In the following sections, the basic instrumentation required for seismic and microseismic studies is described, and some of the presently used applications of these methods are discussed to exemplify the state of the art. INSTRUMENTATION Seismic disturbances in a rock structure generate two types of seismic wave radiation, body waves and sometimes surface waves, which radiate outward in all direc¬tions from the source of the disturbance. Underground mining applications are generally concerned only with discerning the characteristics of the resulting body waves, i.e., the compressional (p-wave) and the shear (s-wave) energy. As these two forms of energy travel through the rock structure, the particles of the rock mass are caused to vibrate, and the vibration character¬istics resulting from each of the two types of wave are distinct. Some important differences are: 1) Compressional and shear waves travel at different velocities through the rock structure. 2) The frequency at which each wave causes particles to vibrate is different, and may range from about 50 to 100 000 Hz. 3) The amplitude or energy level of each wave is different, with the shear energy usually being the greatest. These differences form the basis for equipment se¬lection for individual studies and for modern data analysis techniques. The following sections describe the basic equipment necessary to detect and record seismic wave energy data and show several examples of analysis procedures and how these procedures have been used. In principle, seismic equipment is very simple. It consists of a geophone (or geophones) to detect the seismic energy vibration and convert that vibration to an electric signal, an amplification system to increase the level of that signal, and a means of monitoring and/or recording the signals detected. Fig. 1 is a block diagram of a typical system. The following sections offer a very brief discussion of system components and their individual functions. A more complete discussion is given by Blake, Leighton, and Duvall (1974). Geophones The function of the geophone is to detect the vibrations caused by the passing of the seismic wave energy and to convert that vibration into an electrical signal that displays both the amplitude and frequency characteristics of the vibration. Particle motion or vibration can be quantified and measured by measuring displacement, velocity, or acceleration of the particles. Thus, there are three types of geophones: displacement gages, velocity gages, and accelerometers. The choice of gage depends on the characteristic frequencies of the seismic energy to be monitored and the sensitivities of each type of geophone. In general, displacement gages are used for low-frequency monitoring (periods to 1.0 Hz), velocity gages for medium-frequency monitoring (1.0 to 250 Hz), and accelerometers for high-frequency monitoring (250 to 10 000+ Hz). Experience has shown that in underground studies, the choice of which gage to use lies between velocity gages and accelerometers. An easy, accurate method for selection of gage type is discussed by Blake, Leighton, and Duvall (1974). Once the type of geophone has been selected for use, it must be properly installed, and in the installation procedure the most important step is insuring that the gage is firmly attached to a competent portion of the rock structure. Poorly mounted geophones may entirely fail to recognize low-level seismic signals and will distort the information from signals they do see. Amplifiers Seismic events associated with mine structures occur over a very broad range of energy which results in a broad range of geophone output levels. In general, geophone output levels occur in the microvolt to low milli-volt range, and it is necessary to amplify these signals in order to drive recording or monitoring equipment. Because either an accelerometer or a velocity gage might be used as the geophone, the amplification system must

Jan 1, 1982

By N. J. Peberdy, C. N. Moreton, K. C. Garner

Introduction During the past decade a major objective of the process industry has been to use digital computer technology to improve plant operating efficiencies. This objective implied some form of optimization, a concept that has various interpretations depending on the view of the prospective user. For the purpose of this paper, optimization of a process plant is defined as the establishment and setting of plant operating conditions that maximize some mathematical yield function, i.e. maximum profit, minimum residue, etc. Analysis of these objectives and the available design and implementation techniques led to the conclusion that digital computer and optimization techniques are not the stumbling blocks, but rather the development and derivation of the mathematical models of the unit operations and process plants to be optimized. Such models should not only describe the optimized (steady-state) objective, but also how one steers to this state (control algorithm). Due to the multidisciplinary nature of the skills associated with the design and operation of process plants, the development of suitable models by a single discipline, such as the process control engineer, was found to be not only difficult but often impossible, due to budget and human resource limitations. To over-come these limitations, a computer aided design (CAD) tool has been developed. It aims to provide a productivity tool to the various disciplines, at the same time coordinating the technical input from each. The system described is but the starting point in an evolutionary development of a tool that, with use, is becoming more efficient and cost effective to use. Development has become an application engineering activity rather than the preserve of the computer specialist. Project phasing The development of a mathematical description of a process plant requires coordination of information from conceptual design to operation management. The activities required to build and operate a process plant are divided into four basic chronological activities or phases. These activities are often undertaken by different organizations and disciplines. As a result, continuity is often lost with the resultant loss of critical design data. The major activities are considered to be: conceptual and flowsheeting; detailing around the P & ID; building and commissioning; and plant operation. The CAD system described provides a design tool to be used for each of these activities, as well as providing continuity between the activities and the disciplines involved. The heart of the system is the dynamic simulation of the flowsheet. Each of the activities will be discussed, leading to two simple examples that demonstrate the use of the simulator. Figure 1 shows a schematic format of the various activities and the path followed by the dynamic flowsheet simulator in the life of a project. Flowsheet development The prime requirements in the design and develop¬ment of a process flowsheet are • selection of the correct unit operations to achieve the most economic (capital and operating) beneficiation of the specified reserve ; • the sizing of the unit operations to achieve the desired results, as a function of the projected feed rates etc., to handle the time related (dynamics) of the process; and • the production of a set of engineering documents showing the drawn and labeled flowsheet with an equipment list and process specification for each of the unit operations. The question may well be asked at this stage why dynamic flowsheet simulation should be considered when steady state modeling has been found to be adequate to date. With the increases encountered in the cost of capital, one often cannot afford the luxury of designing around the compounding worst case technique. Further, a more accurate design of the control surges can be achieved. No information is lost in that the steady state solution is in fact a subset of the dynamic model. In generalized state space modeling, the differential equations describing the process dynamics are illustrated in the following matrix notation: XDOT=A.X+B.U(1) Y =C.X+D.U(2) where XDOT describes the set of first order derivatives of the system state Vector, and X- is the system state Vector; A - is the system matrix operator which in the general nonlinear case is both a function of X and time ; U- is the process input vector; B - is the input mapping matrix; Y - is the set of observations; C - is the output mapping matrix which maps X - onto Y; and D- maps the input onto the observations. Thus, by time integration of the system dynamic equations, described in (1), the dynamic trajectory away from any set of initial conditions can be deter¬mined. Further, by finding the conditions at which XDOT = 0, the steady state solution can be determined.

Jan 1, 1987

By Robert F. Goss

OCAW appreciates the opportunity given to us by the sponsors of this Conference to present our position and policies on the issue of radiation hazards in mining. Our principal concern is the health impact that the mining of uranium has on our members. OCAW represents 1,500 underground uranium miners and more than 10,000 underground miners with 3,000 in the Rocky Mountain region. The U.S. Public Health Service has determined through mortality studies that the number one cause of death among uranium miners is lung cancer. It was also determined that exposure to radon daughters and mine dust correlates with the lung cancer experience of uranium miners. Data from the U.S. Mine Safety and Health Administration has also shown that not only uranium underground miners, but all underground miners, are exposed to radon daughters -- especially underground miners in the Rocky Mountain region. It is our position that any OCAW underground miner is at potential lung cancer risk. The dosages of radon daughters that our miners are exposed to are very many times the background levels of radon exposures in the communities where they live. We are also aware that cigarette smoking accelerates the onset of lung cancer; however, it has to be clear that the available scientific evidence shows that alpha radiation does initiate lung cancer and that cigarette smoke, as a recognized co-carcinogen, promotes cancer already initiated by radiation. It is true that cigarette smoke increases the risk of cancer significantly for miners exposed to radon, but nonsmoking miners have experienced lung cancer rates twice as high as the comparable members of the U.S. population. OCAW's position is that the occupational regulatory agencies should concentrate on the exposures that can be controlled; that is, occupational exposures rather than life-style exposures. Our Union has maintained a consistent posture in relation to carcinogens in the workplace -- that is, exposure to cancer-causing agents should be limited to the [lowest feasible level]. OCAW has interpreted lowest feasible level as the lower limit of detection of the collection and analytical method used to detect the carcinogen. Our posture is based on the available scientific information on carcinogenesis. We have asked the scientific community, many times, to provide us with safe levels of exposure to carcinogenic substances, including radon daughters. The answer has been: "We cannot determine levels of exposure low enough to assure that no cancer will occur." In short, there is not a "safe threshold" for any carcinogen. This statement does not come from one of the few so-called "pro-labor scientists," it comes from the National Cancer Institute and the National Institute for Occupational Safety and Health. I don't need to be a scientific sage, then, to conclude that the lowest level of exposure corresponds to the lowest risk of developing cancer. That is, then, our policy on exposure to carcinogens. It seems there has been an attempt to ignore the fact that lung cancer in uranium miners is the principal cause of death. Uranium miners are no exception from workers exposed to carcinogens. Our policy applies to them. Uranium miners should be exposed to the lowest feasible level of radon daughters and any decrease in the permissible exposure level is a decrease in their lung cancer risk. Accordingly, OCAW has petitioned the Department of Labor for a new permissible exposure limit to radon daughters in uranium mining, which lowers the current exposure standard from 4 Working Level Months (WLM) per year to 0.7 Working Level Months per year. We made our demand to the Department of Labor on April 20, 1980. We are still awaiting action from the Federal Government on our petition. OCAW is also very concerned with other important health impacts of uranium mining. We are concerned with a rate of disabling accidents and fatalities which is twice as high as the same rate in other underground mines, excluding coal. We are also concerned with the rate of respiratory disease fatalities among uranium miners which is almost four times the rate among a comparable U.S. population. We have expressed those concerns when the U.S. Senate proposed a Federal Compensation Act for uranium miners. That proposal, by Senator Dominici of New Mexico, found a quiet death in two Congressional sessions. In conclusion, our position on lung cancer induced by radon daughters is the same position we have taken with all other industrial carcinogens: The lower the exposure, the lower the risk. OCAW is demanding a drastic decrease of the permissible exposure limits. OCAW will never accept that a segment of our membership which mines uranium should take the lion's share of the risk while the uranium mining companies take all the benefits.

Jan 1, 1981

By Karl W. Mote, R. K. "Ivan" Urnovitz

One of the Northwest Mining Association's guide-lines is: "If you always do what you've always done, you'll always get what you've always gotten." Another way of saying this is, simply, "when the road curves, don't refuse to turn." Our industry has been roundly criticized for living in the old political world, and, I believe, justifiably so. Until about 20 years ago, members of both houses of Congress deferred to specialists among their members for mining legislation. This was true for all public lands and natural resource issues, and were referred to as the mining bloc, or the timber bloc, or the ranching bloc. We felt safe, and to a large degree, were safe in the sense that western members of the US House of Representatives and of the US Senate understood the need for multiple use of public lands, and the need for the natural resources they supply. They also knew what it takes to stay in business and how to create the economic atmosphere that encourages private industry to find and provide the natural resources. Almost daily, members of our industry bemoan the destructive effects that current regulations, let alone new proposals, are having on American industry. And, there are some intriguing aspects for us to ponder. First, the specific complaint about the government is not regulation per se, but rather that the rules or policies don't make sense. We recognize that few understand our industry, but we have assumed that the government will rely on people experienced and knowledgeable in the area being brought under government controls. There was a time that this assumption was valid, but it is no longer true since we have entered the era of the professional regulator. Much like the MBA who believes that a professional manager can manage anywhere, the new breed of regulator sincerely believes that process is the key to developing effective regulations or programs. They reason the process virtually assures that all the pertinent facts will inevitably come to light and, after receiving appropriate considerations, will result in the proper decision being reached and implemented. Unfortunately, following the process also allows politics to play a new and greater role than ever before, since information based on what can only be called weird-science can now be presented to the professional regulator and accepted as fact. Regulators do not have the background to question the basic tenets of the conclusion being presented to them. To add insult to injury, testimony from people in the industry involved, intended to set the record straight, is too easily discounted as being merely self-serving. A second aspect is that the question of why the associations aren't doing more to stop poorly conceived legislative and regulatory proposals, or to counter the ever-increasing influence of the selfappointed public interest groups is being increasingly discussed. This is occurring despite the fact that there is a higher level of involvement and closer cooperation between the various associations than ever before. This raises a most important question: where are we going wrong? In the last 20 years or so, major changes have taken place. A new philosophy, taking advantage of the public's recognition of need to protect our environment, has swept through the nation, resulting in a whole new body of law enacted to assure environmental protection. We all acknowledge the need to prevent undue degradation during the course of mineral development, and we all consider ourselves to be responsible stewards of the land, because we live and work in the outdoors and know the importance of environmental protection. The changes in the federal statutes apply now to every aspect of mineral discovery, development, and production, and assure adequate reclamation of lands when we are done. Along with the positive changes, are a constant barrage of attempts to stop public land use, or stop mining, or stop any new economic development, ostensibly to "protect the environment," but often reflecting a hidden agenda of social reform. These types of extreme, irrational, and just plain impossible proposals are everyday occurrences, but we have not geared up to challenge all the proposals which could end mining. Instead, we did what we always had done. The road turned, but we didn't. And we wondered why those who had always protected mining weren't doing their job, and why we were having such limited success in reining in the regulators. The Congress had changed, the public demands had changed, the recognition of importance of natural resources had

Jan 1, 1991

By R. Poulin, R. Pakalnis, D. Peterson

Introduction A comprehensive project has been completed at the University of British Columbia that focused on the development of a commercially competitive composite cable bolt. Laboratory evaluation consisted of pull tests, preliminary shear tests, grout column testing and scanning electron microscope evaluation of the composite cable bolt. Field implementation at five underground mines throughout Canada were completed for technical and overall performance evaluation of the new product. The project began in 1989 under the sponsorship of HDRK Ltd., a consortium of mining companies including Inco, Noranda and Falconbridge Ltd. The objective was to develop a cuttable cable bolt that would serve as support for a continuous miner within a hardrock environment (Mah, 1994). The bolt was required to be cuttable by a continuous roadheader without adverse effects to equipment, personnel and the milling process. From the beginning, the focus was to employ existing cable bolt installation technology (Goris et al., 1994) whereby the cable bolts are grouted within a 4.8-cm-diam (1.08-in.diam) or larger employing a portland cement grout having a water: cement ratio between 0.35:1 to 0.4:1. The prototype, developed from Phase I, was technically comparable to that of conventional steel cable bolts. However, high cost and limited accessibility hindered its commercial acceptance. The second phase of the project was sponsored by the British Columbia Science Council and Pacific Pultrusions Ltd. of British Columbia to locally develop a glass fiber cable bolt. This phase of the project has resulted in the DAPPAM cable bolt (Pakalnis et al., 1994). DAPPAM Cable bolts are the second most used support method in Canadian underground mines after mechanical rock bolts. It is projected that 870,000 m (2.9 million ft) of cable bolts are installed annually in Canada (Poulin et al, 1994). The objective of the research program was to develop a glass fiber cable bolt with similar pullout strengths to the conventional steel cable bolts (267 kN - 60,000 lbf). More than 200 laboratory pull tests were conducted employing standard cable bolt test procedures, as outlined by the US Bureau of Mines (Goris, 1994), to assess optimum rod and surface composition, geometry, number required, pullout and shear characteristics, among others. This resulted in the DAPPAM cable bolt (Fig. 1). It consists of 10 individual 0.6cm-diam (0.25-in.-diam) rods enveloping a 2.2-cm (0.8-in.) outside diameter, high-density polyvinyl grout tube. The composite used in the construction of the rods is comprised of about 65% glass fiber within a 35% polyester resin matrix (by volume). Through pultrusion (a continuous fabrication technique), the fibers and resin are combined into a composite material. The strength of the overall composite is largely governed by the material properties, the alignment [ ] of the fibers relative to the applied load and the density and interaction of the overall components within the composite. A further modification to the composite was the addition of a sand grit to the surface of the individual rod. This increased the overall bond strength between the rod and the cement-based grout, used to solidify the cable bolt to the rock mass. The grout tube has an inside diameter of 1.9 cm (0.75 in.) and is rated at a pressure of 1.72 MPa (250 psi). During installation, the grout is pumped through the annulus of the grout tube thereby filling the hole from toe to collar. Laboratory and field trials concluded that a minimum drill hole diameter of 4.8 cm (1.08 in.) be used to ensure optimum grout coverage. Table 1 compares the design strengths (as tested) for DAPPAM vs. the ultimate strengths for a conventional 1.6cm (0.6-in.) steel cable bolt. The characteristics for the steel cable have been derived directly or indirectly by Goris et al. (1994). A design tensile strength of 289 kN (65,000 lbf) was proposed for the composite developed. However, recent tests have indicated maximum breaking strengths of more than 400 kN (90,000 lbf) (Fig. 2). Figure 2 shows the load vs. displacement pull out curves for steel and DAPPAM cable bolts. The methods of testing were standardized so as to be able to compare the results directly. The shear strength of the fiberglass cable is largely a function of the surface area sheared. Shear strengths of 89 kN (20,045 lbf) were obtained for the configuration shown in Fig. 1 when sheared normal to the axis of the cable. The testing procedure was similar to the shear tests performed on steel by Goris (1993). It is important to reinforce the implication that critical bond strength has on the success of an overall support system.

Jan 1, 1995

By P. Stanley Jacobsen, Hemant B. Gala, Richard E. Hucko

Coal is the nation's most abundant fossil energy resource. However, its use is limited because of the nature and amount of associated impurities, such as mineral matter, sulfur, trace elements, and moisture. The moderate-to high- sulfur content of many U.S. coals is of particular concern as current and newer, potentially more stringent, regulations require reduction in sulfur oxide emissions. Release of particulates, NOx, and other pollutants during the coal-use cycle is also of concern. The current practice of controlling sulfur oxide emissions in the major steps of the coal- use cycle is limited mainly to coal preparation by physical beneficiation and, to a smaller extent, to post-utilization cleanup (flue gas desulfurization). Approximately two-thirds of the coal from states east of the Mississippi river is physically cleaned, whereas only 15% to 20% of the operating boilers in the utility industry are equipped with scrubbers to desulfurize the flue gas (Henzel, 1988). Physical beneficiation methods can be very effective in removing mineral matter, including pyritic sulfur, from coal. These processes take advantage of the differences in physical proper- ties between coal and mineral matter to achieve the separation. In a typical physical cleaning operation, raw coal is crushed to about a 50-mm (2-inch) top size. It is then screened into coarse, intermediate, and fine sizes. Crushing the coal liberates some of the ash-forming minerals, including a portion of the sulfur that is not chemically bound to the coal. The per- centage of impurities freed from the coal matrix generally increases as the coal particle size decreases. The coarse- and intermediate-sized coals are typically subjected to a cleaning pro- cess based on differences in specific gravity between coal particles and the associated impurities. Such cleaning processes used include jigs, dense-medium baths, cyclone systems, and concentrating tables. The fine-sized (typically less than 0.5-mm) coal, if beneficiated, is typically cleaned using a froth flotation technique based on the differences in surface proper- ties between coal and its associated impurities. At fine sizes, greater cleaning potential exists because a greater portion of the impurities is released from the coal matrix. However, this potential is not fully realized owing to inefficiencies in separating fine particles using state- of-the-art flotation technology. Overall, current commercial coal preparation processes are estimated to remove up to 60% of the ash and 10% to 30% of the sulfur, depending on the coal characteristics, with product weight recoveries rang- ing from about 60% to 90% and thermal recoveries of about 85% to 98% (Versar, 1985; Science Applications International Corporation, 1987; Hoffman- Muntner Corporation, 1978). For many U.S. coals, grinding to sizes finer than those used in current coal preparation plants is required to liberate most of the physically bound impurities from the coal matrix. As noted, conventional coal-cleaning methods are not very effective at these fine sizes. Accordingly, the expanded use of domestic coal requires the development of advanced coal cleaning technologies that can provide economic performance while reducing environmental impacts. Program Goals and Objectives The overall goal of the U.S. Department of Energy (D0E)'s Coal Preparation Program is to provide the scientific and engineering knowledge base that industry can utilize to produce economically competitive and environmentally acceptable clean-coal products for introduction as market conditions warrant. The coal preparation process(es) should remove undesirable components from coal and improve its handling characteristics, thus making the product coal technically, as well as economically, viable as a competitor with petroleum-derived fuels and natural gas, and with alternative environmental control technologies. The coal preparation process must be applicable to a wide variety of bituminous and low-rank coals. The cleaning process should reduce the variability among clean-coal products from various raw coals, thus significantly increasing the range of potential end uses for each raw coal. To the extent that deeply cleaned coal products have greatly reduced levels of

Jan 1, 1988

By J. Weyer

"A lot of incorrect, dangerous and misleading statements about the effect of CO2 on the human body can still be found in the internet, the media and in books. It is a fact that already low concentrations can lead to critical reactions, higher concentrations cause a narcosis. The real effect is not a lack of oxygen but an inner suffocation. The breathing reflex is decreased and the transport of oxygen in the blood is reduced due to a low pH value of the blood. Furthermore a reduced oxygen content in the air can also lead to the failure of diesel engines. MOTIVE Different gases can appear in active mines as well in abandoned mines, adits and caves. The most dangerous gases are CO, NO/NO2, H2S, SO2, CH4 and CO2. Although the general danger arising from these gases is mostly known some aspects are underestimated. Furthermore misleading or wrong statements exist about the detection and the effect of gases on the human body. This is especially true for CO2. In the last years a couple of accidents happened with a great number of injured persons and unfortunately also some fatalities. As a result wrong information and comments about the effect of CO2 have been published in the media. A huge unexpected CO2 outburst in a potash mine showed again some additional problems and limits for rescue troops. ORIGIN OF CO2 AND MISLEADING STATEMENTS ABOUT THE EFFECT ON THE HUMAN BODYCarbon dioxide can appear as free gas in the earth crust. It is especially known in significant concentrations in some potash deposits, in hard coal mines and some caves. Furthermore it can be formed due to complete oxidation of carbon. It can appear in technical processes. It is also used for or in technical processes, as fire extinguishing agent, as cooling agent, it is contained (added) in sparkling water or other soft drinks and it is formed by fermentation processes. Up to today a lot of wrong information about the effect of CO2 on the human body is published or is still in the mind of many people and unfortunately even in the mind of members or leaders of rescue troops. On August 16th 2008 a faulty CO2 extinguishing installation in the city of Mönchengladbach/Germany caused the injury of 107 people and the evacuation of further 150 people /1/. People on the streets became unconsciousness and felt down. Fortunately all could be rescued due to the immediate rescue operation of local fire fighters. Furthermore the CO2 concentration on the ground was diminished by help of a helicopter. What could have happen with the engine will be discussed later. After 6 days a second similar accident took place /2/. In the news members of the rescue teams explained that CO2 leads to suffocation due to a lack of oxygen. This is wrong! Also cave explorers often say that a burning candle can detect a low oxygen and/or high Carbon dioxide content in the air. This is also wrong and deadly dangerous! Similar statements we find nowadays in the internet /3, 4 and many other/. In a forum “science and mathematics > chemistry” several people say that winegrowers go with a candle into the vine-cellar to detect CO2 /5/. Again: this can be deadly! Another dangerous statement is that CO2 in higher concentrations leads to headaches so people should leave such areas. A last example found in the literature is that at 3 % the breathing volume is doubled and at a concentration of 20 % one can felt some warmth on the skin. Both is not wrong but can lead to false conclusions. /6/"

Jan 1, 2015

By James M. Barker, Richard O. Y. Breese

The recognition of perlite as a distinct volcanic glass, and of the arcuate fractures that often characterize these natural glasses, dates back at least to the late 19th century and perhaps into the 18th century (Howell, 1974). Some authors suggest that recognition of perlite dates considerably further into antiquity, perhaps to the 3rd century BC (Caley and Richards, 1956). Most recent definitions identify perlite by the presence of vit¬reous, pearly luster and by the presence of characteristic concentric or arcuate perlitic fractures. As discussed more fully below, arcuate fractures in perlite need not be megascopic. The textures of perlite which commonly occur in deposits range from dense to highly vesiculated to pumice-like. Thus, the classical definition of perlite is descriptively restrictive compared to the natural range of textures and neglects to address the issues of genetic origin and geologic occurrence. In spite of the petrological usage of the term perlite, the in¬troduction of expanded perlite aggregate into industrial markets caused the term to be applied as well to the lightweight, cellular aggregate that is produced through the rapid thermal expansion of milled perlite ore. Perlite is distinguished petrologically from other natural glasses by a silicic or rhyolitic composition by the presence of 2 to 5 wt% total chemical water held within the glass structure and often by the presence of pearly luster and onionskin-like perlitic fractures. Worldwide, occurrences are associated with Tertiary through early to middle Quaternary volcanics and with the glassy portions of silicic domes and flows, with vitric tephra, with the glassy chill margins of dikes and sills, and with the glassy portions of welded ash flow tuffs. Less commonly, perlite is reported in association with the glassy portions of volcanic plugs and lacco¬liths. The perlite carapaces that partially or fully comprise extru¬sive domes and flows are often thick and areally extensive, and thus provide commercially attractive targets for low cost open pit mining. Perlite tuffs and tephras are also important commercial sources. Mining costs are minimized by open pit quarrying and by either bulldozer ripping or blasting. Crushing and sizing facilities are generally close to the pits. Due to the low weight and large volume of expanded perlite, unexpanded perlite that has been crushed and sized is usually shipped nationwide directly to local markets where it is expanded and processed for distribution to end users. The vast majority of international trade is in unexpanded perlite. In the United States, the evolution of the expanded perlite industry can be traced to the late 1930s and early 1940s when expanded perlite aggregate was introduced into gypsum plaster and concrete markets as a substitute for vermiculite (Schundler, Jr., oral communication, 1991) and was promoted for use in a much broader spectrum of products (Howell, 1974). By the early 1950s, end user markets were well established and small expanders were numerous east of the Mississippi River. Today, mining and expan¬sion is undertaken throughout the world. Worldwide production of perlite is dominated by the former Soviet republics, the United States, and Greece. Total worldwide production of processed perlite during 1990 is estimated at roughly 1 778 000 tons (Davis, 1991). The physical properties of expanded perlite that give this prod¬uct unique commercial value include: low bulk density, chemical inertness, high insulating ability, nonflammability and fire resis¬tance, and its ability to retain water. These and other physical properties give expanded perlite value in such uses as: lightweight acoustical and thermal insulating aggregate; component in plaster, insulating cements and lightweight concrete; loose fill insulation; as aggregate in horticultural applications; and in milled form as filter aid. In the United States and Europe, the largest end use is within building construction products.

Jan 1, 1994

By Robert W. Lawson

INTRODUCTION The two basic types of mine-air compressors are the positive-displacement compressors and the dynamic compressors. Positive-displacement compressors con¬fine successive volumes of air within a closed space and increase the pressure by reducing the volume of that space. Reciprocating and rotary compressors operate on this principle. Dynamic compressors utilize an impeller to acceler¬ate the air, with the increased air velocity converted to pressure through a stationary diffuser. Centrifugal compressors and turbocompressors operate on this principle. Reciprocating Compressors Reciprocating compressors include both single-stage and multistage units. For mine service the most com¬mon is the two-stage intercooled compressor having an atmospheric pressure intake and a discharge pressure of 689 to 861 kPa (100 to 125 psi). The cylinders of a reciprocating compressor may be either single acting or double acting. In the single-acting cylinder, compression occurs on only one side of the piston, while, in double¬acting cylinders, compression occurs on both sides of the piston. Fig. 1 shows a cross-sectional view of a typical two-stage double-acting water-cooled reciprocat¬ing compressor. Fig. 2 illustrates the sequence of opera¬tions occurring in the cylinder of a reciprocating com¬pressor. Most modern mine compressors are double acting and are driven by electric motors. Induction motors are used on low-powered compressors, and synchronous motors are used on larger compressors. Steam-driven compressors once were common and still are applied where a reliable source of low-cost steam is available. For this type of compressor, steam is a desirable power source since a governor that senses the discharge pres¬sure can be used to control the steam inlet. This can be used to regulate the compressor speed to compress only air sufficient to meet the mine's requirements (up to the capacity of the unit). Unlike a constant-speed electric¬driven compressor, no other regulation is required for the cylinders of a steam-driven unit. Rotary Compressors Rotary compressors include sliding-vane, liquid¬piston, two-impeller straight-lobe or cycloidal, and helical-lobe units. Each of these has specific advantages and applications. Of these, the helical-lobe or "screw" compressor has the greatest application for supplying "100-lb air" [689 kPa (100 psi) ] to small- and medium¬size underground mines. Fig. 3 illustrates the operation of a sliding-vane compressor. Fig. 4 illustrates the operation of a helical lobe or screw compressor. Fig. 5 illustrates the operation of a two-impeller straight-lobe compressor. Fig. 6 illustrates the operation of a liquid¬piston compressor. Centrifugal Compressors Centrifugal compressors deliver 0.94 m3/s (2000 cfm) or more of air, and they are best suited to supply¬ing large volumes of compressed air for base loads. The efficiency of a centrifugal compressor declines rapidly when its operation deviates from its design point. If the power cost is not a critical consideration, a centrif¬ugal compressor is attractive as a result of the low initial and installation costs and the high degree of reliability. Fig. 7 illustrates a typical four-stage centrif¬ugal compressor, and Fig. 8 is a cross section of this type of unit. COMPRESSOR SELECTION The process of selecting a compressor requires sev¬eral considerations, including the capacity, the control system, the location, the power supply, and the cooling provisions. Compressor Capacity To determine the required capacity, the total air quantity and pressure requirements must be calculated. The air consumption of various tools, as listed in Table 1, must be included in this calculation. Some of the other considerations include: 1) The correction factor for altitude, as listed in Table 2, must be included in the calculation. While pneumatic tools require additional compressed air as the altitude increases, the compressor produces less. 2) A correction factor must be incorporated for the use of multiple tools. Table 3 lists appropriate correc¬tion factors for rock drills.

Jan 1, 1982

By J. A. Marshall

"INTRODUCTION The purpose of this paper is to investigate the potential for use of UAVs in underground mines and present a prototype design for a novel autorotating UAV platform for underground 3D data collection. The mining industry has recently shown increased interest in the use of unmanned aerial vehicles (UAVs) to assist in everyday operations [20, 11, 4]. Above ground, small UAVs are in some cases a more efficient, inexpensive, and safer alternative to manned aircrafts currently used for photography, inspection and security [6, 10]. For example, attaching camera, infrared and LiDAR payloads, UAVs can provide a low-cost method of obtaining highly accurate 3D photogrammetric data and aerial photography. UAVs are now commonly used in open pit mining operations for applications that include stockpile surveying, 3D pit modelling, facilities management, accident reporting, progress monitoring, and environmental assessment [11]. There are numerous companies offering UAV services for mining applications; e.g., UAV Geomatics, Leica Geosystems, Sky Futures, SenseFly and FlyTerra, to name just a few. Common services include stockpile and open pit volume computations, environmental assessment, aerial mapping and photogrammetry. Although the aforementioned services are thosemost commonly provided by UAV surveying companies, they may also add value to other daily mining operations. For instance, SenseFly offers pre and post-blast monitoring in order to identify the presence of misfires and wall damage. This data can also be used to reconcile the blast results with expected results. GEM Systems offers a UAV equipped with a magnetometer for mineral exploration surveys. Companies such as Barrick have even used UAVs for solids modelling at tailings dams and stability monitoring [11]. UAVs in mining have so far been mostly limited to surface applications. Harsh underground environments pose many obstacles for flying UAVs. The confined space, dampness, reduced visibility, air movement, and lack of control signal propagation hinders most operators from being able to fly a drone underground. It may be that truly practical uses for UAVs underground will require either autonomous or semi-autonomous flight capabilities. Although there are many difficulties with flying underground, the potential benefits from a working system could greatly improve mining operations. The potential benefits of deploying UAV platforms underground include access to unreachable and dangerous locations and removing personnel from unsafe situations. These benefits have the potential to greatly improve mine monitoring and mine safety. Research has shown that current UAV technologies exist that allow for autonomous indoor flight. Extensive research has been done to develop UAV systems that are capable of performing on-board simultaneous localization and mapping (SLAM), which can allow them to navigate and map a foreign environment autonomously [5, 7, 1, 13]. Grzonka [7] successfully used an open hardware quadrotor to autonomously navigate and map an office building. The research outlines the localization, mapping, path planning, height estimation and control of the autonomous quadrotor. Other research has been done that exploits autonomous UAVs for search and rescue. Kassecker [9] proposed a software and hardware framework for a quadrotor capable of indoor and outdoor urban search and rescue and Rudol [19] developed a system for human body detection and geolocalization using an autonomous UAV.The use of autonomous UAVs in search and rescue has the potential to improve situational awareness and surveillance for a rescue team. Similar research has been done that uses autonomous UAVs for indoor exploration. Pravitra [15] outlined a strategy for autonomous exploration of miniature air vehicles (MAVs) within indoor environments and Rudol [19] developed a method for using an unmanned ground vehicle (UGV) and a UAV for cooperative indoor exploration. These studies show that aut"

Jan 1, 2017

By S. Frimpong, D. Ali

For achieving bulk economic excavation, large dump trucks are being used in majority of the earth moving operations resulting in high impact shovel loading operations (HISLO) which exposes the operators to high levels of whole body vibrations (WBV). As the truck ages, hydro-pneumatic suspension struts loses their capability to effectively attenuate the vibration. A system is required to monitor and predict the performance of the suspension struts in real time. Machine learning and artificial intelligence (AI) has been applied for modeling and predicting the suspension system performance for lighter/smaller vehicles. However, no work has been done to implement machine learning or AI for modelling and predicting the performance of hydro-pneumatic struts in large dump trucks. Therefore, current work is going to be a pioneering effort towards developing machine learning and AI models for solving this problem. Support vector machine (SVM) and regularized dual non-linear regression has been implemented in order to achieve the goal. 3D virtual simulator for CAT793D was used to conduct experiments in MSC.ADMAS environment. The two main recorded outputs were the RMS accelerations in the vertical and horizontal directions at the operator seat for characterizing the performance of the suspension system. During the development and the training of the models, eighty percent (80%) of the total experimental data was used and the remaining 20% was used during the testing and validation of the developed models. SVM model showed the desired accuracy in terms of hydro-pneumatic suspension system performance prediction for large dump trucks. These models can be implemented in the dump truck controller which can then be used to monitor the performance of the suspension system in real-time, and with that proper maintenance and/or replacement can be scheduled by the maintenance personnel. Workplace safety, operator’s health and the overall system efficiency can be greatly improved with an implementation of such an intelligent system. INTRODUCTION Large capacity dump trucks are being employed in any earth moving operation throughout the world for achieving bulk economic excavation [1, 2]. A large dynamic impact force is produced with large capacity shovels dumping material weighing 100 tons or more in to the dump truck under gravity. This dynamic impact force generates high frequency shockwaves which travels throug0068 the truck body, chassis and reaches the operator’s cabin and thus exposes the operator to severe levels of whole body vibrations (WBV). Recommended safe limits have been listed by the International Standards Organization (ISO) in sections 1, 2, 4 and 5 of ISO 2631 (1997, 2003, 2001 and 2004). An operator can got through a long term back, shoulders, arms and neck disability and disorders if exposed to the WBV levels beyond those limits. The dynamic impact force generated as a result of the high impact shovel loading operations (HISLO) was modelled by Ali and Frimpong [2]. Furthermore, the complete methodology for optimizing the dumping process and reducing the impact force has been provided by Ali and Frimpong [1]. The vibration attenuation system present in the dump truck requires improvement in addition to the dumping process optimization. The vibration attenuation system in the dump truck commonly consists of hydro-pneumatic suspension struts. As the dump truck ages, its suspension system deteriorates and therefore loses its capability to effectively attenuate the shockwaves and the resulting vibrations produced by the dumping process. The deterioration tales place due to the contamination of the hydraulic oil or oil and nitrogen mixing or due to the gas accumulator diaphragm rupturing. The monitoring of the output vibration levels can be used to evaluate the effectiveness of any suspension system. With the dump truck suspension system, the vibration levels at the operator’s seat can be used for the evaluation of the effectiven

Jan 1, 2019

By James R. Arnold

Gold Fields Mining Corporation started processing ore from it's Chimney Creek operation in northern Nevada in November, 1987. Since that time production has averaged about 200,000 ounces per year from the milling and dump leaching circuits combined. The events that led up to commissioning of the plant started in January of 1985 when the Gold Fields exploration team drilled their first hole in an area now known as Discovery Hill. The major orebody was located later in the year in what is currently the major focus of activity, the North Pit. Subsequent drilling since plant startup has found another 14 million tons of ore two miles south of the original orebody. This South Pit orebody is currently being stripped and metallurgical testing is underway. Engineering Engineering on the Chimney Creek plant started in earnest in May, 1986. Davy McKee Corporation from San Ramon, California, did both the engineering and the construction management. Before engineering began the entire senior staff of the Plant Department was hired and assigned to the project. Senior staff included the Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist. Gold Fields does not employ an internal central engineering staff to design their plants. We feel that the people who design the plant should be the ones who will be required to run it once it has been commissioned. This philosophy affords a continuous accountability from start of design through commissioning and operations. All four of the senior staff spent the most critical phase of the engineering stage at Davy McKee, so the plant is truly a reflection of all four disciplines. In bringing on the Maintenance General Foreman from the start, there was a focus placed on ease of maintaining the plant. Another philosophy that was adopted from the start was that the plant be easy to work in and maintain for both operators and maintenance personnel. Numerous design criteria were given to Davy McKee to ensure this. For example, we required that every piece of equipment could be reached easily with a crane or fork lift. All conveyors were required to have walkways on both sides for ease of maintenance and a minimum of 30 inches of clearance under them for ease of cleanup. Visibility for operators was considered very important and the control rooms and general plant layout were designed with this in mind. A list of about 100 general rules were given to Davy McKee during the early layout meetings and these were held as the general design criteria throughout engineering and construction. While engineering was progressing in San Ramon, California, dump leach test work was being done at the site. Three 6 foot diameter fiberglass columns were constructed and filled to a 20 foot height with low grade ore. Leaching characteristics were determined and the dump leach design was based on information collected from these columns. Construction The major contractors on the project were The Industrial Company for the bulk of the piping and structural, EMKO on the electrical/ instrumentation and Hunewill for most of the earth work and leach pad construction. All plant operating and maintenance personnel were employed eight weeks before startup for intensive training. Instructors were Gold Fields personnel, including foremen, and Davy McKee engineers. On occasion, selected suppliers representatives were brought in to discuss major pieces of equipment. Project completion was targeted for January, 1988 at a projected cost of 78.9 million dollars. Actual startup was about 1 1/2 months ahead of schedule and slightly under budget. Full design capacity was achieved the first full month of operation. Crushing Run of mine ore is delivered to the crusher and dumped into a 90 ton hopper which is topped with a 30" X 30" opening stationary grizzly. Ore is fed to a 50" X 60" Telsmith jaw crusher with a 5' wide apron feeder. The closed size setting of the

Jan 1, 1989