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By Carl Barchet

INTRODUCTION Located on the Colorado River on the Nevada-Arizona border about 56 km (35 miles) southeast of Las Vegas, Nevada, Hoover Dam is a gravity arch dam 221 m (726') high for electric power generation and water supply. The head waters of the dam form a large reservoir, Lake Mead. The dam is equipped with two emergency spillways, one in the Nevada abutment and one in the Arizona abutment. Both spillways which are concrete lined and 15.25 m (50') in diameter, were intended for use only if Lake Mead is full and incoming flow of the Colorado River exceeds the discharge capacity of the turbines and regular bypass conduits. In the spring of 1984, a combination of rain and unusually rapid snow melt required the opening of the spillways for only the second time in the 50 year history of the dam. As in the previous use in 1942, the massive discharge of water through the tunnels resulted in severe damage to the concrete Linings of both tunnels. This damage is the result of cavitation which occurs as the discharge water flows at very high velocities over minor irregularities in the concrete lining. The U.S. Bureau of Reclamation has long been aware of the cavitation effect which occurs during these high volume discharges and has been studying solutions. One solution to the problem is to induce air bubbles into the flowing water to eliminate the vacuum associated with cavitation. This solution had been implemented at Yellowtail, Flaming Gorge, Blue Mesa, and Glen Canyon dams by constructing an aeration slot in the inclined portion of their spillway tunnels. Since the spillways of Hoover Dam were already in need of repair, the BUREC decided to construct similar facilities in each of the two 15.25 m (50') diameter spillways coincidental with the other repairs required.

Jan 1, 1987

By Viggo C. Bertelsen, David G. Hammond

LOCATION AND EXTENT OF BART TUNNELS The San Francisco Bay Area Rapid Transit District (BART) during the past six years has become the Owner of 23.5 miles of single-bore tunnels. The tunneling experience gained while working closely with our Engineer and Contractors in the design and construction of these major transit facilities provides a basis for our "Owners" viewpoint. The 23.5 miles of tunnels are divided into two categories. Fifteen and one-half (15-1/2) miles of tunnels were constructed in soft ground underlaying the downtown areas of San Francisco, Oakland and Berkeley. Eight (8) miles of tunnels were built in "hard ground" in the hills separating Oakland and Berkeley from Contra Costa County and in the San Francisco hills on the Mission Line. Location of tunnels within the BART system are shown on Figure 1. The tunnels comprise 15.7% of BART's 75-mile dual track system. Their 176 million dollar costs (excluding rail, electrification or ventilation features) are 12.4% of the system's total estimated cost of $1.4 billion. The tunnels constructed in soft ground were shield and boring machine driven with ground support provided by steel liner rings. The exterior diameter is 17 foot, 6 inches in the majority of tunnels; however, on two tunnels with a 500 foot radius of curvature, the diameter was 18 feet. BART awarded 14 contracts for soft ground tunneling. The work was accomplished under free air or compressed air depending upon the ground water and soil conditions, safety requirements, and considerations for protecting adjoining property from damage either from the de-watering operations or settlements due to excavation. The tunnel bores, except on curves, were normally parallel being spaced at 32 foot centers on the Mission

Jan 1, 1997

By Francis R. McAllister

SYSTEM ENGINEERING Years ago, while attending Harvard's Advanced Management Program, Professor Robert Hayes introduced our class to the, then not-so-familiar, concept of systems engineering. First assignment-dinner at The Benihana of Tokyo Restaurant! Delighted, but a bit mystified with the assignment; we had dinner at Benihana. Back in class, Professor Hayes inquired about the assignment and persuaded us; that Rocky Aoki's business formula for Benihana was not necessarily Japanese cuisine, but rather the systematic way in which he moves more people-if you will, processes more people through his restaurants than his competition, by systematically linking all restaurant functions. For instance, the bar is a holding area in which patrons ideally wait no more, or no less, than 20 minutes. When seated, your table, which accommodates eight, is filled with eight even if you are a party of two. The Japanese chef is also the entertainment, and when, at the end, he cleans the grill, you are being subtly prompted that dinner is over. The bill accompanies ice cream, systematically delivered shortly thereafter. You are expected to pay promptly and depart, as there are eight more, mostly unacquainted, people who have waited 20 minutes in the bar, and who are assembled ready to be seated at your table. For me-this was electrifying. I understood the concept, concluded it was very powerful and felt prepared to conquer whatever, by means of its application. That was years ago and I can safely say I have yet to fully realize on the euphoria of that moment. Why? Well, I simply conclude: • The concept of systems engineering is simple and not revolutionary, but The application of systems engineering is far from simple and is revolutionary.

Jan 1, 2004

By D. A. Barr

The Galore Creek copper deposits lie in the Stikine River district of northwest British Columbia, a largely undeveloped region which possesses rugged mountain terrain with deeply incised valleys which support dense undergrowth. Timberline is normally at about 1300 m (4265 ft) and the higher levels contain glacial systems and large ice fields with peaks rising to 3500 m (I 1,483 ft). A location map of the Galore Creek area is given in Fig. 1. In the late 1920s, field parties under the direction of F.A. Kerr (1948) of the Geological Survey of Canada completed geological surveys of the district. Primary access for this work was provided by riverboats on the Stikine and Iskut Rivers. Kerr's party had to contend with extremely difficult travel conditions that required considerable backpacking, augmented by pack dogs in certain cases. Absence of ground control led Kerr to establish his own topographic maps. Among the geological data established was the recognition [ ] that small bodies of orthoclase porphyry in the district are often well mineralized with base metal sulfides. Copper mineralization in the upper Galore Creek valley, a tributary of Scud River which flows into the Stikine River some 80 km (50 miles) south of the small settlement of Telegraph Creek, was originally discovered during aerial reconnaissance by M. Morrison and W. Buckholz while prospecting for Hudson Bay Exploration and Development Co. Ltd. in 1955. Malachite- and azurite-stained out- crops in several parts of the Galore Creek watershed are clearly visible from distances of 500 m ( 1640 ft) or more. Claims were staked in 1956 and the copper showings were partly mapped, trenched, and investigated by shallow diamond drilling. Because of the remoteness of the area and higher priorities for expenditures elsewhere, no further work was undertaken and most of the claims were allowed to expire.

Jan 1, 1991

By B. Pollack, L. G. Stevens

INTRODUCTION In other chapters of this volume many of the important technical aspects in the design of concentration and dewatering circuits have been discussed. This chapter will deal with the steps necessary to bring that information together in order to produce the specifications necessary to purchase the equipment and complete the design. It deals with the plant as a whole rather than just concentration and dewatering, since the steps are essentially the same. Prerequisites to Engineering Design Feasibility studies will have been completed and design parameters established prior to the start of the engineering design stages for a metallurgical plant. The data base for the feasibility studies will have included laboratory and/or pilot plant test work, historical data, proposed plant site data such as climatology, topographic mapping, geological and geotechnical information. These studies will have established the general project design criteria as well as the process design criteria leading to the preparation of preliminary flowsheet ( s) and/or alternative process flowsheets. These will have incorporated metallurgical and material balances, allowing identification of major equipment items, and development of preliminary plant site plot plans. Preliminary technical and economic assessments and recommendations will have been prepared during this phase for Owner's review and evaluation of the technical, physical, and cost effectiveness of the overall project and its component process, process facilities, and supporting facilities elements. After Ownert s approval and decision to proceed, the engineering design can be initiated, using the above information as the basis. Engineering Design Phase The engineering design of metallurgical plants and their elements are generally developed in two sequential stages, namely: (1) Basic ("Front end") Engineering Design (2) Detail Engineering Design Figure 1 relates these engineering stages to the entire program from feasibility through construction.

Jan 1, 1986

By Wenhua Wan, Jan M. Mutmansky

Many concerns have been expressed worldwide regarding the use of diesel equipment in underground mines due to the potential health risks associated with diesel particulate matter (DPM). In assessing the risk to the health of miners and designing control strategies for DPM, it is important to know the concentration of DPM in the working areas of a mine. In some mines, significant stratification occurs due to the characteristics of the diesel exhaust, affecting exposure and the ability to measure DPM concentrations accurately. This study outlines some of the concentration profile data obtained during sampling at three mines where diesels were employed. Mine #1 (a gold mine) employed large openings in which a diesel-powered roof bolter was used. Mines #2 and #3 are both room-and-pillar operations with varying conditions. Mine #2 is an experimental mine where two stationary diesels were utilized for sampling purposes, while Mine #3 is a traditional evaporite operation using two mobile diesel vehicles for section haulage. These field studies indicated that some level of stratification occurred in all the mines, but the stratification was not as pronounced in the case of Mine #2 due to the complex pattern of exhaust emissions generated by the vehicles. The stratification seemed to depend on the distance from the vehicle, the mine air velocity, whether the vehicles were moving or stationary, and whether the exhaust was diluted by equipment design features. The concentration of DPM was generally higher at samplers located near the roof and lower at samplers located near the floor. However, this was not found in all cases. This paper provides conclusions on how knowledge of stratification patterns can be utilized in sampling and control of DPM in mine atmospheres.

Jan 1, 1997

By Carmine Nappi

This article examines some important trends in East-West trade in non-fuel mineral products, in particular the recent decline in Eastern bloc exports of such products. It relates this decline to an improvement in Soviet terms of trade; a deterioration of geological and locational conditions in Soviet mineral production; an increased integration of the CMEA-countries based on a regional pooling of resources; and a surge in demand for mineral commodities in Eastern countries coupled with a slowdown in Western economic growth. Finally, it evaluates to what extent these determinants may evolve in a different direction during the eighties and thus foster East-West mineral trade. Eastern1 bloc trade in mineral products is a subject of perennial interest in the West. When exports increase, there is much talk of the dangers of dependence on the East for critical fuels and materials. When imports increase, talk turns to the danger of Soviet dependence on Western supplies. A common fear is that the USSR may not only seek to meet its own resource needs but also deny the access of Western countries to these resources. This paper, like three other recent additions to the lite5ature on Soviet foreign trade of commodities2, tries to put matters in clearer and cooler perspective. The first section evaluates the structure of East-West mineral trade and tries to identify some important trends of recent years. The second section relies on supply and demand factors, as well on the Comecon connection (the intra-Council for Mutual Economic Assistance mineral cooperation) to explain why there has been a reversal in the mineral trade trend since the end of the seventies. The last section evaluates the factors which could stimulate East-West mineral trade in the future.

Jan 1, 1986

By A. C. Lyons, A. J. Reed

INTRODUCTION A trend which is obvious in the United Kingdom and is thought to be evident in other parts of the Western world is the disproportionate rise in the cost of tunnel mining labour compared with other tunnel cost increases. This disproportionate increase in labour costs has been accompanied by, and may be the result of, a shortage of skilled working labour. To combat this problem there has been a steady increase in the use of machines for excavation and for the erection of tunnel linings to keep the amount of skilled labour working below ground to a minimum. The typical tunnelling cycle, where the ground needs continuous support can be excavation, temporary support, temporary lining, grouting and permanent lining. It is obviously an advantage to combine as many of these operations as possible and carry out as much as possible by machine. Machines have been developed which will, either alone or used in combination with a shield, excavate and, when necessary, temporarily support the excavation. If a segmental lining is used, it can be so designed that it will form the permanent lining and reduce the cycle time and skilled labour requirement to a minimum. In cases in which the ground can be cut accurately to a smooth profile and will stand without support for a sufficient time, the lining can be expanded against the ground and the necessity of grouting eliminated. In cases in which ground will not stand without support, the usual method is to use a bolted lining and to grout behind it preferably as the shield is shoved forward. Whichever technique is used (this is largely determined by the ground) it is an advantage to make the segments as large

Jan 1, 1974

By D. C. Torre

Ever since the introduction of the first successful continuous mining machines some twenty odd years ago, there has been an obvious, and steadily increasing, need for some form of continuous haulage machine, or system to work with them. There is universal agreement that the continuous mining principle, and the various continuous mining machines in particular, have done a great deal for mining. However, coal mining men, who have the longest and broadest experience with them, agree that continuous miners do not truly mine continuously. In fact, their estimates, and actual studies, show that the real mining time of these machines ranges, on the average, from 20% or less to about 30% of the available shift time. Part of the failure of the continuous mining machine to mine continuously may be ascribed to breakdowns in its intricate machinery. Most of the failure, however, is blamed on a variety of other factors. Of these, there are two which are given particular blame: face haulage and roof control. The relative importance of the two as the major negative factor in continuous mining is a moot point. However, there is an extensive concensus that haulage is the number one problem. Roof problems and their control are inhibiting factors regardless of the type of mining or mining machines employed. But the truth of the matter is that a continuous mining machine, like any type of loading machine, whether underground or on surface, is only capable of mining or producing in direct ratio to the mineable, or loadable, material that can be kept ahead of it and equally, if not more, important, that can be taken away from it. While the first commercially successful continuous miners were developed for, and proved successful in, coal mining, they were soon

Jan 1, 1974

By J. Georgi

A number of infrastructure projects are on progress all around Greece at the moment. The common propose of these a project is the promotion of quality in: Design, construction and maintenance stages. They also comply with environmental management, which is, a method of organizing and implementing environmental protection measures. The paper includes by special reference on two important modern transport infrastructure projects for the development of Greece and its communication with Europe and the Balkans. Their landscape types and natural vegetation, local climate, geological conditions, altitudes, aspects and soil types vary As a result, upland landscapes in particular, are endangered. Due to the complex of nature of this environment, the development and the landscape design of the projects is a high challenging task. Wherever the nature of the projects requires intervention to the environment, they intend to follow the direct restoration of landscape and method for ecological balance. The implementation of landscape design of infrastructure projects, aims at restoring the landscape affected by construction works, incorporating transport works into the landscape, emphasizing on areas of particular interest (e.g. archaeological sites), while at the same time observing the principles of an efficient operation and safe use of the alignment. Landscape design provides, among others for the protection of slopes against surface erosion and the reinforcement of their stability, the improvement of road functionality, the aesthetic improvement of the landscape, adaptation of the road landscape to the greater environment, visual guidance of the user, screening of lights of the vehicles moving in the opposite direction, the protection against wind etc. The landscape design strategy and landscape applications are specified by special landscape guidelines and specifications that can be considered as part of the sustainable development of a alignment creating a high a esthetic scenery, ensuring the security of transportation and the environmental protection.

Jan 1, 2002

By D. Koren, D. Carson

Disposal of smelter slag may be subjected to regulatory leach tests to deter- mine whether they produce a leachate containing contaminants in excess of stated limits. Contaminants that are most often of concern for smelter slags are As, Cd, Cr, Cu, Ni, Pb, and Zn. The present regulatory tests are based on the assumption of storage in a municipal landfill dump, where wastes are in contact with organic acids. Since slags are not normally stored in such an environment, these tests do not adequately assess the true leachability potential of the slags. Ten slags from six major Canadian nonferrous sulfide smelters were studied. Leaching was done following regulatory and modified tests, including leaching in columns with synthetic acid rain for a 25-week period. The slags were characterized by mineralogical analysis before and after leaching with synthetic rain. The synthetic mineral equivalents (hereafter referred simply as phases) in the slow cooled slags consist of fayalite grains and dendrites, amorphous glass, some spinel and small quantities of complex sulfidic prills and metallics. The phases in the quenched slags consist of amorphous glass, some spinel and small quantities of complex sulfidic prills and metallics. Some metal dissolution took place during leaching. However, for all leaching tests, the concentrations were always below the regulated limits. In general, the content of the potentially leachable elements remains unchanged in the fayalite, in the amorphous glass and in the spinel, whereas the quantity of sulfidic prill and metallics decreases after leaching. Therefore, most of the deleterious element dissolution comes from the sulfidic prills and metallic phases. Metal dissolution increases by crushing the slag, probably due to an increased exposure of the sulfidic prills to the leaching solutions.

Jan 1, 1998

By A. Redfern, B. G. D. Smart

The ultimate objective of the development work presented in the paper is to devise an appropriate method for relating powered support specification to geological conditions and mining practice, thus enabling the longwall method of coalmining to be successfully implemented throughout the varied coalfields of the world. The paper begins with a review of various methods previously used to determine support rating, each method being analysed with regard to its source, the logic and theory which constitute the method, and subsequent recommendations. An alternative approach is then proposed, the Powered supports being considered to be one of three foundations, sharing load with coal ahead of the faceline which acts as another foundation, and the waste behind the supports which acts as the remaining foundation. The load/deflection characteristics of the coal and waste, which in turn are related to geological conditions and mining practice, are shown to affect the load transfer on to and hence the specification of the powered support. In particular the characteristic of the waste foundation is shown to be affected by the caving mechanism, an alternative form to that mechanism accepted for British conditions being proposed and designated "parting-plane-controlled" caving. A finite element solution for a beam on elastic foundations was used to translate the above concepts into numerical results, these results being compared with data from Indian and Chinese operating conditions. The Indian case study examines the effect of using a conventional chock type of support in a one-web back mode when mining towards a roof break generated ahead of the face, while the Chinese case study compares measured subsidence over a longwall with "blocky" roof conditions with the model results. The comparisons are encouraging, suggesting that the approach used is worthy of further development.

Jan 1, 1986

By R. Gong, R. Bhaskar

[Full-scale laboratory experiments were performed to evaluate the effect of ventilation and water spray pressure on the potential dust capture efficiency of a ventilation system comprised of blowing ventilation and an underboon scrubber. The tests were performed for a 1.52-m (5-ft) roof height and 2.36 m-3/s (5000 cfm) blowing ventilation volume, and a 2.44-m (8-ft) roof height and 4.72 m3/s (10.000 cfm) blowing ventilation volume at three face positions of the miner. The scrubber suction rate was varied for the tests to study the effect of scrubber flow rate on capture efficiency. The airflow patterns and scrubber efficiency were determined by performing tracer gas tests. Tracer gas was released through a manifold placed in front of the rotating miner drum. The gas was sampled in a grid fashion using a real-time multi-point gas sampling system. The capture efficiency of the scrubber was evaluated based an the peak gas concentration of the gas laden air drawn through the scrubber and the amount of gas passing, through the scrubber as a fraction of the gas released at the face. The data were analyzed to yield scrubber peak concentrations, potential exposures at various operator locations and the gas removal percentage (GRP). The underboom scrubber appears to work best in the end sump position, relative to the start sump and start slab positions. The capture efficiency scrubber was higher for 2.44-m (8-ft) roof heights compared to 1.52-m (5-ft) roof heights. The data also showed that the scrubber air discharged behind the mining machine may be drawn into the face area by the action of the blowing ventilation system, resulting in recirculation.]

Jan 1, 1994

By F. C. Turin, T. L. Friend, A. G. Mayton, J. S. Jaspal, W. J. Wiehagen

This paper describes the serious injuries that have occurred to bulldozer operators working at domestic coal, metal and nonmetal mines in the United States. A serious injury is defined as one that results in either a fatality or in lost work time. Examined were 873 injury records compiled by the Mine Safety and Health Administration (MSHA). The time period covered was 1988 through 1997. These injuries resulted in 18 fatalities and 31,866 lost work days. All of these injuries occurred to dozer operators while they were performing common production tasks. An injury classification system was developed to code the narrative information that describes the circumstances surrounding the injuries. Jolts and jars accounted, for 604 or 70% of the serious injuries to bulldozer operators, and jolts and jars accounted for 75% of the workdays lost. Vertical jars comprised the largest set (354 incidents) of jolting and jarring injuries. In more than half of these incidents, the operator was backing up the dozer. Working near an edge carries significant risk for fatal injuries. Of 116 incidents where the dozer fell over an edge, rolled over or fell into a hidden void, 14 (12%) resulted in fatal injuries. In cases where the dozer operator either jumped or was thrown out of the cab in a fallover or rollover, seven out of 17 of the operators were killed. This study finds that further reductions in injury risk would require more widespread use of seat belts; consideration (and maintenance) of the seat and seat suspension systems that might reduce injury risk from jolts and jars; research interventions to assess the effect of alternative engineering designs to dampen or isolate the effects of shock and vibration; and continued attention to skilled performance, i.e., the integration of hazard awareness, recognition and response with dozer operational skills.

Jan 1, 2003

By Roger G. Haga, Richard P. Campbell

Introduction Project scheduling . . . . • What is it? • Why is it done? • Who does it? • How is it done? • When is it done? These are some of the questions that will be addressed in this paper. It is intended that by exploring these questions, an understanding of project scheduling, its purpose and benefits can be achieved. The purpose of this article is not to describe in detail "how to schedule", but rather to give an overview of the scheduling function on a project. A successful scheduling effort is dependent upon the people responsible for developing, implementing, and monitoring the schedule, and the tools and procedures that are used. These three elements are integral components of project scheduling. What is it? A schedule is the integration of network relationships, resource constraints and the calendar. The schedule identifies when the activities "will" be done, while the network states when they "must" be done. Networking techniques are the most widely accepted scheduling tool for complex projects. The benefit of the network is the simplicity in illustrating, presenting and analyzing the dependency relationships of a project. The logic network is the foundation of the schedule. When calendar dates are applied to the logic network, it becomes a schedule which provides the basis for evaluation of project status, forecasting the occurrence of events and controlling the project. The sequence of activities that requires the longest time to complete determines the overall duration of the schedule. This longest sequence of activities is called the "critical path". The "critical path method" (CPM) is used to describe this particular schedule network methodology. The two common conventions of CPM networks are the "arrow diagram met hod" (ADM) and the "precedence diagram method" (PIM). There are arguments in favor of each of these two methods, and usually the one that is used is a matter of personal preference, project directive or corporate procedure.

Jan 1, 1986

By Eduardo Poltronieri Trés, Anderson Denadai Thomazini, Rodrigo Boyer Fernandes, Francisco de Assis Dutra Macedo, Gilvandro Bueno, Maycon Athayde, Rossano Augusto Pilon Nunes

The earliest industrial application of high pressure grinding rolls (HPGR) at comminution was in 1984 in the cement industry. Since then, the equipment has been widely applied in mining activities. Despite the rapid spread across the industry, several challenges are still present in the equipment’s application, especially considering the complex ore breakage behaviour reported for this process. The pellet feed’s HPGR comminution was recognized as fundamental to efficiently increase the particle surface area (e.g., blaine specific area, BSA) at a lower energy and water consumption level (Chapman et al. in J Southern Afr Instit Mining Metall 113:407–413, 2013). The present study considers a novel milling process applying only HPGR as a re-regrind stage after concentration. The amount of successive ground product recirculation into the machine to achieve the required particle size for the pelletizing process was investigated. The study compared the ball wet milling process with the innovative processing technology for hematite–goethite ores. Green pellet balling and induration processes were simulated in bench and pilot scale. The milling results showed a considerably steep increase in BSA after each recirculation step into the HPGR, although compared with the ball-milled product the size fraction <45 μm% increased a modest five percentage points after three recirculation cycles, remaining steady until the last recirculation step. As a result, the green pellet generated presented higher concentration of pellets between 10 and 16 mm (%10-16 mm), with a higher drop number than typically observed when produced with ball-milled feedstock. Consequently, on the induration process, a higher pellet deformation was observed during the pot grate loading. Despite that, the fired pellets still presented an outstanding performance in cold crush strength (CSS) considered suitable to an industrial application.

By John Ralston, Jinming Duan, David R. E. Snoswell, Daniel Fornasiero, Jingwu Yang

The colloidal stability of synthetic silica spheres with clean, methylated and dehydroxylated sur¬faces was studied at different concentrations of dissolved gas and KCl electrolyte at a fixed pH of 4.2. A classic stability ratio/electrolyte concentration analysis shows that hydrophobic, methy¬lated particles undergo faster rates of aggregation with increasing concentrations of dissolved carbon dioxide. Similar data for hydrophilic particles and dehydroxylated particles show no change as a function of dissolved carbon dioxide concentration. Zeta potential data behave simi¬larly, showing a strong influence of dissolved gas only for methylated particles. The results are interpreted in terms of DLVO theory, with the surface-to-surface interaction dominated by the presence of very small protruding bubbles. Direct evidence for the existence of these very small gas bubbles (or "nanobubbles") was obtained using tapping mode atomic force microscopy. Small bubbles do not form on smooth, hydrophilic, or dehydroxylated silicon oxide water sur¬faces immersed in aqueous solutions under known levels of gas supersaturation. Randomly dis¬tributed small bubbles were observed over the whole surface of observation on methylated surfaces of controlled roughness. Bubbles formed on rough, methylated surfaces were larger and less-densely distributed than those on a smooth surface of similar hydrophobicity. The process of bubble coalescence was observed as a function of time. The macroscopic contact angle, measured with respect to the aqueous or gas phase, is very different from the microscopic contact angle detected by TMAFM and appears to be due to the influence of line tension at the pinned three¬phase contact line. The latter has a value of -3 x 10-10 N and acts to stabilize the small bubbles, flattening them and thereby reducing the Laplace pressure.

Jan 1, 2005

By Jack J. K. Daemen

A wide variety of subsidence prediction methods exist. They can be classified into empirical methods and mechanistic models. The empirical methods include profile functions and influence functions (Brzuner, 1973a,b; Kratzsch, 1974; Berry, 1978; Hood et al., 1981), the two empirical approaches of most practical value, if only be- cause they are readily programmed on calculators or computers. The profile functions have the advantage that they allow for ready determination of the necessary empirical parameters from field observations. They have the disadvantage of being restricted to subsidence calculations at a very limited number of points or lines above very simple (rectangular) mine geometries. This reduces greatly their practical applicability for damage predictions, which usually require estimates of strains, slopes or curvatures at structures that have arbitrary position and orientation with respect to planned mine excavations. Influence functions can be used to calculate subsidence at any point near an excavation of any in- seam outline. The main difficulty associated with empirical influence functions is the lack of uniqueness in the relation between subsidence and its cause, the convergence distribution over a mined excavation and its immediate surroundings. A secondary practical difficulty is the need for fairly arbitrary adjustments near the excavation edges, usually in the form of compensation zones (e.g. Brsuner, 1973; Hood et al., 1981). Determination of the appropriate empirical constants in influence functions is facilitated greatly when a direct mathematical connection exists between influence and profile function (several examples are given by Brauner, 1973 and by Hood et al., 1981), so that field data can be used to determine the parameters through backfitting of the profile functions, which is considerably simpler and more direct than backfitting of influence functions.

Jan 1, 1982

By V. F. Hollister

Copper and gold were discovered in the Copper Canyon area of the Battle Mountain district of Lander County, Nevada, in August of 1866. By 1876 a gravity plant was con¬centrating oxidized Copper Canyon copper-gold ores. The history of mining in the district to 1967 has been summarized by John E. Frost (in press) as a case history of the discovery of the Battle Mountain porphyry copper deposits. His summary is not repeated here; rather, the present case history covers exploration since 1967, primarily for gold ores. Subsequent to the construction by Duval of facilities to mine the porphyry copper deposits in 1967, Duval established a geologic staff at the mine to continue district exploration. The staff was guided by David W. Blake, who acted as chief geologist for Duval&apos;s exploration in this part of Nevada. Blake&apos;s activities led to the publication of numerous technical articles, including his contribution to the US Geological Survey Professional Paper 798-B. A part of that volume is attached as a reprint (pages 23 through 66) as an explanation of the geological setting of mineralization in the district. The work by Blake and his staff recognized metal zonation in the skarn environment at Copper Canyon, and their studies were incorporated in the development of the Tomboy-Minnie gold deposits that are peripheral to the porphyry. Exploration continued beyond the immediate vicinity of the porphyry and the Fortitude ore zones were discovered. These skarn deposits are described in the attached article by Wotruba, Benson, and Schmidt, published in the July 1986, MINING ENGINEERING. The continued exploration successes at Battle Mountain reflect on the well organized and effective exploration team developed at the mine. The team has provided successive discoveries in a district known mostly for ore bodies that are exhausted in a few years&apos; time.

Jan 1, 1990

Because of the different types of reagents used for the flotation of sulfide and nonsulfide minerals, it is convenient to consider these two groups of ores separately. The metal sulfides and native metals are substantially all floatable with xanthate and similar collectors.

Jan 4, 1982