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By J. Partyka, J. S. Stachulak

Dust produced during blasting in a copper/nickel based mine was studied and characterized. A stope utilizing the crater blasting technique, or the vertical retreat mining (VRM method), was examined. Two dust monitors were used to measure an airborne dust produced during blasting operations. The data collected in situ was primarily used to formulate the boundary value problem for a generalized equation that describes the transportation process of airborne dust from all places where dust is produced by the VRM mining method. The field data was also used to define the analytical forms for the initial and boundary conditions, to provide a simple and reliable equation for flow of dust in a stope vicinity, and to identify the non-measurable constants. The effects of the amount of blasting material and of the ventilation system in use on dust concentration are also discussed.

Jan 1, 1995

By Felicia F. Peng, Lili Li

The dispersion model of nonideal flow was applied to describe the hydrodynamic state within the flotation column. Residence time distribution (RTD) data of a laboratory flotation column were measured to determine the parameters of the model. The effects of operating variables and column geometry on the Peclet number which reflects the extent of axial dispersion were investigated and a semi-empirical expression of Pe was formulated. The dispersion model was validated for the column flotation of ultrafines coal. Under the conditions of sufficient aeration rate and frother addition, a good agreement between the measured recoveries and predicted data was obtained. The dispersion model with first-order flotation rate process of the flotation column developed in this study is useful in predicting the collection zone recovery of fine coal, and for the flotation column scale-up.

Jan 1, 1995

By Jaroslaw Drelich

This report briefly reviews the measurements and interpretation of adhesion forces measured between microscopic probes and substrates having micro- and nanoscaled roughness characteristics that were recently conducted at Michigan Tech using atomic force microscopy. The laboratory study concentrated on both rigid and deformable materials, with the probes larger than the size of roughness features of interacting surfaces. For such systems, a detailed surface characterization in terms of roughness parameters, asperity size and shape, and spacing between asperities, as well as analysis of deformation of materials is required. Analysis of the experimental data can be done with theoretical models, which are suitable for rough surfaces and therefore allows to normalize the measured forces. Alternatively, loading of the probe during adhesion force measurements can be managed properly to control deformations of the nano-scaled surface irregularities. Controlling the deformation reduces and even eliminates the effects of nano-roughness on the probesubstrate contact area, promoting conditions described by the contact mechanics models for the sphere-smooth substrate system.

Jan 1, 2005

By R. Hogg, R. H. Weiland, P. Bunnaul

The dewatering of flocculated clay suspensions by compression in a centrifugal field was investigated. The results were analyzed with respect to the direct applicability of centrifugal dewatering (as in solid-bowl centrifuges) and indirect applications to gravity sedimentation systems. An important objective of the study was to evaluate the self-weight effect using centrifugal forces to simulate consolidation of deep sediments. Experiments were conducted in which the sediment mass and initial structure were maintained constant while the effective sediment weight was varied by centrifuging at different speeds. The relationship between the limiting sediment density and the applied consolidation stress was found to be the same for both gravity and centrifugal consolidation, indicating that small-scale laboratory tests can be used to predict consolidation in deep beds. The effects of the initial structure were also studied using controlled variations in flocculation conditions. The results imply that, for the flocculated suspensions studied, the longterm consolidation behavior is relatively insensitive to the actual extent of flocculation.

Jan 1, 1995

By Rick T. Richins

Despite what may be overwhelming evidence that natural degradation of cyanide and cyanide attenuation in hydrogeochemical processes are capable of reducing or eliminating potential water quality problems associated with heap leach facilities, the recent sharp increase in the price of gold and wide interest in the cyanidation process has focused all regulatory eyes on cyanide leaching. Regulations continue to increase, design criteria are more strict, and detoxification requirements approach or exceed the level of detectability. Current governmental regulation of mining and milling operations using cyanide varies considerably from state to state. The lack of regulatory consistency at this level is largely attributable to the current Bevill Amendment exclusion of mine wastes as potentially hazardous wastes, the undefined status of the Environmental Protection Agency's ongoing mine wast study effort, and the mining industry's reluctance to develop a consistent policy regarding state versus federal mine waste regulation.

Jan 1, 1998

By Peter J. Huck, Swpan Bhattacharya, Madan M. Singh

This paper reviews the state-of -the-art techniques for estimating the duration of post- mining subsidence arising from different mining techniques and evaluates existing post-mining liability systems in the United States and abroad. For each mining technique (room-and-pillar and longwall) the mechanism of subsidence development is examined and the critical factors influencing the duration of subsidence are evaluated. These are followed by a comprehensive summary of field data on the duration of subsidence movements collected in various parts of the world. Using this information, graphical charts are developed to depict the relationship between the time duration and critical controlling factors. Also included are any mathematical relationships that may exist between the different variables. Finally, a survey of current state, federal and foreign legislation on mine operator liability time spans is provided. These laws are evaluated for their applicability in light of the data on duration of residual subsidence. On the basis of this evaluation, guidelines are suggested for the establishment of the mine operator liability structure for the United States.

Jan 1, 1986

By Dick Ruby

Most miners know that Latin America's major mineral areas are in the mountains of Mexico and down the backbone of the Andes in South America. These areas comprise some of the more rugged and challenging parts of the world's geography, and the risks of doing business in these areas are just as difficult as the impressive task of extracting and processing ores. However, Latin America does contain mineral deposits, some untouched and many to be discovered, that convince venturesome people that the tremendous effort required to mine them will be worthwhile. The rewards have often been adequate in the past. Hopefully, they will be in the future. Background The safest statement that can be made about Latin America is that the "ballgame is constantly changing." Here are some examples: Chile. Under President Frei in the 1960s, Chile welcomed foreign mining interests, even though the government was acquiring significant ownership interests in the big copper properties. Then, under Allende in the 1970s, the big mines were taken over and compensation was denied on various grounds including the allegation that extraordinary profits had been made that should have been taxed. After Allende, an about-face was made in policy; former owners of the expropriated mines received some compensation and new foreign investment was seriously welcomed. Many major companies announced intentions to return to Chile. Some have. Peru. After a military takeover in 1968, Southern Peru Copper

Jan 7, 1983

By C. D. Elifrits, N. B. Aughenbaugh

Introduction Federal and state laws enacted to regulate coal mining and the accompanying public concern about the adverse effects that mining might have on land use have focused much attention on the subsidence of the land surfaces over underground mines. Some questions asked are • Will abandoned mines subside in the future even though they have exhibited no surface adjustment to date? If so, when will subsidence take place? • Can a specific surface area be affected by subsi¬dence more than once? • What surface form will the subsidence feature take? • Is there any way to predict when subsidence will occur and end? An attempt is made here to address these questions insofar as it is possible with available data and experience of evaluation of subcritical extraction conditions. Mining that results in subsidence features can be classified as supercritical, subcritical, and critical with respect to the minimum span lengths of the worked-out rooms (Fig. 1). A complete basin will not develop fully immediately over subcritical conditions because the span lengths of the entries are less than the critical dimension (Wardell, 1983). Passageways, entries, and room-and-pillar panels are all included in the subcritical category. Because of the short span lengths, subcritical mining conditions tend to be stable over the life of the mine unless adverse geologic conditions exist (Gray et al., 1974; Dunrud, 1976). Various roof control and support procedures are employed to maintain entry stability during mining. Little information exists concerning the time factor of subsidence over areas of subcritical extraction. Data that have been collected relate to longwall mines, whereby the cavity sizes are designed large enough to induce caving behind the working faces so that complete subsidence at the surface occurs within a short period of time. The UK's National Coal Board (1975) has reduced the enormous amount of data it has collected to empirical correlations and has published these correlations in its Subsidence Engineers Handbook. The handbook states that "None of the foregoing remarks applies to pillar and stall working," with respect to the section discussing the time factor of subsidence caused by longwall mining (National Coal Board, 1975, p. 40). To understand and evaluate how time relates to subsidence above room-and-pillar mines, one must first recognize how mine cavities deteriorate and fail. The propagation of failure to the surface must then be studied, and, finally, the failures that are manifest at the surface in the form of depressions accompanied by both compressional and tensional movements must be investigated. Failure of mine cavities Any cavity beneath the earth's surface represents a condition of nonequilibrium and tends to close with time. Failure of cavities in a subcritical mine can be classified as failure of the roof, failure of the floor, failure of the supports (pillars, cribs), and elastic rebound of the rock. One mode of failure may dominate, but generally combinations of the four will occur. Also, initial instability of geologic materials and structural features and in situ stress may be unknown factors that can contribute to cavity failure. Roof failure is the most obvious opening stability problem in most operating mines. The bridging capabilities of the roof rock are dependent on the rock's material strength, geologic discontinuities, and reaction to humidity changes in the air of the mine. This type of failure may vary in magnitude from gradual flaking to massive roof falls. Mine floor instability can be manifest in the form of a general heave over a broad area or of individual pillar punching. The weaker and thicker the immediate floor material, the more prone a mine is to experience floor problems. Water seeping into an area can soften the underclay, reduce its bearing capacity, and precipitate failure of support for the overburden.

Jan 1, 1987

By S. S. Peng

An expert team consisting of geology, civil, and mining engineer professors was organized to investigate several problems of longwall mining in three mine sites including: site investigation, panel layout and support selection, overburden strata movement, surface subsidence and structural damages, slope, stability, well and stream dewatering, etc. The results include the development of a PC software package such as powered support selection, Surface subsidence prediction, optimum panel dimension, and pillar design.

Jan 1, 1989

By S. Harteis, D. Dolinar

Many mining operations rely on bulkheads to provide a barrier between impounded water or slurry and active mine workings. However, bulkhead failures could cause and have caused catastrophic flooding that puts the underground workforce at risk. The National Institute for Occupational Safety and Health1 (NIOSH) in collaboration with the Mine Safety and Health Administration (MSHA) is conducting research to evaluate the adequacy of existing design practices for water and slurry bulkheads. A key component of this research effort is documentation of the performance history and design parameters for bulkheads installed in underground coal mines during the last 20+ years. This research is part of a larger effort to develop general design guidelines and procedures for constructing and maintaining bulkheads that will help to ensure their long-term structural integrity, while significantly reducing the inundation risk for miners. This paper presents the results of NIOSH’s research to identify state-of-the-art bulkhead designs, including design criteria, leakage monitoring systems, and emergency warning systems. Underground observations and evaluations of existing bulkheads suggest that the most important design factors influencing their performance are the interface of the bulkhead with the surrounding strata and the potential magnitude of the hydraulic pressure to which they may be subjected. When a bulkhead has failed, leakage has generally been through the surrounding strata or along the bulkhead/strata interface, with the failure potential along the interface increasing with hydraulic head. It was also determined that development of a monitoring program to ensure the long term structural integrity of bulkheads is an important safety consideration. Current monitoring procedures range from weekly visual inspections to constant monitoring via pressure transducers and fluid level indicators. Where possible, these devices are used in conjunction with a computer based mine monitoring system to alert mine personnel when an emergency condition exists at the bulkhead installation.

Jan 1, 2006

By Kenneth L. Clifford

We have heard a lot during the recession about the mining industry now competing in a world market for sales of such commodities as copper, lead, zinc, molybdenum, iron ore, and phosphate. If the US mining industry is to survive, do we not have to reduce direct operating costs to a minimum? How significant is labor cost compared to consumables, electricity, fuel, maintenance, materials, and transportation? Ore grade is ultimately the dominant factor; but can we offset its dominance by improving productivity? Is the US mining industry perhaps only 25-30% as productive as it might be? What prevents the industry from being more productive? Can we retrofit older plants to be more productive or does it take a new plant with the largest equipment available for each unit operation? Present Productivity A copper industry study a few years ago covering 13 US copper concentrators ranging from about 10-45 kt/d (11,000-50,000 stpd) showed that productivity did not vary with capacity above 18 kt/d (20,000 stpd). Operating productivity varied significantly, however, from about 180-635 t/d (200-700 stpd) per operating hourly worker in most units, but one operation had 900 kt/d (1,000 stpd). Concentrators with more than 545 t/d (600 stpd) per hourly worker were in the 23 kt/d (25,000 stpd) or more category. Maintenance productivity covered about the same range, but most concentrators were in the 270¬635 t/d (300-700 stpd) per hourly maintenance worker. Again, one company stood well above others by exceeding 1 kt/d (1,100 stpd). Total hourly productivity ranged from 125-280 t/d (140-310 stpd) per hourly worker for most concentrators. The company with the most productive operation, however, was above 455 t/d (500 stpd). Salaried worker productivity for most operations was in the 725 t/d to 1.2 kt/d (800-1,300 stpd) per salaried worker range. The best operation was 2.2 kt/d (2,400 stpd). The end result is, of course, tons per day per total personnel. Most operations fell in the 110-220 t/d (120-240 stpd) range with the best operation at 400 t/d (440 stpd). This is the productivity definition that is most directly related to unit costs. Many of these operations have made some productivity progress

Jan 11, 1983

Minter's excellent recent paper (1990) on the great paleoplacer gold deposits of the Witwatersrand and its subsequent discussion (Cheney, 1991) and reply (Minter, 1991) provide an informative, interesting and up-to-date review of the geology of these unique ores and some long enigmatic questions regarding their origin. Clearly summarized by Minter (1990) are the compelling sedimentologic and stratigraphic relationships indicating that the Central Rand Group conglomerates are ancient placers and documenting the fluvio-deltaic conditions of their transport and deposition. As Minter (1991) indicates, Cheney's discussion broadens Minter's 1990 coverage of the subject by considering higher and lower stratigraphic units and by adding detailed comments concerning recently published information about metamorphic and age relationships. Cheney's discussion and Minter's reply also readdress the old controversy of epigenetic hydrothermal vs. syndepositional placer origin for these deposits. This controversy has been resurrected by: • recent recognition of hydrothermally-altered granites (HAGS) in the older Archean basement rocks that lie unconformably beneath the Dominion Group rocks north and west of the Witwatersrand depository; • recognition of greenschist metamorphic assemblages in the Witwatersrand strata; and • the presence of interpreted epigenetic mineral assemblages and textures in some of the deposits. These relationships are briefly discussed by Cheney who suggests they may result from later, superimposed metamorphism and deformation and do not prove that the gold was introduced by epigenetic/metamorphogenic processes. In his paper, Minter (1990) cited the 1986 abstract of a subsequently published paper by Hutchinson and Viljoen (1988) who also considered many of these aspects and attempted to integrate them all in a broad genetic hypothesis. In addition, they emphasized the additional important problem of determining the source for the Witwatersrand gold. Although this issue was partially addressed by Reimer (1984), Mossman and Dyer (1985) and Reimer and Mossman (1990), the question has been generally under emphasized in earlier work and was, therefore, not stressed by Minter and Cheney. Reimer (1985) also emphasized the importance of weathering of the basin's Archean hinterland and its affects on Witwatersrand sedimentation.

Jan 1, 1992

By Ernest H. Gilmour

Introduction An earth science teacher at the junior high school level has the potential of working with 450 students per year. And these young people are at an age where they can be properly informed about the importance and necessity of a viable, domestic minerals industry. It was with this philosophy that a course was developed at Eastern Washington University for science teachers entitled, "The Minerals Industry of the Pacific Northwest." The two-week course is sponsored by the Northwest Mining Association and the Idaho Mining Association. It was first taught during the summer of 1971 and will be taught during the summer of 1986 for the sixteenth straight year. It is a two-week mini-version of the very successful six-week course taught at the Colorado School of Mines. In 1971, Russ Chadwick was President of the Northwest Mining Association. He suggested that such a course should be developed and offered at Eastern. It was determined that tuition scholarships would be necessary. This would ensure that a reasonable number of teachers would register for the course. Say a teacher has five classes a day, with 30 students in each class. Some 150 students come into contact with that teacher each term. With three terms a year, an instructor works with about 450 students per year. So, with 20 teachers in the course, it was possible that 9000 students a year could learn from our informed teachers. And now, after 15 years, more than one million students have learned about the minerals industry from teachers who have completed the course. The course is designed to take the teachers through all aspects of the minerals industry. Most of the lectures are given by the working geologists and engineers who talk about their jobs in the industry. This is extremely important for the credibility of the course. The teachers realize that the lecturing professionals really believe in what they profess. Industry fundamentals The course begins with a lecture on the need for minerals in a modern society. Most of the teachers do not realize how critical minerals are to every aspect of our daily lives. The lecture is followed by the US Bureau of Mines movie, "Minerals Challenge." An exploration geologist then discusses what is involved in the exploration for a mineral deposit. This includes a case history that the geologist has actually been involved in

Jan 9, 1985

By James D. Bennett, David L. Passmore

The Coal Mining Section of the National Safety Council has selected hand and finger injuries of coal miners as a major topic for safety policy analysis [Tinney, 19851. So that facts about hand and finger injuries can be used to guide the Committee's work and the work of policy-makers, we present in this paper detailed analyses of the following factors related to hand and finger injuries received by coal miners between 1975 and 1983: occupations of the injured miners; activities of the miners when injured; machines involved; objects, substances, exposures, or bodily motions that produced or inflicted the injuries; principal physical characteristics of injuries; and events directly resulting in the injuries. We analyze hand injuries for underground and surface mines separately because of the remarkably different production methods and hazards evident in these two operating environments. We do not, however, provide information about hand and finger injuries that occurred in surface operations at underground sites (such as yards, shops, or tipples), auger mining operations, culm banks, dredging operations, mill or preparation plants, offices, or laboratories. The most comprehensive information about the safety of coal mining in the United States is assembled by the Health and Safety Analysis Center (HSAC) of the Mine Safety and Health Administration of the United States Department of Labor. Mine operators in the United States report to HSAC all injuries as required under Part 50, Title 30, of the Code of Federal Regulations (mandated by Section 103 the Federal Mine Safety and Health Amendments Act of 1977, Public Law 95-164). These regulations stipulate that injuries must be reported within ten days of their occurrence on a United States government data collection form, Mine Accident, Illness and Injury Report (MSHA Form 7000-1). HSAC maintains these data primarily to produce federally-mandated reports. HSAC resolves injury reports that are incomplete, checks for errors in data coding and handling, and corrects inconsistencies in data submitted by

Jan 1, 1986

By Robert Seal, James J. Rytuba, Michael P. Hunerlach, Roger Ashley, Dean Enderlin

Cessation of open pit mining in 1996 at the McLaughlin hot-spring type gold deposit has resulted in the development of two pit lakes that will be managed as permanent lakes with a total storage capacity of 10,000 acre-feet. The south pit lake developed from alkaline ground water infiltration and had an initial pH of 7.2. It contained high concentrations of boron and chloride, indicating that it consisted of a mix of meteoric and evolved connate waters. Numerous carbon dioxide gas vents were present initially in the shallow south pit lake, but as the pit lake increased in depth, gas venting ceased and carbon dioxide is retained in the lower layer of the chemically stratified lake. Continued increase of dissolved carbon dioxide in the lower layer increases the potential for a limnic eruption. A chemocliie occurs at a depth of 47 m. The upper layer of the lake has lower pH, temperature, and conductivity than the lower layer. Effluent from ore and waste rock piles directed into the south pit lake has contributed to high levels of antimony, arsenic, cobalt, copper, mercury, nickel, and zinc. Iron oxyhydroxide is abundant in mine effluent that enters the pit lake and adsorption onto this phase has lowered the arsenic and antimony concentrations. The north pit lake formed primarily from winter surface runoff. The acidic water, pH 4.56, initially had high concentrations of cobalt, mercury, and nickel derived from weathering products of pyritized serpentinite and shales. Buffering by wall rocks has established a slightly alkaline pH of 7.62. The concentrations of trace metals in the north pit lake have substantially decreased due to continued dilution of the lake by surface runoff and remediation of parts of the pit walls that have been covered by backfill.

Jan 1, 2000

By K. H. Bettles

The Goldstrike Mine is in Eureka County, Nevada, approximately 23 miles northwest of the town of Carlin. The mine is at the north end of the Tuscarora Spur of the Tuscarora mountains at elevations of 5,300 feet to 5,950 feet. It is part of the Bluestar/Goldstrike subdistrict of the Carlin Trend of sediment hosted, disseminated gold deposits. The Carlin trend comprises over 15 major gold deposits with production, reserves and geological resources of approximately 50 million ounces of gold (Fig. 1). In the Bluestar/Goldstrike subdistrict, there are four major deposits; Post, Betze, Bluestar, and Genesis, plus numerous smaller deposits (Fig. 2). Betze is located on property controlled by Barrick Goldstrike Mines; Bluestar and Genesis on Newmont Gold property; and Post on both Barrick and Newmont property.

Jan 1, 1989

By E. G. Joe

Introduction Uranium mining has been actively carried on in Canada for over twenty-five years mainly in Ontario at Elliot Lake in Bancroft, and in Northern Saskatchewan at the operations of Eldorado Nuclear Ltd. During this period, approximately one hundred million tons of mill tailings have been created and lie mainly on the surface close to the mining and milling installations. If all of the present reserves of uranium ores in Canada are mined, a further eight hundred million tons of mine tailings could be created in Ontario, Saskatchewan, the Northwest Territories, Labrador and British Columbia. The present control of effluents from existing tailings is managed by the operating mines under the regulations of both federal and provincial control agencies. It is the long-term effect of these tailings on the environment that has not been clearly defined or measured. Several years ago, the Canada Centre for Mineral and Energy Technology of the Department of Energy, Mines and Resources, embarked upon a program of research directed at the long-term effects of uranium mine tailings. Much of the research to date has been on tailings in the Elliot Lake, Ontario area where some tailings piles have existed for over twenty years. The program consists of two projects, one to address the problem of existing tailings and the other to examine alternatives in the management of current and future tailings. TAILINGS CHARACTERISTICS Uranium tailings are generally the result of sulphuric acid leaching in which the uranium and alkaline minerals such as carbonates and chlorites are leached out. If the ore contains sulphides, these minerals potentially present an additional hazard of acid generation over time and could increase the mobility of heavy metals and some radionuclides such as thorium-230. Except for the extracted uranium, most of the radionuclides from the ore remain in the tailings pile. Uranium ores in Canada span a wide variety of mineralization and associated gangue composition. Uranium minerals include pitchblende, uraninite, uranothorite, brannerite and the phosphate uranium minerals autunite and saleeite. Gangue minerals vary from acid consuming carbonates, and chlorites to acid producing sulphides and inert silicates. With such a wide variety in ore composition, the characteristics of the resultant tailings in Canada are in many cases site specific and the solutions to the long-term effects of these tailings may also be site specific.

Jan 1, 1980

By E. M. - Technical Papers Spokes

A.B. Rushton The work and resulting publications of the Committee on Underground Coal Mine Safety and The President's Commission on Coal were appropriate at the time of publication. The statistics cited in the publications appeared to be valid, and the actions proposed by the committee had merit. Although the actions proposed to achieve further improvements in safety may have been worthy of repeating, to associate statistics (Table 1) that are seven to eight years old with company names is bad. Statistics are readily available from the Mine Safety and Health Administration (MSHA) and are usually within three months of being up to date. A review of the most recent Coal Controlling Co. Report will reveal significant changes in the rankings. The more recent statistics should have been cited, or at least compared with those used in the article. Westmoreland Coal Co. is listed as having a 21.1 disabling injury rate for the period (1/78-12/80). The disabling injury rate for the Company for the year ending 1985 was 7.4 for the underground mines and 6.25 for the total facilities. Through March 1986, the dis- abling rate for the underground mines was 2.86. The management and employees of Westmoreland Coal Co. are committed to safety. The improvements achieved over the past five years attest to that commitment. We are justifiably proud of the achievement and dislike being misrepresented as in the April 1986 article Reply by E.M. Spokes I regret that the lapse of time since the original work gives a picture of safety practice that is no longer true today. That Westmoreland has reduced its disabling injury rate by 65% to 7.4 in 1985 is most commendable, and I congratulate you for this accomplishment. It is typical of the progress sought by the National Academy of Science and National Research Council when they funded the original study and published "Toward Safer Underground Coal Mines" in 1982. If a significant number of companies do as well as Westmoreland, the report will have served its purpose. Since the paper was primarily a condensation of the original report, it would not have been appropriate to seek access to later statistics. The purpose in seeking its publication at the time of presentation was to reach those who might have missed the original report. The fact that publication was delayed has not voided that purpose and may have reminded some who were aware of the original that its recommendations still apply to improvement of underground coal mine safety

Jan 1, 1987

By W. A. Crosby, A. Bauer

The recent increase in size of open-pit mining operations has resulted in major improvements in efficiency that are beneficial to the mining industry. Higher bench heights, larger diameter blast-holes, and more powerful explosives have played an important part in reducing mining costs. These measures have, however, resulted in increased energy concentrations in the blast area, which can result in severe backbreak problems for final pit walls. If backbreak is not controlled, a decrease in the overall pit slope angle will ultimately be necessary, and consequences, such as decreased recoverable ore reserves and increased waste-to-ore ratios, will result. Greater amounts of loose face rock will be produced, and planned safety berms will be less effective or nonexistent. Hazardous working conditions can also result. Remedial measures, such as scaling large areas and using wire mesh or other artificial support, are very expensive and difficult to implement. There must be a trade-off between the money saved by using larger blasts, and the time spent to maintain pit wall quality. The best approach is to control blasting effects so the inherent strength of pit walls is not destroyed. Presplit and trim blasting are two control blasting techniques used to accomplish this. Both techniques are designed to create a low explosive energy concentration per unit of wall area at the perimeter of the pit. Energy concentration of the main production blast must also be controlled to avoid damaging the final pit wall. A low-energy concentration at the final wall can be obtained by decoupling charges, decking charges, using less powerful explosives, decreasing blasthole diameters, and changing the burden and spacing. Rock properties-fragmentation and the final state of the pit wall-influence the success of a blast. The most important properties to consider are: in situ rock strength, and the nature, frequency, and orientation of structural features. Since these variables cannot be controlled, they must be evaluated by suitable field tests, and then a control blast can be designed using controllable variables, such as spacing, burden, and hole diameter, so the blast will be successful for appropriate rock conditions. One main thrust of this article is to show how control blasts can be successfully designed for larger hole sizes, up to 381-mm-diam. While efforts to produce a good control blast in shovel/truck pits have been ongoing for some time, recent use of a modified presplit technique in dragline operations has helped reduce water problems, improve safety, and increase dragline productivity in a number of surface coal mines. This article presents examples on designing and using presplit and trim blasting for large hole operations in both shovel/truck and dragline mines. Presplit Blasting As introduction to presplit blasting, Fig. 1 illustrates a typical presplit blast layout using 102 mm-diam presplit holes for 381¬mm-diam production holes. For this type of blast, presplit holes would normally be drilled first, ahead of main production holes. The choice can then be made between loading and firing the presplit line or infilling the main blast. In the latter case, the presplit line would be fired instantaneously 100-150 milliseconds before the main blast. As shown in the figure, the presplit line is formed ahead of the main

Jan 2, 1982

By K. Hansson, A. Jedborn

Background In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore, from a system's point of view, slurry transporta¬tion would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alterna¬tive examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasi¬biliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. l . The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then tempo¬rarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situ¬ations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe¬line, and personnel are needed. In a first phase (Fig. 2), a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1989