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By Jalal Vakili

In designing very large excavated slopes which are becoming increasingly common in both mining and civil engineering projects, the most important decision of the engineer is the selection of a slope angle. The purpose of the geotechnical study of the slope is to ensure reasonable stability of the slope in the most economical way. Because the rock mass composing each slope is unique, there are no standard solutions for slope stability analysis. A practical solution is formulated from the basic geologic data, engineering properties of the rock, geometry of the slope, and ground water observations. Stability analysis consists of analyzing the forces causing and resisting slope failure. There are several methods of slope stability analysis. The limit equilibrium method evaluates the overall stability of the sliding mass just on the verge of slip, using some or all of the three equations of static equilibrium of a plane problem. The stress-strain relationships of the rock mass are not considered. Method of analysis include the assumption of the shape of the slip surface. Peterson (10) introduced the Swedish method of analysis in 1916, which is based on cylindrical sliding surfaces. Fellenius (4) in 1926 extended this method and introduced the method of slices. This method was refined by Bishop (I) in 1955 and later used by Morgenstern and Price (8) in 1965. In many cases layers of extremely soft strata or strata weakened by neutral stress cause failure to take place by translation along plane of weakness instead of by rotation on a circular surface. In this paper, a slope with a horizontal layer of weak meterial is selected for analysis. This is a usual case for open pit lignite mines where lignite seam may be non-cohesive and 'weaker than the overburden rock or a soft clay layer underlies the seam. It is also a common case of rock fill dams supported on a foundation stratum of clay where the strength of the clay layer is less than that of the rock fill. It is shown that utilizing this method, the location of potential failure surface and the value of the minimum factor of safety can be calculated directly and the trial computations of factors of safety for a number of assumed failure surfaces are not necessary.

Jan 1, 1982

By Thys B. Johnson

This paper presents a new algorithm for solving the Controlled Flow Mine Ventilation Problem in an optimal manner - minimum power consumption. The use of booster fans and restrictions on the locations of ventilation regulators are dealt with in a straightforward manner. This new approach is based on a common Operations Research algorithm for which there are many fast computer programs available.

Jan 1, 1982

By Richard L. Bullock

At the beginning of the previous chapter, variation in the types of room-and-pillar stopes were briefly identified so that the reader could begin to understand the extensive application of this mining system. These same variations will be reviewed and elaborated upon in this chapter while discussing the methods by which the material is extracted. FULL-FACE SLICING BY DRILLING AND BLASTING When the entire mineralized thickness is taken in one pass of mining, it is known as full-face slicing. There is no mineral of economic value intentionally left either in the floor or the back. Theoretically, there is no limit as to how high the face could be in a single pass. But there are practical limitations to equipment size, pillar stability, and control of local loose slabs on the face, roof, and pillars that dictate current practices. The range of face height of the 15 typical room- and-pillar mines covered in Table 8 (Sec. 1, Subsec. 7) from the Dravo report, 1974 was 1.7 m (5.5 ft) to 9.8 m (32.1 ft), the average being 5.2 m (17.1 ft). In the Gaspe copper mine, Murduchville, Que., the drill jumbos were constructed with an extendable tower for full-face mining up to 15.2 m (49.9 ft) if necessary. However, a more common practice at that mine is to take the ore in multiple slices (Hall, 1959). In many metal mines the normal practice when starting stoping is to drive a single development drift into the ore zone a distance that will allow at least four or five rooms to be opened on each side of the initial drift. Since this opening will probably serve as the main haulage drift, it should be kept as straight and as level as possible. As to the other criteria for the stope developments, it might be well at this point for the mine planner to review the remarks presented earlier (Sec. 1, Subsec. 7) on drifts, entries, and crosscuts for production. After the initial drift is driven, it should be slabbed to the full room width if it is not that wide already. Next, the future pillars should be marked on the ribs and the rooms driven between them. If the ore extends beyond the length of the initial drift and is to be mined in the normal sequence of mining, care should be taken in a random room-and-pillar system to see that the extension of the initial roadway remains straight. If pillar "spotting" is left to the shift foreman or to the miners, they probably will not realize the importance of maintaining a uniform pillar line next to the roadway, and of keeping the line straight. Invariably, a pillar will end up in the middle of the future drift extension resulting in a "dogleg" in a possible main haul- age road. While this error is not catastrophic, too much weaving in and out of pillars will certainly slow down future production haulage and needlessly increase costs. Face or Breast Drilling and Blasting Practices To drill and blast the initial advances into the rock, usually some form of cut pattern must be used. In room-and-pillar mining, when a pattern is drilled in a face and that face is the only exposed surface to which all of the rock must be broken, the pattern is known as a "face round" or "swing." The face to which the rock must be broken is referred to as a "free face." Thus the most common way to advance a room into virgin rock with only one free face is by drilling swings. If, after breaking around a pillar, taking down back, or taking up bottom, there is a second (or third) face exposed, a group of holes drilled nearly parallel to a free face and breaking to it is known as a "slab round" or "slabbing." In some mining districts it is also known as "slashing." Obviously, because two faces exposed offer less resistance to fragmentation than a single face, slabbing requires less drilling and fewer explosives than required to break the same tonnage with the same degree of fragmentation with swings. Therefore, it behooves the driller (or supervisor) to plan the combination of

Jan 1, 1982

By Davies L, Jeffreys J

The Broken Hill Proprietary Company Limited (BHP) has been successfully using mini computers in its open cut strip and pit coal planning for the past six years. A recent development within the computerised Seam deposit User oriented Mine Planning System (SUMPS) provides the facility for the generation, evaluation, and simulation of colliery workings.Existing computer programs for collieries were studied, but none met the requirements of BHP's underground mine planners.Consequently, it was decided to develop programs, based on those existing within BHP for open cut strip and open cut pit mine planning. The new planning facility was designed to enable any number of proposed colliery developments and operating plans to be evaluated against optimistic, most likely and pessimistic geological lease area interpretations.In mine planning, sensitivity studies should be conducted to determine the likely effects on production, development costs and colliery economics of changes in mine design and budgeted equipment productivity. By automating the manual calculations involved, the computer programs provide the ability to conduct numerous and extensive sensitivity studies. Hence an optimum mine design for budgeting purposes can be derived quickly with an improved degree of confidence in the long term viability of the proposed development.

Jan 1, 1982

Over the past few years, the School of Mines of Paris has carried out a program of research* on the grade-tonnage relationships in the border zones of porphyry copper deposits. The initial problem was quite difficult : how does the grade vary at the border of the known orebodies? Does it decrease regularly, or do some thresholds exist, where the grade diminishes abruptly? The answer to these questions have important repercussions on the prediction of the future reserves in the fringe of the orebodies. If the grade decreases regularly, very important reserves could exist; if not, there would be strong limitations to the possible recoverable reserves. The first step of our study was a critical study of previous work done on the subject. It appeared that, following Lasky's pioneering paper (Lasky, 1950), most of the authors used what we can call an historical approach to the problem. In this, the grade-tonnage relationships are calculated from the available figures for production and reserves and the results are then extrapolated to the future. The strong limitation of this model are brought out in this article.

Jan 1, 1982

By J. Wong

We report a successful use of x-ray absorption spectroscopy to probe the chemical and structural environment of trace elements in coal. When applied to vanadium in coal, it was found that vanadium exists in at least two environments: a trivalent site 6-fold coordinated by oxygens at ca. 2.0~ and a tetravalent site 5-or 6-fold coordinated also by oxygens, one of which forms a short V=O bond at ca. 1.6a. No evidence of nitrogen (porphyrin)1'or sulfide environment was found. It was also found that the vanadium environ-ments in the raw coal did not survive unchanged in a liquefaction process. These findings have implications for coal cleaning processes and for trace element release into liquefaction process streams.

Jan 1, 1982

By K. P. Sinha, J. F. Schatz, T. F. Borschel, S. Demou

Triaxial creep behavior of oil shale was investigated in the laboratory at simulated in-situ conditions. A range of temperature and stress conditions were chosen to represent those within the inter-chamber pillars of a modified in-situ retorting operation. Oil shale in grades up to 60 gallons per ton have been tested. Experimental data and a constitutive model have been presented here.

Jan 1, 1982

Mining operations have been carried out at Mt. Lyell for the past century. Until the 1970's the location of the mines in relation to the surface topography has enabled the mines to drain their copious inflows of water by gravity. Deeper mining has forced the Company to undertake pumping of increasing inflows of water. A major pumping facility is being established near the base of the known ore reserves with a view to serving the future requirements of the mines. The system has a design capacity of 600 litres per second. Run of mine water will be treated in Vertical Flow Settlers to produce clear water and thickened mud, each of which will be disposed of separately. The clear water will be pumped by six 1000 kW conven- tional centrifugal pumps. The thickened mud will be removed by a Hydraulic Displacement System. The rising mains for both the mud and the clear water are stainless steel cased bore- holes.

Jan 1, 1982

By Christian F. Baiz

Industrial minerals can have similar financing requirements to those needs of metaliferrous and coal mining projects. Although, one may see many similarities in the fundamentals of the various risks that must be evaluated in lending to any specific project, one of the outstanding differences is the understanding of the many different markets that these industrial minerals represent. There are in excess of 30 principal industrial minerals ranging in terms of annual production from the largest quantity such as sand and gravel, to possibly the smallest quantity of production of a legitimate industrial mineral which may be defined as industrial and rough gem diamond. If we assume that it is necessary to finance independent projects of ore bodies of industrial minerals for development, the financing techniques are similar to the fundamentals that we would use for financing coal projects or metaliferrous deposits. Therefore, let's begin by redefining project financing in the traditional banking or private sector area, and then in the later development of this paper, we will apply these techniques to industrial minerals in general and cite it with several specific examples.

Jan 1, 1982

By Ernest W. Loya

A method for preparing gravimetric primary standards of helium in the parts per trillion range has been developed by the Bureau of Mines. The standards are used to calibrate a specially modified mass spectrometer used in determining the abundance of helium-3 and helium-4 in geochemical samples and of helium-3 in isotopically purified helium-4. The desired gas standards are prepared in 5-kg metal cylinders filled to 1,800 psia. Relative errors of less than 1 pct have been attained in the parts per trillion range using a successive dilution technique.

Jan 1, 1982

Topography, location of coal bearing land and existing built-up areas severely restrict locations for overburden disposal and dumping sites north of the Latrobe River clear of the coalfields are assumed until internal dumping areas can be established.

Jan 1, 1982

US 4,201,749-In the Bayer process production of alumina wherein aluminum hydroxide is precipitated from a sodium aluminate solution for use as seed, sodium oxalate and caustic components are removed from the precipitate by washing it first with a washing liquor which is saturated with sodium oxalate but has little content of caustic, and secondly with hot water. Substantially no sodium oxalate is recycled, yet loss of sodium aluminate is minimal. K. Yamada, T Harato, and H Kato, assigned to Sumitomo Aluminum Smelting Co Ltd May 6, 1980 8 pp. (423-127). Filed. 5-8-79 (37,145) Priority. Japan, 5-8-78 (53/54,803). Same¬British 2,020,642, dated Nov. 21, 1979.

Jan 1, 1982

By Mann LD

The Haveluck gold deposit, located 2.5km north of Meekatharra, contained 1,284,000 tonnes at an average grade of 1.8 g/t gold of drill delineated proved and probable ore prior to commencement of production. Following successful laboratory and pilot scale testing a decision was made to establish a 300,000 tonnes per year heap leaching operation. The ore, crushed to minus 100mm was stacked on an impermeable membrane and irrigated with a dilute cyanide solution. Gold was recovered from the resulting pregnant solutions by adsorption onto activated carbon and gold bullion produced in a modified Zadra process desorption plant. From the commencement of mining it was obvious that a high percentage of fine material was being produced which would adversely affect the percolation of the cyanide solution-through the leach heaps. A screening plant was incorporated into the crushing circuit which provided an oversize product satisfactory for heap leaching and an undersize for carbon-in-pulp treatment. Plant design provided for the slurrying of minus 0.8mm ore before feeding to the pachuca tanks. Initial practice was to use four tanks for gold dissolution and two for carbon contact. Vibrating screens were used through- out for the carbon-ore separation and the loaded carbon was stripped in the existing desorption plant.

Jan 1, 1982

By Kirby D. Mellegard, Ralph A. Wagner, Paul E. Senseny

Six creep laws for salt, each fitted to the same laboratory data base, are used in a numerical model of the deformation of an opening in salt to determine the influence of creep law form on the predicted closures. Results show that predicted closures can be very sensitive to the functional form of the creep law chosen to model the laboratory data.

Jan 1, 1982

Provide roof to floor closure rate measurements during underground mining, especially retreat mining of coal pillars, thereby helping the miners predict an imminent roof fall. Approach An electronic device and rugged potentiometric extensometer were developed to provide a means fur measuring total closure and the rate of closure. How It Works An intrinsically safe, battery operated instrument is enclosed in a rugged, hand carried field case. The unit is constructed using low-power, integrated circuits to provide minimal battery drain. Two digital Light Emitting Diode (LED) readouts are used; one displays the rate, the other displays the total closure. The unit will measure one extensometer continuously or two extensometers alternately. Two potentiometric extensometers with a closure measurement range of 6 inches are attached to the readout unit. As the roof to floor closure changes, the potentiometric resistance

Jan 1, 1982

By Gerald R. Coonan

Laboratory research is being conducted to substantiate suggested levels for voluntary standards on human response to transient vibration. The application of those findings to actual field conditions, however, will be limited because they will not include social and psychological factors existent in the real world. It is necessary, therefore, to examine available empirical information on complaints in an attempt to define the influence social and psychological factors might have on the laboratory results.

Jan 1, 1982

To develop a system which will effectively detect and extinguish fires in underground fueling areas while also providing a warning for nearby miners. Approach A system involving optical fire sensors. automatically-released dry chemical and aqueous film-forming fluid (AFFF) fire suppressants, and automatic alarms was designed from off-the-shelf, state-of-the-art components, and was in-mine tested at two mines in the two typical underground fueling area situations: fuel transfer areas and fuel storage areas. How It Works As mines become more mechanized, the number of fire hazards increases since fueling areas are more abundant. Fire hazards are especially serious in deep mines, since escape time is longer. To help solve this growing problem, the Bureau of Mines, in cooperation with the Mine Safety and Health Administration (MSHA) and the National Fire Protection Association (NFPA), has developed and in-mine tested a cost-effective fueling area fire protection system. The system was designed to be rugged enough for reliable, long-term use, yet to utilize state-of-the-art, off-the-shelf components to maintain cost-effectiveness. The system consists of cross-zoned, ultraviolet flame sensors used for quick response to fire, and a dual-agent fire suppression subsystem. A dry chemical is used for fast fire -knockdown," and AFFF (or "light water") is used for long-term, fire-securing capability. The basic components of the system are shown in the schematic illustration.

Jan 1, 1982

By Randal J. Barnes

Computer based techniques for solving the Ultimate Pit Limit Problem have appeared in mining literature since the early 1960's. The majority of papers pertaining to this subject can be partitioned into one of two categories: discussion of the pros and cons of a particular algorithm, or a case study presenting specific applications of an algorithm. Subsequently, there has a1 so been considerable debate concerning the best techniques between, essentially, two camps: those favoring the moving cones techniques, and those favoring the techniques which yield true optimal pit designs. The purpose of this paper is not to add to this debate, but rather to present a few new ideas (partially a new look at an old idea), which can be easily implemented to noticeably improve any ultimate pit limit approach independent of the particular algorithm used. These ideas fall under the heading of "bounding the optimum" (Barnes and Johnson, 1980).

Jan 1, 1982

By Robert L. Grayson

An application of goal programming (GP) in evaluation of proposed purchases of major mining machinery is presented. It incorporates historical operating data compiled on mechanical delays to production time. The goal programming technique, which is an extension of integer programming, is used to maximize an objective function. The objective function combines the capital budgeting evaluation methods of net present value, payback period, profitability index, and total net earnings as goals weighted according to a management preference. The GP method gives an optimal solution to a multiple-goal problem. A sensitivity analysis is performed to determine the effect of an increased production forecast or a tightened budget on proposal selection.

Jan 1, 1982

By T. P. Lim, N. K. Davé

"A modified accumulator method was devised to measure surface radon flux from porous media. It consists of a collector hood type of accumulator with a homogenizer pump and a pressure balancer. The latter was used to compensate any pressure changes produced during sample withdrawal and hence eliminate additional radon transport from the media. Errors resulting from radon back diffusion due to decreasing concentration gradient with time across the interface were also eliminated by measuring radon concentrations in the accumulator for two different periods of time and calculating the true radon flux, J(o), from the solution of a time-dependent radon transport equation in the absence of any pressure and temperature gradients.Core samples from the observation site were also withdrawn for laboratory determination of porosity, moisture content, bulk and grain densities, Ra-226 source concentration and the fraction of Rn-222 produced in the void volume. The data were used to calculate the radon diffusion coefficient for the media.Using the present method, the true value of surface radon flux for an abandoned uranium tailings pile was obtained as J(o) = 2020.2 mBq/m2/sec (54.6 pCi/m2/sec) compared to the value J = 1406.0 mBq/m2/sec (38.0 pCi/m2/sec) obtained by the conventional method. The bulk diffusion coefficient of radon for the media was obtained as 1.2 x 10-7 m2/sec.IntroductionRadon-222 (hereafter called radon) released from uranium mill tailings is a continuous source of environment radioactivity in terms of its transport as an airborne gas and its decay products. The particulate daughter products readily attach to aerosols or dust particles and can further be transported by various other environmental pathways. Typical radon release rates from existing tailings piles vary between 740 and 14,800 mBq/m2/sec (20-400 pCilm2/sec), depending on the nature of the tailings material, moisture content, etc. The United States Nuclear Regulatory Commission (NRC) is proposing radon release rates of 74 mBq/m2/sec (2 pCi/m2/sec) above that of background."

Jan 1, 1982