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By William Lyman

GENERAL DESCRIPTION Shrinkage or shrinkage stoping refers to any mining method in which broken ore is temporarily retained in the stope to provide a working platform and/or to offer temporary support to the stope walls during active mining. Since ore "swells" when broken, it is necessary to shrink the muck pile a corresponding amount by draw¬ing some of the broken ore out as the stope is advanced-hence the name. Broken ore retained during stoping is drawn out after the stope has reached its limits. The stope may be left empty or may be filled with waste contemporaneous with, or subsequent to, the final draw. Traditionally the method implies conventional overhand stoping methods with miners working between the muck pile and the stope back, in a space which advances updip with mining and is maintained by balanc¬ing "swell" with "shrink." The shrinkage classification is also applicable to so-called "semishrinkage" methods in open pillar-supported stopes where broken ore is temporarily retained as a working platform but offers no wall support; and to various blasthole shrinkage methods which utilize broken ore temporarily retained in the stope for wall support, but which do not require miners to work from muck pile in the stope. The method is generally applied to steeply dipping veins of strong ore between strong walls. APPLICATION Geometry The geometry of a shrinkable vein is described in terms of dip, width, and regularity along dip. Overall strike and dip dimensions and irregularities along the strike generally impose no restrictions on the method. Dip is ideally 1.2 to 1.5 rad (70 to 90°). As dip falls below 1.2 rad (70°), the shrinkage draw begins to strongly favor the hanging wall side, thus leaving a poor working platform for conventional overhand work. This is particularly true in relatively wide stopes. The sup¬port afforded to the hanging wall also diminishes with decreasing dip, reaching nil as the dip approaches the repose angle of broken ore. Dips below 0.78 to 0.87 rad (45 to 50°) are not generally shrinkable except by open stope "seinishrinkage" methods. Minimum mining width is fixed by working space requirements in the stope-generally about 1 m. Shrink¬age in narrower veins requires that waste rock from one or both walls be broken with the ore and the attendant dilution accepted to achieve the minimum width. Nar¬row stopes are less suitable, encouraging hang-ups and bridging of broken ore, with the attendant problems of erratic draw and incomplete recovery of broken ore. Maximum practical width may be 3 m or less to over 30 m, depending upon the competency of the ore and its ability to stand unsupported across the stope back. This is a vital safety consideration in conventional over¬hand stopes, but is much less of a factor in blasthole shrinkage methods. Very wide veins and massive ore bodies have been mined by transverse vertical shrinkage panels separated by transverse vertical pillars which are either abandoned or recovered later by other methods. Regularity along the dip is a prerequisite of shrink¬age as there must be no serious obstruction to the flow of broken ore downward through the stope to the sill level. Gentle rolls along the dip are acceptable if the local footwall dip everywhere exceeds 0.78 to 0.87 rad (45 to 50°). Off-dip hanging wall and/or footwall splits can generally be mined selectively from a conventional shrink stope as they are encountered without ad¬versely affecting subsequent continuation of shrinkage mining updip on the main vein. Vertical offsets or major rolls along the dip which cannot be "smoothed over" generally require that a sublevel be established with new draw control development. Blasthole shrinkage methods are much less flexible (and thus less selective) in their ability to accommodate any of these irregularities. Ground Conditions The wall rock must be strong enough to stand with the minimal support afforded by the dynamic mass of broken ore in the stope. During active mining, local sloughing from the walls is restrained, but the broken ore affords little, if any, useful resistance to closure of the stope walls. Such squeezing, if present, may bind up the stope and cause the loss of much ore. Pillars left between and/or within stopes are effective in preventing closure but reduce overall recovery. Walls may be re¬inforced by bolting after each stope cut in conventional shrinkage but not in blasthole shrinkage. Ore in place must be strong enough to stand with no natural support across the stope width, although tem¬porary artificial support or reinforcement may be used locally in conventional stopes. Some spalling or sloughing is permissible in blasthole shrinkage as men are never present in the stope. Physical and/or mineralogical characteristics of the broken ore may impose restrictions on stope design and/or operational plan¬ning, and may even preclude the use of shrinkage al¬together. Examples include: ores which, when broken, are cohesive or which tend to pack or cement together under the influence of ground water, wall pressure, and/ or chemical reaction. Such conditions precipitate er¬ratic draw during mining and often result in difficult and/or incomplete final draw; pyritic ores which oxidize very rapidly in the stopes and may generate heat, imposing a fire hazard by spontaneous combustion; sulfide ores which oxidize sufficiently in the stopes to adversely affect mill recovery by flotation; and ores (es¬pecially those containing uranium minerals) which ex¬ude radon gas and thereby impose ventilation constraints on stope design. In most cases these problems can be minimized by limiting the size of stopes, by minimizing the duration of mining activity in each stope, and by promptly drawing each stope empty following comple¬tion of mining.

Jan 1, 1982

By Tom C. Palanglo, Norbert Maerz

Control of crushing and grinding circuits ideally requires continuous particle size measurements on both in- and outgoing streams. Until recently, size distributions could only be obtained by screening methods, which were neither cost effective, frequent enough, nor timely. Input streams are often composed of large fragments, making screening prohibitive. Optical image processing systems, such as the WipFrag System, have been developed to the point that they are an extremely valuable tool to characterize, control, and fine-tune the comminution process. Although accuracy is less than that of screening, measurements are precise enough to detect relatively small changes in size distributions.

Jan 1, 2000

By Zhenqi Cai, Mike Lichte, Kimberly Kennedy

The Ohio River Tunnel is the first of three combined sewer overflow tunnels to be constructed under Allegheny County Sanitary Authority's Clean Water Plan. The project consists of a total of approximately 5.5 miles of 18-ft and 14-ft finished diameter, segmentally lined deep rock tunnels and smaller diameter dewater and flow connector tunnels, as well as 8 deep flow drop and access shafts varying from 25-ft to 70-ft in finished diameter and 130-ft to 170-ft in depth. This paper describes the project components and the primary engineering challenges faced during design, including tunnel alignments and shaft site optimization with tight right-of-way constraints and multiple railroad and highway crossings. Assessment of tunneling methodology as well as tunnel and shaft designs and constructability included overcoming 5-bar hydrostatic pressure, abrasive bedrock, and the selection of the appropriate support of excavation systems in high-permeability alluvial soils adjacent to the Ohio River.

Jun 23, 2024

By P. K. Muduli

The number and the type of trucks and shovels are two important factors in determining the optimal design parameters of an open-pit mining system. Characteristics of trucks' arrival and service-time distributions determine the performance measures of a truck-shovel system when it is modeled as a closed queuing network. In all of the previous analytical queuing studies of truck-shovel systems, trucks are assumed to be identical with respect to their arrival and service-time distributions. This assumption, although desirable for reducing computational difficulties, causes either underestimation or overestimation of the performance measures depending on the number and type of equipment used in the model. The relative error in performance measures could be significant when the number of different types of trucks and shovels is large. In this paper an analytical method is proposed to model the truck-shovel system as a multiple-chain closed queuing network. This approach would provide the capability of estimating system performance measures for planning purposes when the truck fleet is not necessarily composed of identical trucks. An example of truck-shovel systems is discussed to show how the incorporation of different types of trucks does affect performance measures.

Jan 1, 1996

By S. E. Knalafalla

The availability of a fluid whose gross behavior sharply changes in a magnetic field without affecting its rheologic characteristics has been recently demonstrated. A colloidal solution of a ferrimagnetic material, such as magnetite, in a liquid that is immiscible with and lighter than water is projected for use in selective separation of particles according to their density spectrum in a continuous fashion. Magnetic levitation forces were created as an antigravity field when the fluid was properly placed in an inhomogeneous magnetic field. Five spherical non¬magnetic objects with densities ranging from 2.70 to 11.35 gram cm ' were made to float or submerge at will in a 0.96-gram cm-3 density fluid placed at vertical magnetic field gradients ranging from 0.2 to I kilooersted cm-1. The magnetic levitation farce in dyne cm-3 was found to vary linearly as a function of the object-fluid density mismatch with a slope equal to one g, the acceleration due to gravity.

Jan 1, 1972

By C. M. Chang, H Gu, J. A. Barrera-Godínez, T. J. O’Keefe

The inability to economically separate Fe3+(aq) from a variety of liquid effluents is a major Concern for the primary metals industry. A unique solvent extraction process, galvanic stripping, has been developed to address this problem. The galvanic stripping of iron from D2EHPA using only a 5~10 g/L H2SO4 solution is achieved after Fe3+ is directly reduced to Fee, in the organic phase using zinc metal additions. To date, the primary focus of the laboratory research has been on optimizing the parameters that control iron removal. The results from these studies have now been, used to make a preliminary design for a continuous flow process that is envisioned for zinc industrial applications. Specific quantitative data are shown for the use of a Zn-Ph cellos° in the iron-reams al process. Some introductory results on electrochemical potential measurements and impurity distribution with solution cycling are also giiven.

Jan 1, 2000

By L. A. Garfield

Since the start of mining operations at White Pine in 1953, over 22 miles of drifts have been driven for the purpose of developing the ore body. The drifts, normally driven in multiples of three or five headings, serve as conveyor belt routes, ventilation ducts, equipment and haulage roads, and initial access to mining blocks. Development openings in the flat-lying, bedded ore body are driven in rectangular section from 16 to 32 feet wide and from 8. 5 to 16 feet high. The floor level normally coincides with the bottom limit of the chalcocite values and the roof or back is carried at the most competent stratum adjacent to the top of the ore. A development front consisting of five headings and associated crosscuts is advanced at an average rate of about 100 feet per month using conventional drill and blast techniques and equipment. The cyclic nature of the functions involved in driving the openings and the need to include them in production schedules preclude significant improvements in advance rates using current drill, blast, muck, and haul sequencing.

Jan 1, 1969

By B. Hearn, Malcolm J. McPhersqn

Since 1993 Inco Ltd. has been using wet shotcrete at Stobie Mine. Through an evolving process the mine has reduced the thickness and fiber content of its steel fiber reinforced (sfr) shotcrete. However, the shotcrete support is used without any other type of ground support. A process referred to as bolt less shotcrete. Boltless shotcrete is used as the primary ground support method in development headings when the sub level caving front passes through the development drifts which now serve as migrating ore draw points. Although there are references of drifts and tunnels supported by shortcrete only, written accounts are rare. The empirical nature of most underground support designs makes them of limited value in a boltless environment. Considerable amounts of material testing and insitu trials have been completed in order to estimate material property strengths. Material testing, for adhesion, flexural strengths and compressive strength along with observational insitu data have been used to evaluate the performance of shotcrete. Punch tests on panels have also been undertaken in order to gain insights on the limitations of the two dimensional testing common to the laboratory. All these design aspects have been successfully utilized to design an efficient ground support process.

Jan 1, 1998

By G. G. O. O. Uwadiale

Because mineral matter is finely distributed within the matrix, Nigerian Labia coal is nonresponsive to conventional processing techniques. However, testing has shown that Lafia coal can be upgraded by selective oil agglormation. The optimum oil concentration was found to be 10 mL of paraffin per 25 g of coal. A product containing 6% ash and 1.9% sulfur was obtained with a yield of 90%. Optimum results were achieved with a grinding; time of from 2.5 and 3 hrs, which produced a 98% -400 mesh product. However, further grinding wars found to be deleterious to the process. Also, high pH adversely affected concentrate yields. Utilizing kerosene as a bridging liquid, a product containing 4% ash and 0.33% sulfur was achieved at a natural pH of 6.3 with a yield of 85%. At pH 9 and 10, a product containing 4.5% ash was achieved with a yield of >90%.

Jan 1, 1997

By S. H. Amini, A. Noble

"While many classic mineral processing circuit design approaches rely on deterministic methods, numerous studies have shown that the uncertainty inherent to the various input parameters can significantly influence design decisions. In a separation circuit, many of the technical uncertainties can be mathematically represented as statistical distributions for the unit recovery and feed grade. These uncertainties are then propagated by the circuit design and reflected in the circuit performance indicators, such as product grade and global recovery. An effective design strategy entails identifying which units are most influential in this uncertainty propagation and thus merit further consideration. Unfortunately, few analytical tools provide this information, particularly those that rely solely on deterministic methods. However, one promising approach is through the application of Taguchi’s method, in a fashion similar to statistical tolerancing of manufactured products. Since the technical challenges are fundamentally similar, Taguchi’s method shows promise in evaluating the level of compounded uncertainty by various circuit designs while determining the role of each separation stage on the overall selectivity and variability of separation circuits. This paper demonstrates Taguchi’s method in several two and three unit circuit designs. INTRODUCTION AND METHOD DESCRIPTION Uncertainty in Mineral Processing Circuit Design Analysis and estimation of uncertainty propagation in mineral processing separation circuits is an essential and significant aspect of an optimal circuit design procedure. Most of the current circuit design methodologies have been constructed, implemented, and validated using a deterministic modeling approach. However, numerous studies have shown that the inherent uncertainty induced by uncertain design parameters can be a significant factor in many mineral processing applications (e.g. Kraslawski 1989; Xiao and Vien 2003; Ghaffari et al. 2012). Thus, an effective circuit design methodology must be able to quantify uncertainty in separation circuits with limited input data. This paper describes Taguchi’s method, as an experimental design technique, to approximate uncertainty in mineral processing applications, and then the similarity of results between Taguchi’s method and Monte Carlo Simulation is investigated. Monte Carlo Simulation The term “Monte Carlo Simulation” has been used to refer to a variety of stochastic simulation methods spanning a broad array of business, technical, and scientific domains. Most generally, Monte Carlo Simulation refers to numerical analysis which is based on (pseudo) random sampling from a domain of input parameters with a known statistical distribution. This methodology allows one to determine the distribution of expected results from a standard deterministic process model. Execution of a Monte Carlo simulation begins by identifying the domain of input parameters, statistical distributions for each parameter, and a deterministic process model which describes the anticipated result as a function of the inputs. Next, one discrete random sample is taken for each input parameter, considering the input distributions. The process model is then solved for this mix of input parameters, producing a single simulation result. This procedure is then repeated a “large” number of times, utilizing new randomly sampled input parameters for each simulation iteration. Once the desired number of iterations is reached, the results of each independent simulation are aggregated to produce descriptive statistics such as minimum, maximum, mean, standard deviation, kurtosis, and other distribution parameters. Given enough iterations, the Monte Carlo method fully explores potential synergies between the input parameters. The methodology does not require the process model to be differentiable, continuous, or even analytically defined. While the number of required iterations varies in relation to th"

Jan 1, 2017

By S. C. Gambrell

The practice of roof bolting in world-wide mining activities is very well known, having been used extensively for over 50 years. The theory of roof bolting, however, is extremely complicated, and only recently have investigators attempted to explain the theory. Discussions of the mechanics involved usually originate with actual experience gained in the use of roof bolts. The use of roof bolting in mines has grown dramatically during the period after World War II. Other scientific techniques have emerged in mining research in the same period. Included in these techniques is the study of rock mechanics, which has now developed into an impressive research effort. Various means of analysis have been utilized in rock mechanics, most of which reduce rock to an isotropic, homogeneous medium that could be easily analyzed. Because of the inherent nature of rock, direct research methods frequently are not certain to provide satisfactory results, whereas indirect methods must idealize a given set of conditions in order to be useful. One of the indirect research methods is the photoelastic technique, the theory of which is widely known and needs no elaboration. This technique has been used extensively to examine the state of stress around underground openings of various shapes in certain types of strata, and under various configurations (1,2,3,4,5). Much of the theory used now in the design of underground openings is based on photoelastic evaluation. Unless rigorously applied, how- ever, the photoelastic technique is at best a qualitative means of evaluation. It is useful, however, for unusual examples which can be more or less easily analyzed to obtain the overall state of stress which could not be determined by other means. Little work has been done in the photoelastic analysis of rock bolting, mainly because direct research techniques have proven useful.

Jan 1, 1970

By W. G. Kegel

The past few years have produced a rash of demands for changes in America's way of doing things. From manufacturer to consumer, the rules have been changed. Some of the demands were needed and, when legislated to requirements, did manage some good for all concerned. Other demands made by zealots and taken up by regulatory agencies have forced costly changes on our way of life. No segment of industry has been exposed to more new rules and regulations than the deep coal mining industry in the United States. There is no denying that some good has evolved, but most mining men know that the effects have done more to reduce production and increase cost than they were ever intended to do.

Jan 1, 1975

By Tench C. Page

The Yarnell gold deposit is located 1.5 km (1 mile) south of Yarnell in the Weaver mining district of Yavapai County, AZ (Fig. 1). Elevations within the area of the deposit range from 1400 to 1560 m (4590 to 5120 ft). Historic production, principally from quartz veins localized along the Yarnell fault, was about 200 kt (220,000 st) at an average grade of 6.9 to 10.3 g/t (0.2 to 0.3 oz/st) gold. The federal government closed the mine in 1942 due to World War II. Norgold Resources Inc. leased the property in 1988 and joint ventured it with Asarco the same year. Asarco drilled 7822 m (25,660 ft) in 96 reverse circulation and four diamond drill holes and identified 8 t (260,000 oz) gold reserve. IBema Gold Inc. acquired Norgold in 1991 and is currently permitting the project. Production is expected to begin within the next two years

Jan 1, 1994

By L. Frost

The establishment of a retreat system of longwall mining in this submarine coal field has been under study for many years and, although in special cases there have been limited local applications of this method of mining, there has been no attempt until quite recently to establish a uniform system of longwall retreating' in this area (see Figure 1) Where retreat walls were established, it was strictly as a result of physical, rather than economic conditions. In some of the operations- of the Corporation, where the coal was liable to spontaneous combustion- or the area subject to bumps as in No. 2 Mine of the Cumberland Railway and Coal Company in Springhills the recovery of the coal on a retreat system was more or less a necessity.

Jan 1, 1964

By R. Q. Honaker

Although the fundamental principles of hindered bed classifiers are relatively simple, the complex interactions between the operating parameters provide a degree of difficulty that hinders the ability for on-line plant optimization. Thus, a dynamic population balance model was developed for determining the hindered settling velocity of particles varying in density and size. The steady-state results of the model include partition curves as a function of particle size and density as well as metallurgical data. A statistical comparison of the simulation results with those obtained from industrial units reveal a high level of accuracy with adjusted R2 values greater than 0.95.

Jan 1, 1999

By E. Simoncini, S. Ulgiati, M. Raugei

A thorough Life Cycle Assessment of ilmenite extraction and processing to yield synthetic rutile (ERMS/EARS process) was performed according to ISO norms 14040 and updates. The LCA includes the parallel application of selected environmental impact assessment methods (as illustrated in Ulgiati et al. [1]), the results of which are compared and discussed. A MI value of 16 kg of abiotic matter per kg of output rutile is calculated, which is largely determined by the amount of coal that is employed in the reduction and drying steps. The Gross Energy Requirement is 3.33 * 107 MJ/kg, a value which we found to be comparable with that previously calculated by us for alumina. Finally, the increase factor of the Emergy Intensity from the raw mineral to the marketed rutile is 3.19, which points to a comparatively medium-low requirement for indirect environmental support. GWP and AP were calculated to be respectively 1.89 kg(CO2eq)/kg and 12.7 g(SO2eq)/kg.

Jan 1, 2005

By M. Ge

An in-seam seismic (ISS) - based void detection technique was developed and tested for three general mine conditions: anthracite, bituminous and trona. The detection range tested varied from 80 to 100 m for trona mines and from 33 to 66 m for bituminous mines. High quality and broad band signals were obtained from these tests and were used to delineate the voids. The paper provides a brief discussion on the technical features of this void detection technique, the testing conditions and results, and a general assessment of the technique.

Jan 1, 2010

By T. S. Govindan

The first step in energy conservation in any industrial plant involves implementing many of the day-to-day "house-keeping" items such as fixing leaks, turning out lights, etc. When this step has been accomplished, additional opportunities require a detailed analysis of the energy systems, evaluation of risks and costs, and assignment of priorities. Examples of conservation schemes drawn from the experience of Du pont's Energy Management Consulting Services will be discussed. The utility of heat balances for kilns and dryers in minerals processing will be outlined. The inputs needed from Management for the success of conservation efforts will be discussed.

Jan 1, 1978

By K. Bommer

"In today’s fast paced, globally competitive world, reliability and machine uptime are of the utmost importance to a successful operation. Along with these targets more emphasis is being put on cost reduction and decreasing the environmental impact of the overall system. To achieve these goals, automatic lubricators and lubrication systems have been used in many industries for years and are now becoming a common tool in the mining industry as well. These systems are able to provide proper lubrication amounts at the proper times autonomously, helping to keep the machines running smoothly while minimizing the physical support required. Copper mines in the southwest have begun utilizing auto lubrication to avoid mistakes and unplanned shut downs on critical equipment. This paper will describe the engineering behind these products and review case histories where these tools have been successfully utilized in the mining industry. INTRODUCTION In a dirty, dusty, wet environment, where machine uptime is imperative, proper lubrication of bearings is the only way to survive. With the push for cost reduction and mines striving to lower their environmental impact and increase worker safety according to MSHA guidelines, the old lubrication practices have come under scrutiny. It is no longer encouraged or cost effective to have maintenance personal walk from point to point throughout the entire mine pumping a shot of grease into each bearing once per shift. Automatic lubricators and lubrication systems have been on the market for years but recently, as the market focus has changed and technology has developed, the use of these systems has begun to rise in the mining industry. Automatic lubrication systems can be as simple as a single point lubricator lubricating one (1) bearing or a more complex system lubricating numerous bearings. With system design, the automatic lubricator provides a preset amount of grease to a bearing at regular, predetermined intervals. The systems designed to lubricate more than one bearing contain grease pumps of some design and possibly PLC controllers which can be used to remotely adjust and monitor the system. As can be seen in Figure 1, the right amount of lubrication can be achieved much more reliably with automatic lubricators. Depending on application needs, the market now has products to fit almost every wish list. This paper is going to focus on the single point systems that are lower in cost but are still being effective."

Jan 1, 2015

By Jon G. Baird

The suggestion of the use of measurements of trace amounts of mercury as a guide to the location of sulphide ore deposits was first made by the Russian, A. A. Saukov (1), in 1.946. Since Saukov's original work, many others have studied the abundance of mercury in nature, the association of mercury with sulphide ores and methods by which very minute quantities of mercury can be measured in air, soil, rocks, water and plants. The most abundant of the mercury bearing minerals is cinnabar (HgS). Native mercury is also found and sulpha-salts like tetrahedrite are known to have important mercury contents. Other sulphides of base metal deposits contain mercury in amounts ranging from parts per million to as much as one or two percent. (2) The mercury content of sulphide-free rocks is usually less than a few hundred parts per billion (1 ppb - 10-9), Soils usually contain less than 100 ppb and water less than 1 ppb of mercury. Normal atmospheric air contains about 1 nanogram (10-9 grams) of Hg per cubic metre although higher concentrations can be observed near mercury occurrences, sulphide deposits and volcanic exhalations.

Jan 1, 1973