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By J. Stachurski, N. A. Abdel-Khalek

The objective of this investigation was to compare the results of countercurrent, column cell flotation vs. conventional (mechanical) cell flotation for the cleaning of a sulfur ore (-0.5 mm, 27.43% S) and its discrete size fractions. The tests were carried out using diesel oil as a collector and alcohol as a frother. Generally, column flotation produced concentrates of higher grade. For example, column cell flotation produced a nonwashed concentrate of 65.3% S with a recovery of 84.1%. However, conventional cell flotation produced a rougher concentrate of lower grade (57.0% S) and higher recovery (92.5%). At a superficial-wash water velocity of 0.265 cm/sec, the column product grade increased to 76.4% S, with a slightly lower recovery (80.3%). However, a concentrate with a similar recovery but a lower grade (74% S) than that of the washed product was obtained after three cleaning stages with the conventional cell. The column gave superior results for the flotation of fine and ultrafine particles.

Jan 1, 1994

By Roger E. Melley

Since startup of the Syncrude Canada Ltd. (SCL) oil sands facility in northern Alberta 15 years ago, numerous improvements have been made in the ore conveying system and equipment. The unique nature of the material conveyed coupled with the wide range of environmental conditions encountered results in a number of unparalleled challenges. Most of the improvements have been brought about through the use of detailed field evaluation. Numerous publications are available that detail the various aspects of conveyor technology, and a significant list of papers are presented each year adding to the knowledge in this field. In the area of power formulae in the mid- 1970s, there were two major sources for use in the design of the SCL conveyors: the Conveyor Equipment Manufactures Association Handbook, 1988, and DIN Specifications 22101: Belt Conveyors for Bulk Materials, 1982. These sources were supplemented by the knowledge of the belt and conveyor equipment vendors and from the general literature. The SCL design was based on the CEMA formula. However, it was accepted that this empirical approach had been developed for use with more conventional materials in more reasonable climates. To accommodate

Jan 1, 1993

By A. Hayatdavoudi, E. Apande

Introduction Troublesome shales have plagued the petroleum industry for more than 50 years. Unstable boreholes have been experienced when drilled into shale formations; sane have been studied almost at the beginning of rotary drilling. A little progress has been made since then, but much is still to be learned. No simple solution exists so far, but good drilling practices combined with good mud practices have been helpful. It is well known that shales in general have a strong affinity for water, and when water is adsorbed, the stability of the shale section of a well is reduced; cohesive strength will be re- bed, properties will be altered and shale will generally expand and/or crumble. The amount of water a shale will adsorb depends on the characteristics of its clay mineral and the ionic concentration of surrounding fluid. Fluid loss and the resulting hydration of an unstable formation can lead to serious downhole problems. The rate of fluid loss depends greatly on the permeability of the deposited cake. In impermeable formations, however, there is no loss of fluid into the formation. As such, shale, because of its very low porosity and almost no permeability, could be termed impermeable. Yet there is strong evidence that shale surfaces are wetted by an invasion of fluid into the formation through microfractures and jointed planes. It is this wetting that causes the shales to swell and slough into the hole, resulting in problems such as increased mud volume and treating costs, bridges and fill-up, stuck pipe, overgauge hole, poor cement jobs and increased cement requirements, and logging and completion difficulties. Drilling operations have been defeated by severe shale problems, resulting in & hole being plugged and abandoned after several weeks of drilling. Shales are we1l known as being difficult and costly to drill. Numerous fishing jobs have plagued drilling operations, and the cost of reaming tight holes ad weighting up the drilling mud to control shale sloughing has been high. The concept of controlling water loss to stop sloughing was introduced in the 1920's. It was known-then that calcium in fresh water re- strains clay £ran excessive swelling (theory of osmosis) . This reinforced the belief that calcium-based muds such as lime, gypsum and calcium chloride should minimize hydration of clays. Lime muds were then introduced for this purpose and others in the 1930-1940 decade1. They were used for shale inhibition but were not effective. In an attempt to improve them, a special calcium-type mud beam popular during the 1950's. Hole enlargement continued and they too were considered ineffective. During & early 1960qs, papers were published claiming that shale swelling could be prevented by using high concentrations of chemical thinner called Chrane Lignosulfonate1. This began the era of highly treated lignosulfote muds. Unfortunately, the hole enlargement problem continued. During & mid-1960’s, oil based muds were used for the purpose of inhibition of shales. Not until 1966 did Mondshine2 suggest that the water phase be saturated with calcium chloride. The cost of using oil based mud was considered by most operators not worth the additional hole stability. The potassium chloride-polymer muds were also used in many fields in the United States and western Canada. The use of these muds has been claimed to prevent clay swelling by inhibition and encapsulation in sane areas of western Canada. They are also expensive and their ueneral application is still experimental. To reduce filtrate loss when sloughing shale becomes a problem during drilling is a reasonable practice. This study will investigate the use of cationic organic additives to mud to form a protective barrier between the shale - which are primarily composed of clays - and water that will keep water from entering the microfractures and will also prevent them from hydrating. In so doing, an understanding of modern clay chemistry and & dispersive properties of cationic additives in water based systems will be our most valuable tools. This study will also present and solve a mathematical model for simulating cation transport phenomena in fractured shales using the Framdlich equilibrium adsorption isotherm. The studies will analyze wellbore failure and stability as affected by cationic organic material.

Jan 1, 1986

By M. J. Laurila

Utilizing data compiled from a drilling program and bulk samples taken from a coal reserve, flowsheet performance for a coal preparation plant can be calculated using a desktop computer. Sets of equations defining material balances involving coal preparation plant unit operations are employed to depict plant performance. Techniques are discussed for interpolating washability data and extrapolating size distributions and other critical data to expand on oftentimes limited sets of quality data to a useable data base from which projections can be made. Simulation of flowsheets can be done for a number of different conditions and configurations while still remaining in a limited working memory capacity of 32,000 bytes. Presented within this paper are several pertinent examples.

Jan 1, 1982

By Roshan B. Bhappu

In situ mining and extraction is one of the Frontier Technologies that is destined to play an increasingly important role in the recovery of metals and minerals in the future. This premise is based on increasing costs of labor and energy, increasing awareness of environmental considerations and decreasing ore grades in the United States. Up to now, in situ mining has been successfully applied to a few favorable situations for the extraction of some evaporites, copper and uranium. However, because of its unique technical, environmental and economic advantages, this versa tile technology will find broader applications for the recovery of above commodities as well as for others which are currently extracted by hydrometallurgical means. This paper traces the development of in situ mining and extraction; stresses its dependence on interdisciplinary science and engineering; discusses its theoretical, practical and economic aspects; and finally presents potential applications of this Frontier Technology.

Jan 1, 1985

By Robert L. Runkel, D. Kirk Nordstrom, James W. Ball

Low-flow reactive-transport processes in Wightman Fork downstream of the Summitville Mine, south-central Colorado were simulated using the OTIS and OTEQ solute-transport models. The simulation was calibrated using data from a synoptic study conducted during October, 1998. Discharge over the 8-km reach from just below the mine site to the Alamosa River confluence ranged from 0.077 to 0.17 m3/s; travel time was 10.7 hours; pH ranged from 4.6 to 5.7. Simulations revealed that mass balance for SO4 was maintained, and that pH, Fe, and Al were non-conservative. Simulations allowing hydrous ferric and aluminum oxide or hydroxysulfate precipitation matched observed water quality conditions more closely than simulations without precipitation.

Jan 1, 2000

By Paul K. Edwards

The forces of nature have concentrated sand at various geographical locations that are being mined and processed for a large variety of industrial uses. While most operations serve a number of industries, this paper deals only with the foundry requirements. The forces that have decomposed a source rock, transported this material and redeposited it, have created two general types of sand. These types are classified as the (1) Lake and Bank Sands and (2) Silica Sands. The Lake and Bank sands will contain 92 to 94% Sio2 while silica sands are in excess of 98.5% Si02 and most of the producers wash these sands to up grade their products to 99.5% or better. Included in the Lake and Bank sand would be the river sands that are used primarily for construction purposes. The Foundry Industry is broken down into three types of castings : 1. Grey Iron Foundries The grey iron castings largest market is the automotive industry in Ohio, Northern Kentucky, Indiana, Illinois and Michigan. These castings are poured at temperatures ranging from 2400° to 2700°F. The Silica Sands or the Lake and Bank Sands will satisfy the requirements of the grey iron castings. There are 2 ½ to three million tons per year shipped from the state of Michigan far grey iron castings by the Lake and Bank Sand Producers.

Jan 1, 1974

By Dirk van Zyl

It can systematically assess if the control and management systems designed for protection of sensitive ecosystems are adequate. The results of the assessment can help to identify key elements in the project that pose higher risks due to uncertainties in design. These assessments can then be used to make recommendations to further manage and reduce risks as the project is developed. QUALITATIVE RISK ASSESSMENT The general principles of a risk assessment are well summarized by Kaplan and Garrick (1981): "In analyzing risk we are attempting to envision how the future will turn out if we undertake a certain course of action (or inaction)." Fundamentally, therefore a risk analysis consists of an answer to the following three questions: What can happen? (i.e. what can go wrong?) How likely is it that that will happen? If it does happen what are the consequences?" Risk assessments can be performed using various levels of sophistication depending on the amount and accuracy of data available. A qualitative risk assessment uses descriptors throughout to express the level of risk. The definitions of the descriptors are presented in terms of the magnitude or concentration of contaminants. The results of a qualitative risk assessment can be used to: a)Document the evaluation and decisionmaking process. Although differences of opinion may exist about the definitions of the descriptors, the process is well-documented; b)Express the relative risk associated with specific activities or design elements; and, c)Select those failure modes or release mechanisms that may result in significant risks. These mechanisms can then be further developed in quantitative risk assessments. A qualitative risk assessment can be performed in a systematic approach using four steps (Van Zyl, 1987): A qualitative environmental risk assessment can be used to describe the risks associated with the development of a new mine and to make an evaluation of the adequacy of its design for environmental protection. The relative magnitude of risk associated with selected failure modes and exposure/consequence scenario assessments is the final result of a qualitative risk assessment. These failure modes are based on site specific areas and components, and mining and construction phases. The four steps of a risk assessment (hazard assessment, exposure assessment, consequence assessment, and risk characterization) are used in the development of a qualitative risk assessment. Relative magnitudes (using a five point scale of negligible to high) for each step, can be quantified in a qualitative risk assessment. Development of the descriptors for the relative magnitudes in each step is a very important part of a qualitative risk assessment. The success of a qualitative risk assessment depends on the adequacy of the information base and the subjective judgements made in each step. The quality of the risk assessment is, therefore, dependent on the experience of the assessors. A mining project in a sensitive coastal environment is presented as an example to demonstrate the application of a qualitative risk assessment. INTRODUCTION Environmental impacts from mining operations have been a significant factor in the last century. This was especially the case prior to the development of a better understanding of environmental processes and the potential impacts of acid rock drainage (ARD), and associated heavy metals, or the uncontrolled disposal of tailings and waste rock (Smith, 1987). Much care has been exercised over the last two decades in designing and operating mines to limit the environmental impacts as far as possible. All new mines are specifically designed with mitigative methods to manage such potential environmental impacts. Attention is directed to ARD impacts, safe disposal of mine waste rock and tailings, transportation and storage of chemicals and fuels, and final reclamation. A qualitative risk assessment can be used to address all project systems and activities that may potentially pose a risk to the environment.

Jan 1, 1992

By C. Alan Tapp

Introduction Mine sampling determines the practicality of any mining operation. Improper sampling can result in an incorrect appraisal of present production and future potential. Therefore, the mine department in charge of ore-reserve calculations and mine sampling should be overseen by competent and experienced professionals with technical backgrounds qualifying them to produce accurate results. Sampling is a process by which portions of an ore body are collected and analyzed to estimate the average mineral content of the entire ore body. It is incorrect to assume that a large number of samples eliminates any errors in the sampling method. To obtain unbiased samples, proper sample location with respect to rock type and mineralization is just as important. The sampling procedure must yield correct results for the type of mineral deposit, and careful consideration should be given to whether or not the sampling technique has been developed to an extent sufficient to eliminate as much human error and bias as possible. Only after the ore has been mined and milled is the sampling accuracy known. Sampling also provides information about the bulk composition of the ore for mineralogical and metallurgical tests that determine the economic ore-waste boundaries and the geologic trends for exploration. Actual mining plans can be developed from this information to maximize profits. Accurate sampling is critical, and thus must be approached in a scientific manner. Sampling Techniques Sampling practices and techniques are as varied as the mines in which they are used. The method(s) chosen must be tailored to suit the company and mining needs. For instance, tabular uranium deposits, vein gold deposits, and porphyry copper deposits pose special problems in conducting unbiased sampling. The mine geologist or engineer in charge must develop a sampling method, test it in a sample area, and then critically evaluate the results. If the results from the test area are accurate within the economic limits established by the company, they then may be adopted for general use in the mine. Four routine sampling methods are suitable for specific sampling objectives in the daily mine routine: (1) channel sampling, (2) chip sampling, (3) grab sampling, and (4) bulk sampling. The final sampling results depend upon how the four methods are combined to accurately determine the grade of the ore body. When used in conjunction with each other during different stages of mine development, the channel, chip, grab, and bulk sampling methods provide an in-house check or a comparison by which mining methods and sampling procedures can be evaluated. However, the most valid check is based upon the daily mill production. Channel Sampling Channel samples consist of cuttings collected from a groove cut into the rock about 102 mm (4 in.) wide and 19 mm (0.75 in.) deep. Various tools ranging from a 1.8-kg (4-lb) hammer and moil to a pneumatic chisel can be used to cut the sample. Accessibility and rock hardness determine the applicable sampling tools. Before attempting to take a sample, the rock surface must be cleaned thoroughly; the method of cleaning depends upon the amount of mine dust accumulated on the surface or the degree of alteration of the rock surface. Typical cleaning methods 'employ a wire brush, water, or chipping a fresh surface. Next, the sample outline is marked on the prepared surface, taking care to choose appropriate sample loca- tions. After determining the proper sample outline and length, the sample can be chiseled out, catching the rock fragments on a canvas tarpaulin on the floor, in a powder box, in a canvas bag to avoid contamination, or by some other suitable means to capture all of the sample. Vertical veins present a special problem because the drift's back usually is arched, not square. For example, Fig. 1 shows that if sample intervals are measured from A' to D', the length of the sample interval is greater than the true vein width, resulting in an incorrect calculation of the ore reserves. The correct method would collect three samples at A'-B', B'-C', and C'-D', using either actual measurements in the mine represented by A-B, B-C, and C-D or trigonometric calculations to determine the true ore thickness. Then, each sample can be weighted by grade and true thickness for the vein. Channel lengths usually are a maximum of 1.5 m (5 ft), and it is good practice to divide longer samples into smaller intervals according to structures, changes in rock types, or differences in rock hardnesses. The influence of those features on the mineralization then can be determined. Chip Sampling Chip sampling is a variation of channel sampling used when the rock is too hard to channel sample economically or when little variation in the mineral content indicates that this sampling method will yield results similar to those of channel sampling. Rather than cutting a channel in the rock, small chips are flaked off at regular intervals over the entire face or area being the mine. sampled. Care must be taken to assure that the sample is representative of variations in the rock hardness and type. hi^ method is fast and useful in preliminary evaluations, but it should not be used for quantitative ore-reserve calculations. Grab Sampling Grab sampling is a fast method for double checking either channel or chip-sampling procedures, and, in some instances, mine production can be estimated from carefully taken grab samples. Grab sampling takes equal amounts of material at selected intervals over a mine dump, a muck pile, or from an ore car to estimate its Channel Sampling mineral content. Generally, this method is not considered reliable. Many independent variables can affect this type of sampling process. Thus, if the ore occurs in the softer fraction and a proportional amount of the result-

Jan 1, 1982

By A. F. MacDonald

The rugged foothills of the Allegheny Mountains in Washington County, Southwestern Pennsylvania, lend their terrain very well to the use of aerial tram systems to carry refuse from Bethlehem Mines' two preparation plants to disposal areas some distance from the plant site. Bethlehem Nines Corporation has three deep mines in this area, all of which are mining the Pittsburgh Coal Seam, The raw coal is prepared for metallurgical use in a preparation plant at each mine, Trio of these preparation plants, Mine No, 51, Ellsworth and Mine No. 58 Marianna, use aerial tram systems to dispose of reject material.

Jan 1, 1966

Exploration for SX-EW copper deposits and plant operation depend largely on the copper and gangue mineralogy inasmuch as? they govern the feasibility of leaching. Mineralogical parameters are pivotal in identifying if an ore is suitable to dump, heap, thin layer, vat or agitated leaching and/or if the ore requires curing, bioleaching or alternate lixiviants. Problems in metallurgical testing, production leaching and recovery forecasting are very often related to lack of mineralogical data. This presentation will illustrate the use of quantitative mineralogical analysis for exploration, acquisition assessment and optimization of copper leach operations.

Jan 1, 1996

By Norman K. Olson

Hard-rock phosphate was formerly the chief source of all phosphate in Florida. Its mining began as early as 1883 or 1884, reached a peak level in 1907 and ceased in 1966. Production began and ended in the Dunnellon area (Marion County). Nearly all the hard-rock was exported because its P205 value per unit weight is generally more concentrated than that of other phosphate rock. Maximum production, achieved in 1907, was greater than 600,000 tons; the total value for that year was more than $4,000,000. The Florida hard-rock phosphate, considered to be of middle Miocene to early Pliocene age, has an intermittent narrow, linear northwest-southeast trend. The Ocala uplift controls the phosphate distribution for more than 200 miles. Other known, isolated hard-rock areas appear to be related to minor folds on the Ocala uplift. The hard-rock phosphate deposits are a heterogeneous mass of sedimentary fluorapatite, chert and quartz with textures ranging in size from clay to boulder, and in thickness from 0 to 40 feet. Phosphate compounds, principally fertilizers, have been in oversupply since before 1968, but by the middle 1970's some experts forecast a supply-demand balance. A brighter future for exploitation of Florida hard-rock phosphate depends upon one or more of these factors: 1) depletion of present premium grades (74 percent BPL and above) of land-pebble; 2) cheaper electric power from nuclear generators for profitable electric furnace operation; 3) supply and price of sulfur; 4) manageable rail-water transportation costs; 5) development of less costly mining and separation methods; and 6) minimum competition from high-grade foreign phosphate.

Jan 1, 1972

By N. A. Abdel-Khalek

Laboratory flotation tests were carried out on six Polish samples of different rank. Each coal sample was screened into 7 size fractions. The floatability of size -by- size for various size ranges for different samples was investigated as: well as the original samples. The results showed that in flotation of size -by -size, the intermediate size had the maximum floatability. The range of this intermediate sizes varied with the rank of the coal. Entrainment was shown to be an important mechanism to the recovery of fine particles in flotation of the original samples".

Jan 1, 1989

By G. B. Misra

Heat transfer in partially wet mine airways has been the subject of study for quite some time. Of the available models some are quite empirical while some involve highly complicated computation. Models developed by Amano, Mizuta and Hiramatsu and Banerjee and Chernyak as well as that modified by Misra were evaluated for prediction of the rise in temperature and humidity (mixing ratio ) of partially wet mine airways at four different levels in the Mosaboni Mine of Hindustan Copper Ltd. Measurements of barometric pressure, virgin-rock temperature and changes in dry- and wet-bulb temperatures along a certain length of airway were made. The rise in dry-bulb temperature and mixing ratio of air over the sections were calculated using the three models. The study reveals that calculated values of temperature and moisture content agree within possible rage of experimental error with all the three models in case of three levels with low wetness factor, but the agreement was poor for the relatively more wet airway on the 30th level. Errors in the estimation of fw to the tune of 0.025 can lead to fair error up to 0.94 K in the estimated temperature while alternative ways of estimating the heat transfer coefficient [] can introduce errors up to 1.45 K.

Jan 1, 1987

By John P. Dunford, Lewis A. Martin

Researchers from the Spokane Research Laboratory, National Institute for Occupational Safety and Health, in cooperation with engineers from the Department of Mining and Geological Engineering, Montana Tech, Butte, MT, conducted tests to evaluate the tensile strength, fiber count, and adhesion of shotcrete applied to panels at the Chief Joseph Mine, a research and training facility operated by Montana Tech. The tests were conducted in the mine ramp 45 m (150 ft) inward from the access portal at a 15.25-m- (50-ft) long test area divided into five 3-m- (10-ft) long sections. The shotcrete was applied with an Aliva 240.5 machine fitted with a predampener. Different amounts of fiber were entrained for each panel. Background data for the tests were obtained from earlier tests in Nevada, where the typical amount of fiber in the shotcrete is 6.54 kg/m3 (11 lb/yd3). All shotcrete was applied to a minimum thickness of 76 mm (3 in). Cylindrical cores were drilled at the wall-rock interface, and adhesion was tested after 28 days of curing. The goal of this research is to reduce the need to install multiple supports at the advancing face by creating a safer mining system that limits the time miners are exposed to unprotected roof.

Jan 1, 2005

By D. S. Young

The random variable transformation technique was introduced to make probabilistic structural analysis simple for such structures as slopes, dams, foundations, or retaining walls, where two shear strength parameters are required to analyze their stability by deterministic engineering methods. The Gaussian transformation was performed by using the Hermite orthogonal polynomial function. This transforms the experimental histogram into the standard normal distribution, whose mean is zero with a variance of one. Then, the transformed standard Gaussian variable was used for the bivariate normal probability analysis on the structure. The Gaussian transformation technique and the Hermite polynomial functions were chosen because the bivariate joint normal distribution is well-known and easily available for engineers. The binormal probability analysis was made on the probability region of a certain confidence level by superposing the critical line on the same probability region that divides the region into two zones: stable and unstable zones. The critical line was determined from the back calculation of the deterministic engineering method of the structural design. Finally, the probability of structural failure at a desired confidence level was calculated by simply integrating the volume of the binormal density surface above the unstable zone in the probability region. A simple case of slope analysis is introduced to illustrate this probabilistic approach, and its effectiveness and simplicity is demonstrated.

Jan 1, 1987

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

5.1 TECHNOLOGY FOR INJECTING GROUT INTO BOREHOLES The grouting of fractured rock aquifers is carried out in separate stopes' whose size is determined by the hydro¬stratigraphy and the hydrogeologic properties of the rocks. The number of grouting stopes required depends on the number of intervals in the hydrogeological section. Each grouting stope can include one or several aquifers. Com¬bining several water-producing strata into one stope is pos¬sible only when the values of their permeability coefficients are very similar. In order to achieve efficiency, as many aquifers as possible are included in each stope. In practice, the length of a grouting stope in a borehole usually ranges within 10 to 40 m, but occasionally up to 100 m or more. When conducting grouting operations through boreholes drilled from the face of underground workings, the stope's size can be determined by the thick¬ness of each fractured zone (fault) or by the length of the hole if the entire rock is fractured. Injecting grout into separate intervals of fractured rock can be done from the bottom to the top with ascending stopes or from the top to the bottom with descending stopes. Injection from bottom to top is always preferred. Injection also can be accomplished in separate fractured permeable horizons in the section. The STG integrated grouting method is based on the use of packer and injection equipment and technological schemes that facilitate injection in highly isolated fractured aquifers. For this purpose STG has developed a standard series of packers. 5.1.1 INJECTING GROUT THROUGH BOREHOLES DRILLED FROM THE SURFACE Grout can be injected according to one of four techno¬logical schemes depending on the depth of the fractured aquifers, their hydrogeologic properties and the availability of grout injection equipment if the area is remote (Fig. 43). In the first and most preferred technological scheme (1), the grout is injected from the bottom up with ascending stopes. When this scheme is employed, grouting holes are drilled to the entire projected depth based on the field investigations injection tubes is installed above the top of the first hydrostratigraphic unit to be grouted. This scheme al¬lows the grout to be injected only into the permeable stra¬tum or strata isolated by the packer near the bottom of the hole; it precludes the need to clean or redrill the hole before injecting grout into the next overlying stope. This ascending stope scheme is preferred because of its simplicity and ef¬ficiency. It is used in most of the grouting of saturated fractured rock for the purpose of isolating aquifers while excavating deep mine shafts in the USSR. The second technological scheme (II in Fig. 43) injects the grout in sequence from top to bottom with descending stopes. Injection under this scheme is accomplished through the hermetically-sealed mouth of the hole. It is most advis¬able to grout with descending stopes in saturated fractured rock wherein the aquifers constitute thick networks of frac¬tures. Under these conditions it is practically impossible to drill a grout hole to the entire depth of the underground workings due to the large production rate of ground water. In conducting isolation operations with descending stopes, additional work (relative to ascending stopes) is necessary to drill out the grout in order to create and inject the next stope. As a result, the total amount of drilling required to drill out grout can even exceed the volume of drilling of original rock. Consequently it is advisable to use this inefficient scheme only for injecting grout into holes that contain prolific frac¬tured aquifers at depths less than 150 to 200 m. The third and fourth technological schemes (III and IV in Fig. 43) are used in the selective grouting of separate aquifers. For example, these methods would be applied when a grout hole is drilled to its entire depth, wherein grouting of the upper aquifers must be done first so that the sinking of the first shaft can begin prior to deeper grouting. The third technological scheme is used for the selective injection of grout into an aquifer with a thickness of up to 5 m. For this work the DAU-1-2 removable dual packer is installed. It is lowered into the borehole on an injection pipeline. For aquifers with thicknesses of more then 5 m, the selective injection of the grout is carried out according to the fourth technological scheme; a combination of the DAU-2 hydrodynamic plug and the DAU-1 packing device is used. In order to implement this injection scheme, the DAU-2 hydro-dynamic plug is installed at the bottom of the fractured aquifer and the DAU-1 packer is installed at the top of the aquifer.

Jan 1, 1993

By D. Puckett

The feasibility of maintaining equipment location, personnel location and vital statistics in surface and underground mining operations has increased with advances in radio frequency identification (RFID) technology. The paper addresses the use of this technology, relatively new in the mining industry, to track surface equipment in mine pits, loading points, and processing facilities. Specific applications are discussed including a simplified and a complex truck tracking system and an automatic pit ticket system. The paper concludes with a discussion of the future possibilities of RFID technology in mining, such as monitoring biofeedback of personnel (heart and respiration rate, body temperature, exertion level, etc.) repetitious movements to study work habits, and logging air quality via personnel sensors.

Jan 1, 1996

By D. R. Hembree

This paper discusses the mineralogy and metallurgy of the Kuroko-style volcanogenic massive sulfide ores mined at Formosa Exploration Inc. I s Silver Butte Mine located in southwest Oregon. The principle ore type is pyrite + chalcopyrite + sphalerite, while subsidiary pyrite + bornite + sphalerite ore is also mined. The two styles of mineralization have distinctly different metallurgical behaviors and must be treated separately. Selective flotation methods are used to produce copper and zinc concentrates from the first ore type, while bulk concentrates are produced from the bornite ore. Careful mine planning and coordination between the mine and mill are utilized to achieve optimal recoveries.

Jan 1, 1993

By Elizabeth Anne Semon

The marketability of a large volume bulk material is significantly effected by transportation costs. Hence, it is essential for anyone that deals with minerals or mineral wastes to be able to reliably estimate bulk transport costs. The three types of bulk overland transportation modes that will be discussed are truck, railroad, and pipeline. Contract and private carriage are the two types of truck transport generally suitable for movement of bulk commodities. Contract carriers are those transportation firms that service a limited number of shippers and geographic points, and carry a restricted number of commodities. Private carriage consists of control over drivers and equipment by the shipper, ownership or lease of the transport equipment, and shipment of commodities related to the primary business of the shipper.1 A combination of private and contract carriage may also be used. Depending on the size and regularity of the shipment, rail offers single car, multiple car, or unit train service to move bulk minerals. Both long and short distance pipeline transportation has been shown economically and technically feasible for many different solids such as, mine and quarry products, coal, chemicals, copper concentrates, phosphate rock, and sewage sludge. 2

Jan 1, 1973