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By P. H. Lu, M. J. DeMarco, J. R. Koehler

Beginning in late 1985 and continuing through 1987, US Bureau of Mines personnel investigated longwall chain pillar and entry design in two and three-entry gateroad systems at a deep underground coal mine in central Utah. To evaluate their respective stability characteristics, four chain pillars and two longwall panels within the two separate entry systems were instrumented with Bureau hydraulic borehole pressure cells to continuously monitor vertical and horizontal pillar and panel stresses through adjacent panel retreat. In addition, supplemental entry closure information was obtained from sites located in the vicinity of the instrumented pillars. The findings presented in this report demonstrate the practicality of evaluating mine pillar stability using in situ methods and support three significant conclusions regarding ground control aspects of the investigated two-entry system: (1) Total areal entry loading was considerably less for the two-entry system, as opposed to the three-entry system, (2) a smaller percentage of opened ground in the two-entry system contributed to a marked reduction in rooffalls, and (3) lower areal loading in the two-entry system suggests improved ground conditions when mining the underlying seam.

Jan 1, 1987

By R. S. Olsen

Intensely altered rock has had a large effect on slope stability as slopes increase in height at Island Copper mine. Traditionally used, fracture-controlled rock slope stability design techniques have been supplemented with analyses of failure modes which include failure along curved surfaces typical of soil materials. A rock mass rating (Pa) system has been developed that addresses the lower end of rock mass strengths. Rock masses are rated in terms of fabric integrity, alteration, fracturing intensity and gouge strength. Rock mass classes have been defined using statistical frequency distributions of rock mass rating values. Strength values have been assigned to the various classes from shear testing, back analyses of failures and engineering judgment.

Jan 1, 1987

By X. Wu, E. Topuz

Booster fans have been extensively used in underground mines throughout the world. In the United Kingdom, S. Africa, Germany, and other countries, many non-coal and coal mines employ booster fans. These fans help to reduce air leakage, to provide an improved flow distribution and to reduce the overall cost of mine ventilation. Although booster fans have not been used in coal mines in the United States, a recent survey indicated that many coal mines would install booster fans if they were allowed to do so. With the inclusion of coal mines, the use of booster fans in underground mines is expected to increase substantially in order to maintain air quality and quantity requirements. Several techniques are used, with varying degrees of success, in determination of booster fan locations in an underground ventilation network. These techniques include linear programming, critical-path crashing, CPM together with cutset operations, the out-of-kilter method, and simulation techniques. This paper provides a brief comparison of these techniques, and reports on ongoing research for an improved methodology for the determination of booster fan locations.

Jan 1, 1987

By D. R. Forshey

Although the United States has the world's largest coal reserves, most of those reserves will require underground mining. Unless underground injury and fatality rates are reduced, the number of disabling injuries and accidental deaths will grow as coal production increases. The Bureau's Mining Research Program, designed to reduce the exposure of workers to hazards while improving productivity, is described. Major program thrusts are 1) re- moving the hazards or removing miners from the areas in which hazards exist, 2) developing ways to improve management decisions and mining skills so as to reduce the incidence of situations requiring human involvement in hazardous, emergency, or corrective tasks, and 3) improving, or at least maintaining, productivity so that both the tangible (dollar) and intangible (human) costs are reduced. In these ways, projected demands for coal can be met using resources without an increase in human suffering.

Jan 1, 1987

By J. Allen Hennessy

INTRODUCTION The increasing price of gold and silver over the past 10 years has caused increasing activity in the exploration, evaluation and development of precious metals projects. Many of these deposits are relatively low grade and short lived which makes them economically sensitive to the major project variables such as metal price, capital cost, operating cost and of course, tonnage and grade. These deposits often have a high variability (a coefficient of variation -- o/u -- of 2 to 5) and an intermixing of ore and waste, making ac- curate ore reserve estimation difficult, especially when using traditional estimation techniques. Often, the estimation errors are not the 10-20% that we see in base metal deposits, but the errors in grade and tonnage can be as high as 50%, which can obviously have a very substantial impact on project Recent developments in geostatistics when coupled with a sound geologic under- standing of the deposit can significantly improve reserve estimates of these variable precious metals deposits. Examples will be given of actual deposits where the traditional estimates were significantly different from the mining reality. An attempt will be made to explain the applicable geostatistical tools with a minimum of geostatistical jargon and without complicated formulas. The key to improving the reserve estimates is a clear understanding of the deposit geology and especially the structural controls, the use of local estimation techniques that recognize the intermixing of ore and waste and the effect of selective mining on recoverable reserves, and the use of a global reserve estimation technique to check the results of local estimation. Geologic Controls The vast majority of gold and silver deposits being mined or developed in the western United States would be classified as epithermal deposits. These are formed by hydrothermal fluids (hot water) flowing upward through the rocks and depositing the precious metals as the fluids cool or reach areas of lower pressure or favorable chemical sites. These environments are similar to present day hot springs-type environments with some deposits being clearly associated with surface hot springs such as the Homestake McLaughlin deposit in northern California. These hydrothermal deposits can occur in sedimentary rocks resulting in deposits such as those of the Carlin district or the Jerritt Canyon deposit or the hydrothermal systems can occur in volcanic rocks forming deposits such as Round Mountain, Manhattan or Paradise Peak. The mechanisms of formation were similar in either case with the solution flow controlled by geologic structure and precious metals deposited when the conditions of temperataure, pressure and chemical environment were favorable. The gold and silver hearing solutions travel through permeable faults and fracture systems depositing the metals along these channels. The grades are often highest in the central parts of these conduits (veins) with lower grades in the surrounding rocks. An understanding of the nature, distribution and orientation of these plumbing systems is important to a realistic reserve estimate. The point to be emphasized here is that, good, sound geology in the form of geologic mapping, logging of rotary chip samples and a few carefully logged core holes can signifi- cantly improve an ore reserve estimate.

Jan 1, 1987

By Ted D. Maki, Joseph B. Taylor

INTRODUCTION FMC's Paradise Peak mine, located 13 Ian (8 miles) south of Gabbs, Nevada, became opera¬tional in April of 1986 (Figure 1). It was designed and built by Davy McKee, who was instrumental in equipment design and selection. It is the 7th largest gold deposit in the United States with defined reserves of 10.9 mt (12 million st) containing 34 million gms (1.1 million tr oz) of recoverable gold and 933 million gms (30 million tr oz) of recoverable silver. The mine extracts ore by the open-pit method, taking advantage of a 1.5:1 stripping ratio. The mill operates at a 3600 mt/day (4000 st/d) capacity. Crushing is done in three stages to achieve an ore size of minus 0.635 cm (0.25 inches). Grinding further reduces the ore to 85% minus 100 mesh. The ground ore is treated with cyanide in agitated leach tanks and then washed in the counter-current decantation (OCD) thickeners. Zinc is added to the clarified, deaerated pregnant solution to precipitate the precious metals. The precipitate is acid digested to eliminate excess zinc, filtered and retorted to drive off contained mercury. The retorted precipitate is then fluxed, melted and poured through cascade molds. Dore bars are cleaned for shipment and the slag is sent to an in-house slag treatment system. ELECTROSTATIC SEPARATION The dry electrostatic slag treatment system at Paradise Peak is the first installation of its kind. Electrostatic separation has been widely used in mineral processing since the early 1950's. A brief discussion of the theory behind the process is helpful to those not familiar to electrostatic separation. Charging and sepa¬rating slags, dry minerals or other materials by ion bombardment is the most common form of electrostatic separation. Millions of tonnes of minerals are processed each year by this method. In an ion bombardment separation, granular material is fed onto a grounded metal cylinder (or roll) and charged by a corona-producing electrode placed above the roll's surface (Figure 2). While both conductors and non¬conductors become charged, only the conductor is able to lose its charge. The charged non¬conductor, as it rests on the roll's surface, "sees" an oppositely charged image of itself in the metal surface. It is attracted to the image charge, becomes electrostatically pinned to, and moves with the roll's surface. The conductive particle also sees an image and is attracted to it. But upon touching the roll's surface, it discharges rapidly to the grounded surface and is thrown free from the roll's surface with a projectile motion. In the case of precious metal slags, the con¬ductive particles would be metallic prills of dore metal; and the nonconductive particles would be the slag, free of metal. Middling grains, generally, are in the form of a metallic prill encased in or incompletely liberated from slag. LABORATORY TESTING A number of precious metal slag samples have been tested in the laboratory. It has been found that the composition of slags vary widely from one refinery operation to another. Typical laboratory procedure is to crush and size the slag followed by two-stage lab-scale electro¬static separation. The conductor fractions from the first and second pass are then combined as a prill concentrate; middlings fraction and clean slag tailings from the second pass are held separate. Selected results from laboratory testing are shown in Table 1. It is important to note in Table 1 that the assays represent overall silver and gold, not metallic values. From experience, the laboratory results are typically lower in grade and recovery than the industrial installations. This is largely due to the hydroscopic nature of precious metal slags coupled with the high local humidity in Carpco's Jacksonville, Florida, location. Relative humidity in Jacksonville can range form 60-95%, while most mining locations in the

Jan 1, 1987

By B. Arvidson

Computer-controlled machine sorting may, if properly applied, have a dramatic effect on overall economics of a mine operation. Radiometric and photometric sorters have been used for many years and new developments will be reported, such as a sorter for fine particles and a conductivity/magnetic sorter. Application examples with emphasis on precious metals will be given. Economics for some of the projects will be reviewed and discussed. Sorting underground is shown to be particularly attractive, but machine sorting would be profitable in most cases based on savings in transportation and milling/beneficiation costs only.

Jan 1, 1987

By R. E. Wardwell, J. D. Nelson

The design of cover systems for reclamation schemes in tailings requires estimation or determination of tailings materials properties such as compressibility, shear strength, and consolidation characteristics. It is very difficult and expensive to obtain truly undisturbed tailings samples for laboratory testing. In situ testing methods have been developed for economically obtaining large amounts of information. Two of these methods are described, the piezocone and the flat plate dilatometer. Both of these methods provide a rapid means of detailed material identification in a nearly continuous fashion for the entire depth investigated. The equipment is described and typical data from a uranium tailings impoundment is presented. The results are compared and the advantages and disadvantages of these methods are discussed.

Jan 1, 1987

By John T. Bartlett

During the last decade, the increasing price of gold has renewed interest in the various metallurgical techniques for recovering precious metals. Much of the recent technology advance has been aimed at low grade deposits resulting in more complex processes and the necessity of tighter economic control. These factors along with more stringent environmental constraints have made it necessary to optimize each process to minimize energy and raw material consumption and reduce waste discharge. In most cases, detailed process analysis is required to detect inefficiencies and provide alternative solutions for process improvement. Also during this same period, the computer has become widely accepted and computerized mathematical modeling has been applied to an increasingly broader range of problems. Microcomputer based process modeling and simulation systems are routinely used to analyze and improve communition, hydrometallurgical, and pyrometallurgical processes for the recovery of precious metals.

Jan 1, 1987

By D. Farnell

Large semi-autogenous (SAG) or autogenous grinding mills have become the work-horse of the modern high volume concentrator. They typify the economy of scale - fewer, larger machines doing more work, more economically. With fewer machines, equipment reliability becomes of paramount importance. Realizing this, companies installing new or modernizing existing facilities have placed extraordinary emphasis on purchasing a reliable mill. This, in turn, has had a significant impact on mill design. Kennecott And Codelco - Two Different Approaches Last year, two of the world's largest copper producers, Codelco and Kennecott, undertook the modernization of their main grinding facilities at Chuquicamata and the Utah Copper Division (UCD) respectively. Both selected semi-autogenous milling and both, recognized the importance of the reliability of the primary mills. Each, however, decided on a different approach to optimizing reliability. Codelco placed the emphasis on technical innovation in design and opted for a gearless driven mill. Kennecott chose the more conservative approach of active participation and exceptional quality assurance in mill design and manufacture.

Jan 1, 1987

By Carl E. Brechtel, Edward D. Thimons

Providing adequate airflow to effectively dilute and remove diesel pollutants in large mine working headings is a little researched area. A simulated diesel mucking operation was carried out in the Exxon Colony Pilot Mine in Colorado to establish expected levels of diesel pollutants in dead-ended working headings. This heading was nominally 15.3 m (50-ft) wide by 9.2 m (30-ft) high. Two face ventilation systems were tested using sulfur hexafluoride tracer gas to represent the diesel pollutants. The results of these tests showed that conventional face ventilation systems operating at high flow rates could adequately ventilate diesel equipment in the 447-522 kw (600-700 hp) range, even in a dead heading 91.4 m (300 ft) in length.

Jan 1, 1987

By J. V. Byrnark, R. W. von Bitter

The Hibbing Taconite Co. concentrator flowsheet, which is noted for its simplicity, uses single- stage autogenous grinding coupled with two stages of magnetic separation and two stages of size classification to produce a magnetite concentrate. This paper discusses the pilot-plant development of single-stage grinding, the commercial mill grinding performance, and mill liner development. Process innovation discussion includes process control computer implementation, size classification improvements, mill speed changes, magnetic separator improvements, and quality control innovations.

Jan 1, 1987

By John H. Duke

The success of most attempts to evaluate gold orebodies depends largely on the reliability and availability of representative sampling data and the accuracy of geological interpretation. An accurate prediction of the continuity of geology is paramount to a reliable prediction of gold reserves. Many of the problems encountered in gold mine evaluation and planning revolve around the difficulties in interpreting complex geology from sparse drilling data and the eagerness of mine financiers and promoters to start mining in order to derive an early cash flow. Following a program of sampling, geophysics, mapping and geological interpretation, a conceptual orebody model evolves in the minds of those working on the project. It is this model which is used typically in ore reserve computations and mine planning. Computer modelling techniques to represent and evaluate the conceptual model are now available. The main advantages of using computer modelling methods lie in being able to rapidly evaluate, modify and quantify the conceptual orebody model. Ore zones or domains can be recognised and defined both from geology and sampling results. Once these domains are defined as a series of boundaries, they may be modelled as two dimensional grid surfaces or as wire-framed three dimensional geometries, depending on the complexity of the orebody. It is the definition of these domains which provides the sound geological basis for reserves computation. To estimate reserves it is necessary to determine both volume of ore and grade of ore. The ore volume is calculated within each domain as defined by the confining surfaces or wire-frames. The ore

Jan 1, 1987

By J. C. Ransdell

A conventional paddle-type mixer and a gravity-flow mixing device were evaluated for conditioning of fine phosphate feed (- 35 + 150 mesh) with fatty acid and fuel oil, and for acid scrubbing prior to cationic flotation. Gravity-flow mixing re- quired less residence time but slightly higher reagent concentrations to achieve recovery levels similar to conventional conditioning. Grade was less sensitive to reagent concentration using the gravity-flow mixer. For acid scrubbing, the gravity-flow mixer proved equally effective. In this paper, the effect of time, pulp density, reagent concentration, and emulsification on conditioning and acid scrubbing efficiency of the two devices is discussed.

Jan 1, 1987

By Nicholas S. Lynn

Introduction Borne out of the need for reductions in capital expenditure and increased production, several new and innovative carbon systems have been developed for small precious metal heap leach operations throughout the United States. Perhaps the most important aspect of these systems is the carbon column. Though diverse in design, these columns can be categorized into three basic types. The first is known as the gravity flow type or Cortez type. As the name indicates, solution flows via gravity, throughout 4 to 6 columns which are arranged in a normally stepped con- figuration which is conventionally upflow in design. The second type, known as the vertical column, is an arrangement whereby flow is established through pressurization in an upward manner, throughout 4 to 6 columns in a vertically stacked configuration. This arrangement is typically intended for use with small portable heap leach operations and employs a semi-continuous mode thus preventing gravity from occurring. The third type and newest development in carbon columns is an arrangement whereby the solution flows throughout 4 to 6 columns placed in a side by side configuration and pressurized by a single feed pump (Figure 1). A unique aspect to the third type, is that the adsorb and desorb cycles are accomplished in the same column, due to a fixed carbon charge. An in-depth analysis is warranted due to numerous advantages of this arrangement, and is therefore the subject of this paper. Advantages and Variations One advantage to the third type of carbon column is the fact that initial capital expenditures and operating man- power requirements are reduced. Another advantage is that the system can be made fully portable up to a 0.0315 m3/s (500 gpm) capacity. Systems of this nature are operational at several sites. Above this flow rate, the system's portability is one which can be maintained, but disadvantages (i.e. larger diameter carbon column and greater weights) far outweigh the advantages. There are several variations in each of these systems which allow for unique operation. For example, carbon columns made highly portable, can be moved from unitized heap leach operations to central stripping facilities. Although this method has been used in Colorado, transportation of these units has been known to be quite cumbersome. If distances exceed 5 km (3.2 miles) then special front-end loaders are required. This limits heap locations to within 5 km (3.2 miles) of the central processing facility. Another example transports the carbon charge from the heap leach locations to a centralized processing facility; this method is quite effective. For the past 5 years one mine has employed this technique of transporting carbon in 907 kg (1 ton lots), in DOT containers, towed by company pick-ups. This operation uses a vertical column with a 907 kg (1 ton) carbon charge. A third example uses static carbon columns with fixed carbon charges. Carbon Adsorption System The carbon adsorption system (Figure 1) is fed by a single preg solution feed pump. This feed pump pushes the preg solution through all four to six fixed charged carbon columns, which are on line. The sequencing of the carbon columns is achieved using valves. Expansion of the carbon charge is dependent upon the carbon selected. The carbon typically used is a 6 x 12 mesh size, with a nominal flow rate of 0.00135 - 0.00138 m 3/s (22-25 gpm) per square foot of carbon, based on a 40% bed expansion. With 8 x 16 mesh size,

Jan 1, 1987

By J. B. Gwiazda

S. Budirsky J.B. Gwiazda's article deals with an interesting problem that has not been studied thoroughly up to now. Gwiazda has proposed a technical solution that eliminates horizontal load imposed by the supports on the roof. That load is due to a discrepancy between the trajectories of the canopy and the overlying strata. It is true that the improved lemniscate guidance system proposed by Gwiazda could eliminate the horizontal load but, on the other hand, it would represent a higher cost of the supports, and it would complicate the already complicated mechanism of powered supports. That is why we must be convinced that the proposed improvement is actually needed. From that point of view, the following questions should be elucidated. As shown in Fig. 1, the point 0 travels along the curve a, which produces horizontal load on the roof in the direction of the coal face. As stated by Gwiazda, "such a high load is capable of destroying the roof above the support, causing rock debris to be scattered around the face." The results of our measurements carried out in coal mines of the Czechoslovak part of the Upper Silesian Coal Basin show that the convergence in the modern type of powered supports does not exceed several millimeters (not more than 10) per hour. Similar conclusions were drawn from field measurements performed in the Polish part of the Upper Silesian Coal Basin (see S. Romanowicz and H. Szopka in Proceedings, Scientific and Technical Symposium Simmex'85, Katowice, Poland, pp. 73-83). The small vertical closure of the supports also induces a small horizontal displacement of the canopy. In addition to this, the transmission of forces between the roof and the canopy is influenced by the layer of debris on it, i.e., the canopy slides without imposing a greater horizontal load on the roof. This was deduced from field measurements of Voest-Alpine F 4/4500 powered supports equipped with rams instead of the common front lemniscate links (denoted as 1 in Fig. 5). In these supports, the pressurizing of the space of piston rod (1a in Fig. 5) causes the rams to close, which induces horizontal load on the roof in a direction toward the coal face. After the setting of a support unit, the pressure in the piston rod space was found to drop, frequently to zero, because the small horizontal sliding between the canopy and the debris or inside the debris caused the horizontal reaction from the roof to fall to zero, although the debris were pressurized by vertical setting load. As a consequence, there was no horizontal load imposed by the canopy on the roof. A rise of the horizontal load was recorded only after the passage of the shearer and during lowering of the adjacent units. Thorough information on the results of the aforementioned field measurement is given in the paper, "Analysis of the performance of shield powered supports installed in a thick seam," published in Mechanizacja i automatyzacja gornictwa, 1982, No. 12, pp. 38-46. An English translation of the paper is available from the author. We have deduced from our observations that horizontal load imposed by the canopy on the roof in the direction toward the coal face is useful for roof control because it limits the displacement of the immediate roof toward the gob particularly during the rise working. We came to the conclusion that horizontal load on the roof toward the coal face should be induced on purpose. Could Mr. Gwiazda prove with the results of field measurements his contrary opinion? Has the small horizontal displacement of the canopy toward the face actually had an adverse effect on the roof? As shown in Fig. 2, during the setting of a support unit, the canopy imposes a horizontal load on the roof in the direction of the gob. In this case, I share Gwiazda's opinion that the horizontal load could have an adverse effect on the roof. Nevertheless, two important circumstances were neglected - the horizontal compliance of the supports and the interaction of adjacent units. The question of the compliance was analyzed precisely by I. Krumnacker in Gluckauf-Forschungshefte, 1984, No. 5, pp. 219-223. He found that due to the clearance between the hinge pins and the eyes of the sheild and due to the elasticity of the steel structure,

Jan 1, 1987

By Joyce Lappin-Randolph

In 1977, Vertical Crater Retreat (VCR) Mining was developed at the Homestake Mine and by 1983 had become a fully integrated mining method contributing 60% to total underground production. VCR has proven successful for continuous ore blocks exceeding 55,100 tons (50,000 st) with dips greater than 60" and provides a productivity average of 39 tons (35 st) per man shift (tpms) which is 92% greater than Open Cut-and-Fill. Discontinuous ore bodies, small pods, or flat structures not amenable to bulk systems are mined with Mechanized Cut-and-Fill (MCF) and have been able to achieve a 32 tpms (29 st) average productivity. Small, low grade, remnant ore bodies were not economically viable with large scale bulk techniques and couldn't support the development required for a MCF system. As a result, Homestake began researching modified bulk mining methods to increase production rates and maintain safety, while lowering costs. A modified version of VCR with significantly diminished development was prototyped in an effort to reduce initial capital investment and expedite the production phase. This modified method is called Small Block Vertical Crater Retreat (S. B. VCR). Geology The Homestake Gold Mine is located in the northern Black Hills of South Dakota at Lead. Having been active since 1876, it now extends from the surface to the 2,440 m (8000 ft) level underground.

Jan 1, 1987

By V. Rajaram

A mine fire or roof collapse may necessitate the reversal of main fans to isolate the fire or provide an escapeway. Hence, to assist the mine operator in estimating the flow and pressure during reversal, Engineers International Inc. conducted a laboratory and field study of the reverse performance characteristics of main mine fans. Laboratory tests were performed using a 2.4-m (96-in.) diam vane axial fan to establish the forward and reverse performance characteristics under controlled conditions. Subse¬quently, a field test was conducted in a Colorado oil shale mine during which two 2-m (79-in.) diam fans were reversed. The data obtained indicate that reverse performance characteristics depend on the blade angle and the hub-to-tip ratio. The quantity in the reverse mode ranged from 30% to 65% less than in the forward mode, and the static pressure was directly proportional to the square of the quantity.

Jan 1, 1987

By Edward D. Graf

Once the heavy work of mining, supporting and completing an underground structure is finished, there is frequently the problem of handling unwanted water infiltration into the opening. Methods of controlling the infiltration vary from collecting and channelling the inflow to shutting off the water at the surface of the opening or within the ground formation. Discussed are various methods of controlling ground water inflow from within the structure including collector pipe systems, shotcrete, cement grout, bentonite, and various chemical grouts. Case histories illustrating some of the methods and materials will be presented.

Jan 1, 1987

By R. Jin

The hydrophobicity of six coals ranked from lignite to anthracite is studied through contact-angle measurements, bubble attachment-time measurements, FTIR spectroscopy, and Hallimond-tube flotation. The preliminary study shows that the hydrophobicity of coals can be told by the bubble attachment times of coals (a kinetic criterion) or by the contact angles (an - equilibrium criterion). A general agreement between flotation response of coals and measured bubble attachment times or contact angles is observed in the present study. However, the study also shows that the hydrophobicity of coals can be predicted by FTIR spectroscopy work with proper manipulation of the spectra. Both hydroxyl and carboxyl groups in the coals are proven by FTIR work to be responsible for the change of coal hydrophobicity. A new criterion for the hydrophilicity index of coals from FTIR spectra is introduced through the study.

Jan 1, 1987