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At the beginning of the seventies pre- drainage of methane was introduced in West German coal mining in order to reduce the gas emission from the worked seam. After three pilot tests on the "Rheinland" Colliery (86 boreholes in seam "Prasident"), "Bergmannsgluck/ Westerholt" Colliery (23 boreholes in seam "Gustav") and "General Blumenthal" Colliery (27 boreholes in seam "Hugo" and 170 boreholes in seam "Karl") systematic large scale tests with extended investigations were carried out un the "Rheinland" Colliery (127 boreholes in the seams "Anna", "Blucher", Prasident"). During this time an analogous test was carried out on the "Consolidation" Colliery (21 bore- holes in seam "B"). In these studies and tests the technical feasibility, efficiency and economy could be proved, so that pre-drainage of methane starts spreading. On the "Rheinland" Colliery pre-drainage has become common practice in gassy seams. Mainly two points impede a faster spreading of method: 1. in many cases the necessary requirement, that one or both of the gate roads should ire driven 6 to 12 months before the face is worked, is missing. This is mainly due to the difficulty of retreat mining in passe seams.

Jan 1, 1982

By Charles E. Larson

Square Butte has no history of mineral development or production. Surface mapping, radiometric surveys and chemical analyses of rock samples reveal no mineral potential. The area has a speculative potential for petroleum and natural gas. INTRODUCTION The Square Butte Instant Wilderness Study Area, in Chouteau County, Montana, consists of 1,946 acres (788 ha) about 35 miles (56 km) southeast of Fort Benton and 50 miles (80 km) northwest of Lewistown (fig. 1). This study was performed under provisions of the Federal Land Policy and Management Act (Public Law 94-579, October 21, 1976), in conjunction with the U.S. Geological Survey, which is responsible for regional geological, geochemical, and geophysical investigations, and the U.S. Bureau of Mines, which determines the mineral potential of existing mines, prospects, or mineralized areas.

Jan 1, 1982

By R. W. Boyle

"EDITOR'S NOTEPart 1 of this paper, published in the August issue, covered lithochemical surveys based on major, minor and trace elements in whole-rock samples from primary enveloping halos.Part 2, which follows, continues with discussions of surveys based on leakage halos, zonation of elements, stable and radioactive isotopic distributions, and geothermometric, geobarometric, thermoluminiscent, pedochemical, hydro-chemical, atmochemical and biogeochemical methods. It concludes with a selected bibliography.Surveys Utilizing Leakage HalosLeakage halos can be used as scalar or vector quantities in thesearch for deeply buried blind mineral deposits. In the first case, the occurrence of a leakage halo (e.g. a small pitchblende seam) may simply indicate the possible presence of pitchblende mineralization or orebodies; in the second case, the areal or three-dimensional distribution of leakage halos may, by their increasing frequency and I or increasing metallic or gangue element content, provide a means of vectoring in on loci of mineralization and orebodies. Examples of both cases are given in the following text.Leakage halos are of many types, as shown in Figure 2 (Part 1), and are best discussed with respect to their associated mineral deposits. Furthermore, it should be noted that leakage halos are frequently offshoots or extensions of enveloping halos. Where this occurs, careful tracing of the leakage halo(s) downward or upward respectively will often lead to the apex or nadir (bottom) of the enveloping halo(s) and ultimately to mineralization or ore."

Jan 1, 1982

By C. W. Morris

The continuous miner training system provides classroom materials training aids, and a continuous miner simulator to provide a short, intensive instruction and skills development program for training or retraining continuous miner operators. This project included refinement of training objectives, revision of training aids, improvements to the simulator, and evaluation of the entire system. This report describes the revised classroom materials and the validation of the training system.

Jan 1, 1982

By Tobias W. Goodman, George N. Aul, Joseph Cervik

The Bureau of Mines conducted research to degasify a virgin block of coal in the Beckley Coalbed using long horizontal holes drilled in an array from the bottom of a ventilation air shaft. Eight horizontal holes and an in situ pressure hole were drilled in a radial pattern to depths ranging from 418 to 910 ft (127 to 277 m) and 205 ft (62 m), respectively. The investigation showed the shaft was in a basin saturated with water, which resulted in low gas flows from holes. The in situ pressure at 200 ft (61 m) into the coalbed was 79 psig (544.6 KN/m2). The aver- age gas and water flows for the eight holes were 94,100 ft3/day (2,665 m3/day) and 13 gal/min (49.3 l/min) . After 514 days of degasification, 18.7 million ft3 (0.53 million m3) of gas had been drained. Methane face emissions, which were measured as a section advanced toward the shaft, were reduced about 77 pct, emphasizing the value of drainage by this technique. All holes were grouted before the area around the shaft was mined.

Jan 1, 1982

By D. A. Fenton

Two large 34 ½ foot diameter mills have seen seven years of continuous operation before massive fracturing occurred on one of the trunnion cones. These fractures were discovered when liner changes were made. Four modes of failure were identified, three of the modes combining to cause a large circumferential tear in the 4-inch cone plate. A theoretical account is made of the modes of fracture and comparison made with field results. New material was used to repair the fractured cone, and the complete repair took ten weeks. The repair was monitored and lasted twenty-six weeks before the newly designed structure was available.

Jan 1, 1982

By W. N. Hamilton

"Evaporites abound in the Western Canada sedimentary basin, recurring throughout the stratigraphic column and widely scattered over the basin region. Alberta, situated centrally over the basin, includes many of the evaporites within its borders, but only a few have the right geological conditions to form potential economic deposits of salt or gypsum. These include the Middle Devonian Elk Point Group and related evaporites, Upper Devonian Wabamun Group-Palliser Formation evaporites, and evaporites of the 'Triassic Charlie Lake and Whitehorse formations.Salt deposits underlie almost half of Alberta. The major deposits belong to the Elk Point Group, which includes thick salt units in the Lotsberg, Cold Lake and Prairie Evaporite formations. Another, younger salt deposit in the Wabamun Group has only local distribution. Elk Point salts have a maximum aggregate thickness of 430 m in east-central Alberta. The deposits deepen southwestward across the province, from minimum depths of210 m at Fort McMurray to 1820 m at Edmonton.The thickest, most extensive salt deposits are the upper Lotsberg and Prairie Evaporite salts. Both have excellent development potential. The upper Lotsberg is up to 150 m thick, consists of clear, very coarse crystalline halite with only minor impurities, and has a notable lack of non-salt strata in its succession. Analytical data for the Lotsberg Salt show extraordinary chemical purity. Prairie Evaporite Salt differs lithologically from the Lotsberg in having a significant non-salt component. The deposit is up to 200 m thick and consists of medium crystalline halite, brownish coloured and cloudy as a result of finely dispersed impurities. The main impurities-shale, anhydrite and dolomite-are most prevalent as interbeds in the salt succession, ranging from very thin laminae to beds several metres thick. Although of lesser purity, the Prairie Evaporite Salt is shallower, thicker and more extensive than the Lotsberg and the salt quality is quite acceptable for most industrial uses."

Jan 1, 1982

By M. Leroux, J. A. Finch

"Sizing below 38 µm (400 mesh) is an increasing requirement in mill testing laboratories. The Cyclosizer is gaining wide acceptance in this role. A recent alternative is to use electro-formed sieves (micro-sieves), which are available with openings down to 5 µm. The advantage of micro-sieving is a cut size which is independent of mineral density, easier retention of the fi nest fraction and reduced purchase cost. A simple method of microsieving using an ultrasonic bath, which performs readily to 10 µm, is described.A systematic study of Cyclosizing and micro-sieving is reported for three samples; silica, copper oxide and a lead concentrate. Micro-sieving gave comparable results. Comparison with a standard sizing method, the Sedigraph 5000D, revealed misplaced material in each fraction by both methods. The range for micro-sieving was from 20% to 60% depending on the sieve and material; for the Cyclosizer it was a consistent 40-50% for all cones and all materials. The Cyclosizer can be sensitive to feed mass and particle size for dense materials. It is recommended that for minerals of specific gravity greater than 5, no more than 35 g of -38 µm be used.The disadvantage of micro-sieving is that it requires continuous operator attention and sieving time depends on sample mass. To fractionate 35 g-as used on the Cyclosizer and considered at reasonable mass for mill testing purposes-through 25-, 15- and 10-µm sieves requires about 80 minutes.IntroductionParticle size has a pronounced influence on mineral separation processes. Consequently, the measurement of particle size plays an important part in mill testing. By far the most common method is by screening. This, however, is usually limited to 38 µm (400 mesh) or occasionally 25 µm (500 mesh). Increasingly, mineral products are becoming finer than this limit and consequently finer sizing methods are required."

Jan 1, 1982

By Lawrence L. Brown

As part of the Federal effort to help maintain an adequate supply of fertilizer materials, the Bureau of Mines conducted research to characterize complex low-grade Michigan phosphate materials and to devise methods for recovering fertilizer-grade phosphate concentrates. The rock is from middle Precambrian sedimentary formations in the Marquette Range located in the central part of the Upper Peninsula of Michigan. Occurrence of phosphate-bearing sediments in Precambrian rocks is rare. This deposit, containing 12 to 15 pct P2O5 is believed to be the richest Precambrian phosphate deposit in the United States. Petrographic examination showed that the conglomeritic rock contains crystalline apatite both in "black pebbles" and in quartz matrix. Attrition scrubbing and sizing tests showed that there is no tendency for phosphate to concentrate in any of the sized fractions. In beneficiation tests, which included grinding to minus 325 mesh followed by phosphate mineral flotation using a fatty acid-fuel oil collector, a phosphate concentrate was obtained that had a grade of 30 pet P2O5 and a P2O5 recovery of 72 pct. This concentrate should be suitable for fertilizer manufacture.

Jan 1, 1982

By A. E. Kehew

Earth resistivity surveys using the Wenner electrode configuration were made in unmined areas and reclaimed spoils at three surface lignite mines in western North Dakota. The study sites were the Center, Indian Head, and Falkirk Mines. Automatic interpretation of the curves was accomplished with the method of Zohdy and Bisdorf (1975). Overburden materials become more homogeneous with respect to resistivity during the mining and reclamation process. The spoils curves show localized hetero-geneities probably related to discontinuous moisture content variations in the spoils. Consistent resistivity differences in spoils occur between mines with over-burdens of different mineralogic compositions. Groundwater chemical composition in terms of total-dissolved-solids (TDS) is inversely correlated (r=-0.61) with apparent resistivity at the 80-foot spacing from a survey centered at a piezometer screened near the base of the spoils. Better inverse correlation is probably pre-vented by high variability in TDS within the spoils groundwater. A useful predictive relationship may be the plot of mean apparent resistivity at the 80-foot spacing for ar. entire mine against mean TDS for all wells in spoils at that mine. Data from more mines are needed to test the hypothesis.

Jan 1, 1982

By P. Hanniala, S. Harkki, B. Andersson

"The Outokumpu Flash Smelting Process was put into commercial operation in 1949. The original process was based on the use of preheated air. The adoption of oxygen enrichment of the flash smelting process air in 1971 was a crucial improvement, especially for smelter capacity. Oxygen use according to the Outokumpu method means selection of the most suitable oxygen content for the process air on the basis of the concentrate composition, capacity and matte grade. The furnace is not bound to only one fixed degree of process air oxygen enrichment; the oxygen content may vary from 21% to 100%.The degree of oxygen enrichment and the matte grade are the most important process parameters in reducing the operating costs of the whole system, including smelting, converting and acid production. Raising the matte grade allows more efficient use of the reaction heat in the smelting stage. Sulphur is recovered in a high-strength, constant-flow gas stream. The blowing time in the converters is reduced considerably.Oxygen enrichment and high matte grade improve the energy efficiency of the plant complex as long as the surplus heat from the system is not wasted. When the operating costs and investment expenditure of the total plant complex are taken into account, a reasonably high degree of oxygen enrichment (35-80°/0) is the best alternative in practice. Existing operating smelters have achieved single-stream thoughputs of over one million t/yr of concentrates.HistoryLocal capital and operating costs as well as lower energy consumption per unit of metal produced are now the crucial factors in determining the choice of process. In the late I940s, the same reasons prompted Outokumpu to develop the original flash smelting process in Finland. Between 1900 and 1940, very few technical changes in pyrometallurgical copper production took place. However, the commercial use of the flash smelting technique began a new era in metal extraction from sulphide concentrates. The characteristic feature of the method is that the concentrate is smelted in suspension: this offers the most favourable conditions for heat and mass transfer, and thus high reaction velocities. Outokumpu was the first to exploit the known advantages of flash smelting in a commercial installation at Harjavalta in 1949."

Jan 1, 1982

By A. J. Sinclair, N. Champigny

"The Cinola gold deposit in the northern Queen Charlotte Islands, British Columbia, is in a clastic sequence consisting of a lower shale of the Late Cretaceous Haida Formation and an overlying interbedded sequence of pebble conglomerate and coarse-grained sandstone of the Middle Miocene Skonun Formation. Both formations are intruded by a stock and dykes of porphyritic rhyolite. Two K-Ar model ages indicate mineralization and probably porphyritic-rhyolite intrusion at about 14 Ma (middle Miocene). A splay of the Sandspit fault system constitutes the footwall on the west of the deposit.The Cinola deposit can be classed as Carlin-type based on features such as (1) small particle size for gold, (2) Tertiary age of mineralization, (3) element abundances (e.g. high Hg), (4) dominantly argillic alteration, (5) a structural setting associated with major faults, (6) porosity of the host rock and (7) spatial and possibly genetic association with felsic intrusions. Gold is widespread and occurs mainly as grains of< 0.5 11-m in silicified sedimentary rock, and in quartz veins. Locally, coarse > 100 11-m particles of native gold occur in quartz veins, especially in strongly brecciated porphyritic-rhyolite. Ore minerals are mainly pyrite and marcasite, but include small amounts of chalcopyrite, sphalerite, galena, pyrrhotite, cinnabar, tiemannite (HgSe), rutile, magnetite, hematite and limonite in addition to native gold and electrum. No silver minerals have been found; silver was found in gold particles in amounts varying from 6.2 to 24.2 weight per cent. Alteration products are sericite, illite, kaolinite and chlorite, with abundant quartz of several stages.The host rocks, specifically the Skonun Formation, formed as an alluvial plain facies in a braided river system discharging into a marine basin in early Middle Miocene time. During the Middle Miocene, this sequence was intruded by a rhyolitic stock. The highly porous and permeable Miocene clastic sequence apparently provided an optimum setting for the development of a large geothermal system, the energy for which probably derived from the rhyolite stock. Mineral deposition followed a well-defined paragenesis, which, from oldest to youngest, is: (1) precipitation of iron sulphide minerals, and early quartz; (2) several stqges of quartz veins, with deposition of sphalerite succeeded by galena, chalcopyrite and visible gold. Micron-size gold was precipitated throughout these two stages of mineral deposition. Argillitization of the host rocks probably in part coincided with the mineralization and continued during cooling of the geothermal cell.In-situ reserves have been estimated at 41.1 million metric tons averaging 1.85 g Au/tonne using a cutoff grade of0.86 g Au/ tonne."

Jan 1, 1982

By Loren R. Anderson

This is an Appendix of Volume I. It contains the results of the electric cone penetrometer soundings from the first phase of the testing program.

Jan 1, 1982

By Loren R. Anderson

A probabilistic slope stability computer model was developed at Utah State University and independently evaluated by CH2M Hill Consulting Engineers as a design and analysis tool. Volume I of this report describes the model and Volumes III and IV are users manuals for the computer programs. This volume (Volume II) constitutes the report by the CH2M Hill. It includes a description of a field investigation and stability analysis of an existing tailings dam. Based on the results of the analysis, the tailings dam was evaluated for redesign using probabilistic considerations. Conclusions are made regarding the usefulness of the model as a design tool.

Jan 1, 1982

By G. F. Skala

Additional testing of two prototype instruments for the discrete detec­tion of submicrometer particles in metal and nonmetal underground mines is described. The detectors are designed to utilize properties of submicrometer particles to distinguish between those produced by fires from mine ambient particles primarily resulting from diesel engine exhaust. One instrument is based on the ability to pyrolyze fire produced parti­cles, but not diesel exhaust particles. An alarm is produced when the ratio of pyrolyzed to unpyrolyzed particles exceeds a preset value. The other detector utilizes a continuous flow ion chamber preceded by a denuder to remove charged particles. Discrimination against diesel produced submicrometer particles results because the diesel particles are naturally more highly charged than fire produced particles. In-mine testing was conducted over a period of several months. Both in­struments exhibited an ability to discriminate against diesel exhaust. However, problems with the recording equipment, and with the instruments themselves, due to large ambient dust particles, prevented enough data from being obtained to choose between the two methods. Recommendations to correct the problems are made.

Jan 1, 1982

By Charles E. Semler

The possible utility of selected compositions in the systems SrO-ZrO2-MgO and SrO-ZrO2-Dolomite for commercial refractory products was confirmed in this study. Specifically the compositions of interest are located in the sub-systems SrZrO3-ZrO2-MgO and SrZrO3-ZrO2-Dolomite. Both sub-systems have regions that are not affected by hydration, but the sub-system SrZrO3-ZrO2-MgO has a larger candidate compositional region. Solid state phase relations were determined for the subject systems over the temperature range 1600°C to 1300°C, with compatibility diagrams presented for each. The systems appear to show good refractory character with no apparent low melting compositions. As the compositions of commercial interest lie mostly in the higher ZrO2 portion of each system, information is presented regarding the firing treatment required to develop stabilized cubic ZrO2. A list of suggestions for further study of these systems, to further delineate the characteristics required for possible commercial utilization, is presented.

Jan 1, 1982

By E. H. Kurt

INTRODUCTION The simplest way to drill a hole into rock is to strike a steel chisel or drill bit with a hammer. Early miners used this elementary hand technique so successfully in "single jacking" and "double jacking" that the first mechanical rock-drill designers sought to duplicate it. However, in the early developments, the designers were forced to retreat to a construction known as the "piston drill," wherein the entire drilling element is tied to the piston and reciprocates with it. It took nearly 50 years to devise a method of divorcing the two elements and to achieve the original hammer principle used in hand drilling. The Mont Cenis tunnel, drilled through the French Alps in 1861, usually is considered the birthplace of the mechanical rock drill. The Hoosac tunnel in Massa¬chusetts was drilled at about the same time, and the success of these two ventures paved the way for innova¬tions that produced the Burleigh, Ingersoll, Sergeant, and Waugh piston drills between 1870 and the turn of the century. In 1897, George Leyner of Colorado introduced what has been considered the most significant development in rock-drill history. He devised the hollow drill steel for water flushing; when combined with his free-piston hammer drill, this became the first lightweight and dust¬free underground machine. In addition, Leyner&apos;s drills introduced improvements such as automatic rifle-bar rotation of the drill steel and chuck, automatic lubrica¬tion, and an enclosed throttle control. During the 1920s and 1930s, automatic feeds, centralizers, sliding cones, and the automatic water back head were developed, laying the groundwork for all pneumatic underground drills currently in use. Since these pioneering efforts, rock-drill development has been concerned primarily with refining the designs and improving the metallurgy to make faster, lighter, and more dependable machines. Fig. 1 illustrates the progress in drilling speed during the first 125 years of rock-drill development; that progress is a credit to the persistence and inventiveness of the designers in the rock-drilling industry. ROCK-DRILL CLASSIFICATIONS To meet the variety of conditions encountered in rock drilling, several distinct types of drills have been de¬veloped. In general, rock drills may be classified as either hand-held or mounted, with the hand-held ma¬chines including the jackhammer or sinker, the jackdrill or jackleg, and the stoper. The mounted drills are com¬monly known as "drifters." Table 1 shows that each type of drill is available in several sizes from different manufacturers. Fig. 2 illustrates a typical rock drill, showing the principal components. The jackhammer or sinker, shown in Fig. 3, is used primarily for general mine utility work such as drilling anchor holes (short vertical holes to bolt or anchor machinery), pin holes (short, usually horizontal holes to fasten sheaves, etc. to side walls), popholes (for blasting large boulders), and similar applications. They also are used for shaft sinking. Jackhammers are classified ac¬cording to weight, and they range from 7 to 30 kg (15 to 65 lb). The rock drill originally known as a jackleg was made by clamping a pneumatic cylinder or leg to a jack¬hammer, both to support the weight of the machine and to feed the tool forward in horizontal or uphole work. The more modern jackdrill refined this concept to make the hinged pivot for the leg integral with the drill cylinder and to group all drill and leg controls in the drill back head for convenience. Fig. 4 illustrates a typical jackdrill. Jackdrills are classified according to cylinder bore size, ranging from 60 to 83 mm (2.375 to 3.25 in.). Because of their light weight and versatility, jack drills are very effective in small drifts, small tunnel head¬ings, and stoping. The larger machines are applied to hardrock formations and in applications where drilling speed is a primary consideration. In soft formations or

Jan 1, 1982

By John H. Bassarear, A. A. Dor

INTRODUCTION Primary grinding mills as defined in this paper, are autogenous or semi-autogenous rotating, tumbling mills having a coarse feed with a top size usually varying from 150 to 300 mm (6 to 12 inches). Most frequently, the feed to these mills is the product of primary crushing plants but in some cases run of mine ore is used as feed such as at Benguet, McDermitt, L-Bar Sohio, Bear Creek and others. The size of the largest lumps in the mill feed is principally determined by limits set by the mill feed trunnion diameter and by the type and size of conveying equipment used for handling and of bins and stockpiles used for storage. The use of primary mills in mineral processing plants has expanded very significantly in the last twenty years, and they have been successfully applied to the treatment of a wide variety of ores in both wet and dry grinding applications. In the present state of technological development for primary grinding mills, there is a wide choice of acceptable options with respect to type and size, grinding flowsheet and grinding media. The choice of the optimum combination which would result in the most satisfactory condition of operation, from the standpoint of both metallurgical results and costs, depends on the nature of the ore, the required characteristics of the grinding products and on various geographic and economic circumstances. The effects of such factors and considerations on the choice and sizing of primary grinding mills and flowsheets are reviewed and discussed in the following pages. PRIMARY MILLS SELECTION Inherent Characteristics of Primary Grinding Primary grinding is complex because it involves several comminution mechanisms which occur simultaneously and to some degree interfere with one another. These mechanisms of impact, attrition and abrasion have been described earlier. The relative importance of these three main grinding actions and their mutual balance resulting in a continuous uniform operation of the mill depend to a very large extent on the nature of the ore and its physical properties. Homogenous, single mineral ores will behave quite differently from heterogenous ores consisting of different minerals bound together along defined

Jan 1, 1982

By C. B. Webber, W. A. Grell

"Sherritt Gordon's Ruttan Mine is a copper-zinc operation located approximately 600 miles north-west of Winnipeg. Open pit mining began in 1973 at 10,000 Tpd., with a change over to underground mining in 1981 producing 7,500 Tpd. Ore is crushed to -3/4"" followed by rod and ball mill grinding (two ball mills in series), differential copper-zinc flotation, and dewatering. The copper and zinc concentrates are trucked 75 miles to Lynn Lake, where they are shipped by rail to the HBM&S smelter at Flin Flon.This paper describes a method used to predict the effect of grind particle size on metallurgical performance, as well as details on the copper and zinc metallurgical performance due to fine grinding in actual plant operation.Initial statistical studies on historical data (including ore source, head gr3de, grind, etc.) indicated that copper and zinc recoveries increased with fineness of grind, in the range of 50 to 65% -200 mesh.In order to extrapolate beyond 65% -200 mesh, and to develop expected nonlinear relations between grind, Cu concentrate grade, and recovery, the assays of screen fractions of copper concentrate and tailing were used to determin8 the relation between net smelter return (a function of grade and recovery) and grinding cost. The optimum economic grind would be that grind at which the net smelter return less grinding cost reached a maximum."

Jan 1, 1982

By J. K. Richmond

This is a progress report summarizing more than 5 years of a long-range Bureau of Mines research program. The purpose of this program has been to investigate the fire and explosion properties of Green River oil shale in the Bureau&apos;s Experimental Mine, in laboratories, and in the field. The lean limits of explosibility of oil shale dust, by small-and large-scale tests, are reported as a function of grade, ignition source, and particle size. A limited number of laboratory tests on the autoignition of oil shale dust layers and the spontaneous combustion tendencies of oil shale are described. Moderate-scale rubble fire tests were conducted to determine flame spread rates as a function of ventilation flow. With the aid of a system to continuously monitor methane emissions in a deep oil shale mine, the characteristics of the methane flow are reported as a function of ventilation and blasting procedures, and are compared with gas yields from core samples. Tentative predictions as to the emission of the methane to be expected in deep oil shale mines, far from the outcrop, are offered. Brief reference is made to a parallel investigation into the fire and explosion hazards of oil shale mining and processing by a Bureau contractor; detailed results of the latter investigation are to be found in the contractor&apos;s reports.

Jan 1, 1982