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By Klaus-Peter M. Hanke

INTRODUCTION In early 1979 Jim Walter Resources, Inc. (JWR) of Brookwood, Alabama approached TMCI Construction, Inc (TMCI) to make a proposal on a program that involved the sinking of up to 10 ventilation shafts of approximately 6.7 m (22 ft) diameter and ranging in depth from 500 to 700 m (1650 to 2300 ft) for the JWR coal mines in Alabama. At this time TMCI was already constructing the first spiral underground bunker (capacity 2000 tons) in North America for the JWR organization at their No. 4 mine in Alabama. The TMCI proposal was based on the use of the mos modern large diameter shaft boring machine rather than sinking the shafts using the conventional drill blast-muck technique. The proposal was made based o: the experiences by the parent company, Thyssen Schachtbau, which has been using this type of machin in Germany for shaft boring since 1971. As a result of the TMCI proposal JWR issued a purchase order to TMCI for the construction of four 6.7 m (22 ft) diameter, concrete lined, unfurnished ventilation shafts ranging in depth from 500 to 700 (1650 to 2300 ft). An order was thus placed with WIRTH Machinen- and Bohrgeraete Fabrik GmbH, in Germany for the manufacture of a model 650/850 E/Sch "Schachtbohrmaschine" (Vertical Shaft Borer = V-mole which arrived on site in Alabama in early 1981. The first V-mole GSB 450/500 was introduced in Germany in 1971 and was capable of enlarging in one step a pre-drilled 1.2 m (4 ft) pilot hole to 4.5 - 5.0 m (14.7 - 16.4 ft). This machine has sunk 9 staple shafts and deepened one surface shaft for a total of 2360 m (7740 ft) of shafts. On the last shaft boring operation in 1978 the machine was converted as an experiment to drill without a pilot hol using a hydraulic pumping system to remove the cutting debris. A second generation machine, the SB VI 500/650, was introduced in 1977 for enlarging the pilot hole to a range of 5.0 - 6.5 m (16.4 - 21.3 ft) diameter. This machine is still in operation and has already drilled well over 2000 m (6500 ft) of shaft. The third generation of V-mole, the SB VII 650/85( for diameters from 6.5 to 8.5 m (21.3 to 27.9 ft) was: commissioned in May 1980 and has been used for two surface shaft deepenings totalling 606 m (1990 ft) with another scheduled for 1982. The main advantages favouring the use of such V-moles were identified as: 1) A reduction in manpower to the crew required in a conventional shaft sinking operation. 2) A considerable reduction in time to complete a shaft compared to conventional techniques. 3) The use of the V-mole eliminates many of the hazards encountered in conventional sinking. Based on the successful performance of the first three V-moles in Germany, Thyssen Schachtbau decided to employ this principle abroad. In 1980 a second machine of the third generation was built and is now operated by TMCI Construction, Inc. in Alabama. The first shaft was completed at the end of 1981 and this paper describes the method of operation including some unique aspects not attempted on prior V-mole operations and some of the statistics arising out of the experiences during the first shaft boring operation. THE NO. 7 MINE FAN SHAFT SITE Jim Walter Resources, Inc. was formed in 1970 to exploit the coal field in Alabama on the southern tip of the Appalachian coal field. The coal reserves amount to around 650 million tons of mainly good quality coking coal of which about 350 million tons are to be extracted over the next 30 years. Shaft sinking and preparatory work began in 1972, and at present 6 mines are producing around 5.4 million t.p.a. Annual production is to expand to 10 million t.p.a. as soon as possible, and the ventilation shafts to be sunk by TMCI play a vital role towards attaining this goal. The first shaft site is located at the No.7 mine, near Brookwood, Alabama. The actual location of the shaft relative to the production shafts is shown on the mine plan (Fig. 1), which also shows the room-and-pillar extraction system used at present. The mine plan further shows the conveyor route used for the muck removal. The geological survey showed that the strata consisted of horizontal layers of mainly sandstone, sandy shale and shale interspersed with several coal seams. The seam being extracted at the No. 7 mine is a combined seam made up of the Blue Creek and Mary Lee seams at a depth of 513 m (1682 ft) and having an average seam thickness of about 2 m (6 ft). At the beginning of September 1980 the surface site preparation and pre-grouting work was completed by JWR, and TMCI was able to commence with Stage I of the shaft sinking program - the drilling of the pilot hole.

Jan 1, 1982

By James L. Hendrix

The applicability of a rotating drum with an attached trommel as a flocculation dewatering device of Florida phosphatic claywastes was studied. Several experiments were conducted in order to determine the important fabrication and operating parameters. Results from the experiments indicated flint the dosage of polyethylene oxide required to form a satisfactory product cake could be reduced by using the rotary drum, which introduces less shear stress into the fluid than a constant stirred tank mixer. It was found that the length of the mixer was the most important parameter. Increased length increased the retention time the flocs stayed in the mixer without dewatering. Such increased time decreased the stability of the flocs. Through analysis of residence time distribution data it was found that the longer mixers approached plug flow characteristics. Other important parameters were trommel length and rotational speed. Mixers from 10 cm diameter to 35.6 cm diameter were tested with good results.

Jan 1, 1982

A recent Evening Post Headline read 'Officialdoms Inertia Risking Millions in Growth Strategy'. The article beneath that headline outlined the difficulties encountered by the Government's growth strategy projects and included the paragraph 'public efforts by Government to brush up communication of the widespread nature of its growth strategy is believed to have been matched by efforts in private by some ministers to overcome their frustration at the constant delay while reports and paper work are moved between departments and through the system'. The article went on to point out that these delays could cost New Zealand millions of dollars in lost overseas exchange savings. We as responsible citizens should be appalled or at least concerned that such costly delays should or even could occur. But those on the outside of the system who spend much of their working life prodding, reminding, pleading, cajoling, even begging, the bureaucracy should be excused a smirk or a smile or even a good belly laugh. The Minister is usually a part of the bureaucracy passing judgment on someone else's proposals. This time the Government has plans of its own. Plans which it wishes to see implemented as rapidly as possible and even with the assistance of the precautionary National Development Act with its 'fast track' provisions those plans have got bogged down in the bureaucratic machine which this Government has done as much as any other to create. In the circumstances, the thought of Her Majesty's Ministers frustrated at the constant delay while reports and paper work are moved between departments and through the system is not unpleasing.

Jan 1, 1982

By T. G. Carnahan

The Bureau of Mines conducted research to simulate in situ leaching of a porphyry copper sulfide ore. The purpose was to determine the effects on copper leaching of additions of NaCI to a conventional sulfate leach system and to analyze the oxidant requirements and the oxidation products that resulted from leaching of the sulfide minerals in the ore. Baseline sulfate and sulfate plus 2-pct-NaCl-leach-solution experiments employing 167-kg (367-lb) and 152-kg (333-lb) charges, respectively, were conducted. The minus 25.4-cm (10-inch) ore was leached at 90° C under flooded conditions with 2.8 MPa (400 psig) of oxygen bubbling through the system. Sixty-seven pct of the copper was extracted in 1,236 days of leaching in the NaCI experiment, whereas 47 pct of the copper was extracted in 913 days of leaching in the baseline experiment. Diffusion into and out of the ore fragments controlled the rate of leaching, which precluded enhancement of the leach rate by the chloride ions. The calculated oxygen consumption was 3.17 kg/kg of copper extracted in the baseline experiment and 3.46 kg/kg of copper extracted in the experiment to determine the effect of chloride ions.

Jan 1, 1982

By Masszi D

INTRODUCTION The association of methane with coal has been known almost since coal was first mined and in general it has been considered the curse of underground coal mining. Historically, most of the effort to handle this methane has been directed at diluting it to a safe level and preventing an explosive mixture of methane and oxygen from forming in any part of the mine. This is generally achieved by in-mine activites. This approach has two principal drawbacks, the severity of which are dependent on local conditions. The first is that the activity associated with the handling of the methane is undertaken in the mine and there- fore competes with the mining activity for the limited space available. The second drawback is that the whole methane handling activity is generally a cost to the coal because methane is so diluted that it cannot be used. With the value of energy increasing thi; second drawback is gaining importance so th.u several mines in Europe and North America ;it,, now installing systems to recover the methane in a useful form. Two approaches are

Jan 1, 1982

By Peter J. Szabo

Introduction This is the Pacific Rim. Nearly three fourths of the world's population live on or near its border. In the recent past, on any given day, one could find traversing its„ borders pirates, smugglers, gunrunners, missionaries, slave traders, tourists, and businessmen. If one were not too observant, he would notice that over the last several years the ocean traffic in the Rim has pretty much stayed the same--Polynesian outriggers, passenger ships, general cargo vessels, ore carriers, and oil tankers. If he were to look a little closer, he would notice some coal colliers. A still closer look would reveal that most of them were hauling metallurgical coal destined for the steel mills of Japan, Taiwan, or Korea. If his senses were keen, he would observe that not only were the number of colliers increasing dramatically, but also that a relatively new cargo was being carried to those countries as well as additional areas in the Pacific Rim., This new cargo thermal coal holds the promise of not only exceeding the metallurgical coal trade, but significantly cutting into and decreasing the already vast oil tanker business throughout the Pacific Rim. The high price of crude oil as a result of the oil embargo in the Middle East in 1973-74 has forced the Pacific Rim countries to convert to coal, and as a result, the coal resources of the Rim are being exploited at an ever-increasing rate. Although coal produced outside the Pacific Rim does enter its borders for ultimate consumption from such places as South Africa and the East Coast of the United States, the vast majority of thermal coal that will enter this booming market place will be sourced from countries within the Pacific Rim. Accordingly, this paper will provide an overview of those sources with particular emphasis on the non-communist block nations. This is not to say that countries such as China and Russia will not be players in the Pacific Rim thermal coal arena, but that -most consuming countries in the Rim are basing the majority of their near-term sourcing requirements from countries that provide a stable source of supply. Pacific Rim Coal Resources Figure 1 is a map of the Pacific Rim countries and identification of major coal resource locations within its border. Going in a clockwise direction, the following countries will be the major suppliers of thermal coal through the year 2000: Australia, Indonesia, China, Russia, Alaska (U.S.), Canada, United States (Utah, Colorado, Wyoming, Montana), and Colombia, South America. Other countries such as New Zealand, Malaysia, Philippines, Korea, etc., will play a role in supply, but compared to the ones previously mentioned, will only have a minor impact. When one looks at the vast stretch of ocean that separates the sourcing countries from the consuming countries (Japan, Korea, Taiwan, Hong Kong, Singapore, etc.), one realizes that costs of transportation will play a significant role in determining which countries and in what quantities coal will be purchased. Figure 2 for comparative purposes, shows

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

By James P. Connell

HISTORICAL BACKGROUND The American Concrete Institute defines shotcrete as "mortar or concrete conveyed through a hose and pneumatically projected at high velocity onto a surface." This definition thus includes what is traditionally known as gunite, which is a pneumatically applied mortar. In mining practice, the term shotcrete is restricted to pneumatically applied concrete, and this differentiation will be used in this chapter. In 1914, following the invention of the mortar gun in 1907, then chief engineer of the US Bureau of Mines (USBM) George Rice developed the gunite process for underground test work at the USBM facility at Bruceton, PA. After World War I, gunite was used extensively in American mines and was also utilized for underground civil works such as the San Jacinto tunnel in California. The greatest development was in Europe where, as early as 1911, gunite was successfully used as an overlay for deteriorated tunnel linings. In 1951, the Swiss firm Aliva developed a pneumatic gun capable of handling coarse aggregate, thus making possible the first use of shotcrete at the Maggia hydropower development. Initially, shotcrete was used to reduce manpower requirements for forming and placing conventional concrete. However, by 1954 Sonderegger was reporting that the structural advantages of shotcrete were derived from its flexibility and from the fact that it could be applied almost immediately after the opening had been made. The incorporation of wire mesh into the shotcrete led to the new Austrian tunnel method or NATM. The use of shotcrete in American mines has been implemented more recently. This delay seems to be due to previously unsuccessful experiences with gunite as a structural material and to the US reliance on wood or steel supports in main-line haulageways. The long experience with the apparently more substantial rigid supports led mine operators to be reluctant to accept the new and seemingly unrealistic lighter shotcrete support. APPLICATION REQUIREMENTS Shotcrete is a relatively new material for use in underground support systems. Consequently, experienced miners are not always available who are capable of applying the material effectively. Shotcrete, particularly in the small cross sections typical of mine shafts or haulageways, is applied in cramped quarters under less than ideal conditions. Adequate lighting should be made available. The surface should be clean and free of running or dripping water. It may be necessary to collect flowing water in plastic pipes or water collection devices. Any dry cement dust from previous shotcrete applications should be washed from the surface in order to assure a good bond. The US Bureau of Reclamation (USBR) while shooting test panels at the Cunningham tunnel in 1974, found that experienced shotcrete operators were able to obtain up to three times greater compressive strengths than were obtained by unskilled operators using the same equipment and shotcrete mix. ENVIRONMENTAL AND SAFETY REQUIREMENTS Since sodium and potassium hydroxide, as well as other moderately toxic compounds, are often contained in shotcrete (particularly where accelerators are used), safety precautions must be taken to prevent skin and respiratory irritation. Nozzlemen and helpers are required to wear gloves, protective clothing, and ventilation hoods with a filtered air supply. Respirators approved by USBM, equipped with chemical filters that will not pass the caustic mists, may be permitted in lieu of hoods if goggles or safety glasses are worn. Protective ointments are available to reduce skin irritation. All air and shotcrete feed hoses should be equipped with safety-type couplings and secured with safety chains at each coupling to prevent whipping in the event of a hose or coupling failure. Some environmental effects can take place down-stream from the development face being supported. The accelerator compounds, as well as the portland cement used in the shotcrete, will be found in the rebound material which falls to the invert of the heading. Since these compounds may be leached from the rebound material and carried by the drainage system, it may be necessary to install neutralizing or other water treatment facilities. Investigations may find that the final reaction with other compounds being leached from the mining operations may result in a more or less environmentally acceptable end product. USES OF SHOTCRETE General Uses Shotcrete, as a combination of cement, aggregate, and accelerator, is utilized for underground openings such as shafts, adits, haulageways, and service chambers for the following general purposes : (1) primary sup¬port; (2) final lining; (3) protective covering for excavated surfaces that are altered when exposed to air (the protective covering may be of a temporary or final nature); (4) protective covering for steel or wooden supports, rockbolts and rockbolt plates, heads, nuts, and other mats, including wire fabric, used to prevent rock-falls; and (5) as a lagging material in place of timber, steel, or concrete between steel or wooden supports. These applications can be grouped into three general use categories: shotcrete used as a rock sealant, shotcrete used as a safety measure, and shotcrete used as a structural support. Use as a Rock Sealant Thin applications of shotcrete can reduce or prevent slaking of shales or other rocks that are altered when exposed to the wetting and drying cycles created by mine ventilation circuits. While shotcrete may be effective in preventing such rock alteration, at the present time it is not as economical or efficient as other commercial sealants. However, if the sealant property can be incorporated into the structural support capability, the added contribution is usually helpful. Thin applications are not usually sufficient if the alteration of the

Jan 1, 1982

By Warder A. Shaver

INTRODUCTION Mine No. 95 of the Consolidation Coal Co., a divi¬sion of Continental Oil Co., is located in Shinnston, WV. Mine No. 95 was opened in January 1967 and has been operating continuously since that time. The mine has 544 hourly and salaried employees. In 1974 this mine was the largest coal producer in West Virginia. Opera¬tions include two longwall sections and eight Marietta Borer continuous miners. Originally there were also two Jeffrey Heliminer continuous miners, but use of these was discontinued in June 1978. Production averages 12 727 t/d (14,000 stpd). Two portals now service the mine with a third portal proposed for 1978. GEOLOGY Mining takes place in the Pittsburgh coal seam which dips 0.04 rad (21h°) northwest. The working seam is steam coal with an average height of 2.2 m (86 in.). Immediately above the seam, the overburden is com¬posed of 6.1 m (20 ft) of shales followed by 6.1 m (20 ft) of weak sandstone and 51.8 (170 ft) of shaley limestone. The shale weathers when exposed to mois¬ture. Surface water has been a problem due to a fire¬clay floor which turns to paste when mixed with water. GENERAL DESCRIPTION The present mining plan for Consol's No. 95 mine has a proposed life of 35 to 40 years. One set of main headings has been driven in the north direction and one set in a western direction. These are the areas in which the majority of the virgin coal lies. The panels are driven off of the mains at 2134 to 3048-m (7,000 to 10,000-ft) intervals. The continuous mining sections are set up 133 x 610 m (600 x 2000 ft). Longwall sections are 259 x 1220 m (850 x 7000 ft). All of these mining sections are retreating uphill. Development is with continuous miners. A Marietta Borer miner is used for the main development work. These miners cut 4.1 m (13 ft 6 in.) wide and 2.1 m (7 ft) high, leaving an arch cut. Along with the arch, a 152.4-mm (6-in.) thick layer of roof coal is left in place to protect the roof from weathering. The combi¬nation of the self-supporting arch and the roof coal al¬lows the entries to be driven 30.5 m (100 ft) without bolting. Full bolting, however, is carried outby the last open crosscut. This bolting is done on 1.5 x 1.5-m (5 x 5-ft) centers using 1.5-m (5-ft) long roof bolts with tandem anchors and 152.4-mm (6-in.) steel washer plates. When needed, crib blocks with a 50.4-mm (2-in.) drill hole are used as supplemental supports. Three-hole steel channel with the ends curved down¬ward to conform to the borer arch may also be used. In the mining sections, a 14 BU loader is used be¬hind the Marietta miner to load two Torkar shuttle cars. These haul to Long-Airdox feeder breakers which load onto a 1.07-m (42-in.) wide conveyor belt. With the exception of the shuttle cars, which use 275-V d-c power, the face equipment operates on 550-V a-c power. The main power line is 7200 V ac. The same loading and hauling equipment is used with the Marietta Borer miner for driving both panel and development entries. In the retreat sections, the same equipment with the exception of the miner is utilized. Because of the advantage of a variable cutting hight, Jeffrey Heliminer ripper-type continuous miners were employed in these retreat sections. For longwall mining the first unit uses an Eickoff EDW 300-L double-drum ranging shearer. Drum di¬ameter is 1.69 m (66 in.) with a drum rotation speed of 45 rpm. The drum is equipped with 48 Carmet 34-L-1 bits. The 300-kW (400-hp) motor operates on 1000 V. Tramming of the shearer head is done through a hydraulically driven 26-mm wide chain. Maximum cutting speed is 7.93 m/min (26 fpm), and maximum tram speed is 12.8 m/min (42 fpm) with an average cutting speed of 3.66 m/min (12 fpm). Web cutting depth is 762 mm (30 in.) and the cutting height is 2.13 m (7 ft). The conveyor width is 686 mm (27 in.). It uses a single chain, 26 mm thick, with flights on 0.91-m (3-ft) spacing. The face conveyor speed is 64.3 m/min (211 fpm) when driven by two 112-kW (150-hp) 1000-V Louis Allis motors through torque convertors. The roof support used is Gullick Dobson chocks [6-legged 77.3 t (85 st) per leg, 463.6 t (510 st) total capacity] with double-acting leg jacks. The hydraulic pressure for the legs is 13.8 MPa (2000 psi), and the hydraulic pressure for rams is 6.8 MPa (1000 psi). The yield valves are set at 57.8 MPa (8500 psi). Each chock weighs 4.09 t (4.5 st). The stage loader is an Eickoff drive, with a 68-kW (90-hp), (00-V Reliance motor. A 26-mm thick single strand chain with flights on 0.6-m (2-ft) centers runs at 75.89 m/min (249 fpm). The stage loader is self-advancing and is connected to the belt tail piece with a special fifth wheel attachment so that the stage loader can advance and guide the tail piece. The hydraulic pressure for the roof support is supplied by two 45-kW (60-hp) Reliance motors op¬erating on 699 V, with 1.26 L/s (20 gpm) constant volume Gullick hydraulic pumps. Each pump services alternate chocks, with the two pressure lines connected at the tailgate by 25.4-mm (1-in.) hydraulic hoses, rated at 27.21 MPa (4000 psi). The 6.8 MPa (1000 psi) pressure rams advance is supplied by a 45-kW (60-hp), 600-V Reliance motor and a 1.26-L/s (20-gpm) con¬stant volume Gullick pump. Return lines are 35.5-mm (1.4-in.) diam hydraulic hoses with rated capacity of 27.21 MPa (4000 psi). The fluid used in the hydraulic system consists of 95% water and 5% Hy-chock fire resistant water soluble oil. Water is stored and mixed in a 6096-L (2000-gal) track-mounted tank and goes to a 655-L (180-gal) reservoir at the pump. The belt conveyor structure is 1.07 m (42 in.) wide with 0.47 rad (27°) idlers, bottom stands located on 3.05-m (10-ft) centers, top stands on 1.5-m (5-ft) centers, and is supported on a 19-mm (3/4-in.) diam

Jan 1, 1982

By David H. Carlson, John H. Johnson

The paper presents background information, a description of mine characteristics that affect air monitoring techniques, and a review of Portable measurement techniques and instruments for gas and particulate matter sampling. A general review of the laboratory and field research being conducted at Michigan Technological University in the area of diesel emissions is presented. The facilities involved with this research are also described. These are the experimental mine, the diesel dilution tunnel test cell, the mine air monitoring lab (MAML), portable instruments, and the production mines used.

Jan 1, 1982

As part of the Bureau of Mines program for conserving domestic mineral resources, a survey was made of the methods used for identifying scrap metals. Because of the large number of alloys currently being scrapped, correct identification of these materials is essential if they are to be recycled effectively. The methods and instruments used to identify scrap metals are described and evaluated. These include object recognition, color, density, magnetic testing, spark testing, chemical spot testing, thermoelectric measurements, eddy current measurements, and elemental analysis by chemical and instrumental methods. Other potential techniques are discussed also.

Jan 1, 1982

By Lowell L. Patten

The Federal Bureau of Mines participated with the National Aeronautics and Space Administration and Martin Marietta Aerospace in developing, building, and testing a portable X-ray fluorescence spectrometer for use as an analyzer in mineral-resource investigative work. The prototype battery-powered spectrometer, measuring I1 by 12 by 5 inches and weighing only about 15 pounds, was designed specifically for field use. The spectrometer has two gas-proportional counters and two radioactive sources, 109Cd and 55Fe. Preliminary field and laboratory tests on rock specimens and rock pulps have demonstrated the capability of the spectrometer to detect 33 elements, to date. Characteristics of the system present some limitations, however, and further improvements are recommended.

Jan 1, 1982

By James R. Aggson, Verne E. Hooker

GENERAL CONSIDERATIONS The design of a mine or any underground opening in rock is similar to the design of any other structure, such as a building or a bridge, in one important way. This similarity lies in the fact that the stability of the structure is a function of the loads that will be applied during the structure's expected lifetime. In the case of an underground opening, the applied load is the product of the existing in situ stress field and the concentration of this field that is produced by the extraction of a load¬bearing material. Thus, a knowledge of the in situ stress field is considered a basic requirement in the design process and essential to the eventual stability of the underground structure (Duvall, 1976; Obert, 1972; Merrill and Peterson, 1961; Dahl and von Schonfeldt, 1976; Hedley and Wilson, 1975). The in situ stress field can be used along with the physical properties of the rock masses involved, the ge¬ometry of the various rock types, the geometry of the openings involved, and a failure criterion to evaluate structural stability of a planned mine by theoretical or numerical techniques. This same information can be used to investigate ground control problems in existing mines. In situ rock stress determinations can be separated into three categories: the determination of applied stress, also called the absolute or field stress; the deter¬mination of concentrations of the field stress near open¬ ings; and the determination of changes in rock stress as load-bearing material is removed. The applied stress field is more difficult to determine than are stress changes. Various instruments and techniques have been developed to determine applied rock stress and stress changes in rock. ABSOLUTE OR FIELD STRESSES Theoretical Considerations In the absence of actual measurements, it is often assumed that at any point, the vertical stress is equal to the weight of the overlying material. [ ] where S„ is vertical stress, S is average weight density of overburden rock, and h is depth from the surface. This is normally consistent with actual experience. The horizontal stresses are assumed to be equal and related to the vertical stress through the Poisson effect. [] where Sh is horizontal stress, S„ is vertical stress, and v is Poisson's ratio. The derivation of Eq. 2 is dependent on lateral con¬finement of the rock mass. An increasing amount of information is rapidly becoming available which indi¬cates that the in situ stress field at a particular location cannot simply be assumed to be a gravity-induced stress distribution in which the secondary principal stresses in the horizontal plane are due to confinement and Pois¬son's effect. Rather they can be a complex function of gravity loading, regional stresses, structural features, ex¬cavation geometries, and variations in surface and/or possibly basement rock topography. Furthermore, the underground principal stress dis¬tribution is not necessarily vertical and horizontal as is often assumed. The deviation of the three-dimensional stress distribution from vertical and horizontal can be due to a combination of both topographic (Hooker, Bickel, and Aggson, 1972) and structural features.

Jan 1, 1982

At 5:30 p.m. a social hour was held for all RMCMI registrants on the lawn area south of the Gondola Lift House. This was followed by a western steak fry and dancing to the country western band, "The Cat's Night Out". MORNING SESSION Tuesday, June 29,1982 RMCMI golfers came out in force for a shot-gun start of the Championship Golf Tournament at Vail Municipal Golf Course which began its first session at 7:30 a.m. followed by a second at 1:15 p.m. The tournament was headed by William Carpender, chairman and his committee members Gary Bennewitz, Jr. and Gene Christmann. RMCMI members were treated to early morning rolls and coffee in the Conference Center preceding a showing of the RMCMI slide/tape program on enhancing the image of coal mining, by telling the story of the Institute. Following this, the meeting was called to order by President McKeever at 8:30 a.m. when he introduced Mr. Carl Terzian who presented a workshop entitled "Sure You Have an Image - But What Kind?". MR. McKEEVER: Good morning Carl. I see you're ready to address this bright and busy tail audience. One of the things that I really want to point out this morning and I knew we would have a second chance with Carl Terzian but Carl is someone that will leave a good lasting impression on you and what he will hopefully be talking about, as I remember, is that what he's saying here can be used personally, or for company, or for the Institute, or anything else on image. Carl is Chairman of the Board of Carl Terzian Associates. He is a prominent author, lecturer and consultant. He is headquartered in Los Angeles and he's nationally known: governmental affairs, man¬agerial development and executive speech training and has served nearly four hundred clients since 1969. Each year he travels nearly 300,000 miles to address conventions, conferences and sales and management meetings as a motivational speaker. Carl has been speaking professionally for more than twenty-five years. He was graduate magna cum laude 1957 from the University of Southern California where he was elected Phi Beta Kappa and student body president. He later served as State Department goodwill ambassador for President Eisenhower; public relations director for Lutheran Hospital Society of Southern California; dean and professor of political science and speech at Woodbury University and public affairs director for architect Charles Luckman. He has been honored for his civic and philanthropic and professional accomplishments by the Congress and State Department, California Assembly and Senate, City and County of Los Angeles, Her Majesty the Queen of England, and six times by the Freedom Foundation of Valley Forge. Carl, welcome to our Coal Mining Institute and it's been kind of a historical situation on Tuesday morning that we will probably have people straggling in after that wonderful steak fry that we had last night. You're very welcome to come and takeover the command. MR. TERZIAN: Thank you very much Ira. This gives me an opportunity for those who were in my luncheon audience yesterday to again thank you for giving me the opportunity of that platform and for the very kind comments that were made following the address. Those of us who had been working for some months with Ira, Charlie, Betsy and many

Jan 1, 1982

Comparison of the Zadra, Anglo-American and organic solvent procedures for desorption of gold from activated carbon loaded in carbon-in-pulp gold circuits indicates that the Anglo American and the organic solvent procedures are considerably more effective than the commonly used Zadra procedure. Organic solvent elution procedures, particularly those using acetonitrile and acetone, can provide effective gold elution and very high concentrations of gold in eluate solutions. The operational simplicity and effectiveness of the Anglo American procedure would appear to make it the most satisfactory procedure available at the present time but this may change as the solvent elution procedures become better developed. INTRODUCTION The carbon-in-pulp technique has been introduced into the gold processing industry to by-pass the often lengthy CCD and filtration steps required in the conventional Merrill Crowe cyanidation procedure for gold recovery. Adsorption of gold on to carbon is also a valuable technique for recovering trace amounts of gold from low level waste streams from conventional processing procedures.

Jan 1, 1982

By S. K. Schwartz

"This paper describes the general layout of the environmental monitoring scheme at Devco 's No. 26 Colliery, including wall face ventilation, fan performance and methane drainage, and then goes on to discuss the sensors, the outstation layout, the surf ace control room and the protection of the system.IntroductionCape Breton coal mines have been pioneers in the use of the latest technology in mining coal on this continent. To mention just one example, continuous methane monitoring at the coal face was started in 1976 and the idea of automatically stopping the face conveyor during high gas levels using the face signalling system was conceived and developed by CBDC engineers.With the introduction of remote environmental monitoring, No. 26 Colliery will be the first underground coal mine in Canada and possibly in North America to have computer-based comprehensive environmental monitoring.Traditionally, surveillance of the mine environment has been based mainly on periodic checks by portable or hand-held monitors. Although this method proved adequate, it has some limitations, namely:1) restricted 'number of observations and samples;2) relatively long time between taking measurements and availability of results for assessment.With remote monitoring, the above limitations are overcome because (a) monitoring is continuous and (b) monitored data are updated quickly (typically 5-7 sec.). Remote monitoring also provides a centralized 'information center' at the surface from where the entire mine environment can be monitored. Details of modes in which data can be displayed and features available at the location will be discussed later in the paper."

Jan 1, 1982

By Suzann C. Ambrosio

The mineral industry in Burundi continued to be a minor contributor to the 1980 gross domestic product (GDP) estimated at $293 million in constant 1970 prices.2 Mineral export revenues, primarily from cassiterite, increased by a factor of 15 between 1979-80, reaching approximately $434,000 by yearend 1980. Gold, bastnasite, colombite-tantalite, and wolfremite continued to be exploited on a small scale, using manual and semimechanical methods. Despite a sharp decline in mining activities during 1979, attributed to a lack o? profitability and shortage of construction materials, the mining sector exhibited signs of recovery during 1980. The mineral exploration budget grew 32% over 1979 levels and was equivalent to approximately $42,000 in 1980. In addition, the Burundi Geographical Institute (IGE BU) was created in September 1980 and given responsibilities to collect cartographic, topographic, hydrogeologic, and other geomorphological data., SONIBUROM, the Burundian and Romanian Government joint mineral corporation, conducted mineral deposit inventories during 1979-80 and has been charged with exploiting some of these as yet unspecified resource areas. Another Burundian-Romanian effort, called Mabaya-Butahana, was initiated to search for gold deposits. Meanwhile on the Ruzizi plain, alluvial gold prospecting, with an estimated grade between 0.2 to 1.0 gram per ton, was carried out on a small scale.

Jan 1, 1982

By Robert B. Parsons

"The budget that was tabled by Finance Minister MacEachen on November 12, 1981 will result in a major tax reform of Canada's income and sales tax systems, the likes of which have not been witnessed by taxpayers since 1971. All segments of the economy will be affected one way or another by the far-reaching budget proposals; the mining industry did not escape untouched.Important changes will be introduced which will alter some of the fundamentals of structuring mineral resource ventures in Canada. This month, Tax Notes examines these changes, and also summarizes the other highlights of this key federal budget.IntroductionFinance Minister MacEachen's budget, with its lower personal tax rates, loophole plugging and tax shelter bashing, represents the biggest shakeup in our tax system since tax reform in 1971. In his budget, presented in the House of Commons on November 12, Mr. MacEachen proposed substantially lower tax rates for middle and upper income individuals, with the top federal and provincial tax rate dropping to about 50%, but in return he eliminated most of the opportunities for individuals, and corporations, to reduce or defer taxes through a wide variety of tax shelter and tax planning mechanisms.Although many of those budget proposals that dealt with the resource industries focused on the implementation of the September 1, 1981 agreement reached by the federal and Alberta governments concerning the National Energy Program (see Tax Notes in the December 1981 issue of the Bulletin), the budget also contained some proposed amendments that would have significant implications for investment in the mining industry.This month, Tax Notes examines these latter proposals, and also summarizes very briefly those budget proposals which are not unique to the resource sector."

Jan 1, 1982

By W. F. Cannon

Many high-grade manganese deposits of the world occur in marine transgressive sedimentary sequences deposited on stable platforms. These include the Nikopol and related deposits in the Soviet Union and Groote Eylandt, Australia, among others. We suggest that the Atlantic and Gulf Coastal Plains and the vast area of Cretaceous rocks in the middle of North America are favorable provinces for sedimentary manganese deposits because regionally transgressive units are widespread. North America has 16% of the world's land area but only about 0.2% of known high-grade manganese reserves. This disparity is difficult to reconcile in view of the fact that large areas of North America are underlain by rocks that we consider at least broadly favorable to contain manganese ores, and we suggest, therefore, that important high-grade deposits remain to be found in North America.

Jan 1, 1982

By Dale Dixon, A. R. MacPherson

The present comminution process in principle has remained more or less the same for the last 50 years. The majority of the changes have been in the fields of equipment improvements and size increases of prevailing systems. The development and utilization of completely new systems has virtually been non-existent. As energy and steeel costs increase and lower grade ores are mined, the comminution costs per ton of ore treated must improve accordingly or the operations will not be profitable. The chapter examines further developments in existing systems and touches on relatively unexplored methods of comminution appropriate for future needs.

Jan 1, 1982