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By R. J. Matetic, Peter G. Kovalchik, Jeffrey Shawn Peterson

The National Institute for Occupational Safety and Health (NIOSH) has initiated a study of a roof bolter to reduce noise exposure to mine roof bolter operators. An important segment of this research entails determining the affect of various drilling configurations on the performance (penetration rate) and sound power emissions when drilling into granite. Test conditions included various combinations of rotational speeds and thrusts; hex and round drill steels of one inch (1.) and one and three-eighths inch (1 3/8.) diameters; vacuum and wet conditions; and using drill media of differing compressive strengths. This paper will detail the affects of each of these on the performance and sound power level emission. When drilling into granite, it was found that wet drilling generated lower sound power levels and performed better than vacuum drilling.

Jan 1, 2005

By Drupad B. Desai

Baltimore Region Rapid Transit System project final design began in 1973. The Section A of the planned Phase I is scheduled to go into revenue service in late 1982. The subway line structures include earth tunnels, rock tunnels and cut-and-cover structures. Final designs were performed by both the DMJM/KE staff and by outside subconsultants. This paper deals with Section A of the program. (Fig. 1.) The overall program cost for Section A is estimated to be $721 million. The cost for construction and owner furnished equipment is estimated to be $600 million, including contingency and escalation to time of construction. The remaining amounts are for right-of-way, design, construction management and agency (MTA) costs. The Section A line structures, including station lengths, are divided as shown in Table 1. [ ]

Jan 1, 1979

By J. R. Larsen

The perlite review arrived too late to appear in the June issue. Worldwide production of perlite in 1991 was estimated to be 1.7 Mt (1.9 million st). Events in Eastern Europe and the former Soviet Union caused reporting of production, consumption and reserves from those regions to cease. The US Bureau of Mines reported data only from the US and Greece. Together, the two countries produced about 701 kt (773,000 st), down from 739 kt (815,000 st) in 1990. US production of perlite amounted to about 547 kt (603,000 st) last year. US production was supplied by 10 companies operating 11 mines in seven western states. There are 74 plants producing expanded perlite in 33 states and in Canada. All the companies engaged in mining perlite are also expanding it.

Jan 1, 1992

By George Aul, Richard Ray

Short-hole horizontal degasification systems have been utilized in Island Creek's Virginia Division mines since 1986. These systems remove methane from the coal seam in advance of longwall mining. Without an in-seam methane drainage program, methane liberation rates from the longwall faces may exceed 0.47 cu m/s (1,000 cfrn), requiring 47 cu m/s (100,000 cfm) of air for adequate dilution capability. Optimization of the degas systems has enabled methane liberation levels and the associated ventilation requirements on the longwall faces to be reduced by 50 to 80 percent. Since the inception of the first degas system at the Virginia Pocahontas No. 6 Mine, Island Creek has developed an overall methane drainage program designed to maximize gas production and minimize ventilation requirements for methane dilution. Gas production from the degas systems is affected by four main design criteria: (1) hole location, (2) hole length and spacing, (3) collection system capacity, and (4) time available for drainage. Drilling equipment and materials for the gas collection system are selected to meet the necessary criteria while minimizing manpower requirements. Benefits gained from reducing methane liberation rates from the longwall faces are fourfold: (1) improved safety due to lower methane concentrations in the face area, (2) fewer production delays caused by excessive methane concentrations, (3) lower fan operating costs resulting from reduced air requirements, and (4) enhanced worker comfort resulting from lower air velocities and less dust generation.

Jan 1, 1991

By Tore M. Larsen

The Rødsand Gruber mining company is a subsidiary of the Norwegian industrial corporation Elkem AS, and its mine is situated in the northwestern part of southern Norway. The ore from this mine is a relatively low grade, magnetite ore with a small amount of vanadium. The ore bodies are very irregular in shape, resulting in a high amount of waste rock intrusion in the stopes. The annual underground production in the mines is approximately 1 million t. Because of varying dimensions of the ore bodies, both transverse and longitudinal stopes are employed. The inclination of an ore body is usually 1.04 to 1.59 rad (60 to 90°), and the horizontal width varies up to 100 m. Sublevel stoping has been the traditional mining methods in this mine. Until a few years ago, the ore was primarily blasted using horizontal blasthole fans drilled from Alimak raises. Recently, however there has been an increasing demand for a les expensive mining method. Alternate systems were considered, and finally a method based upon large blasthole stoping was found to be the most suitable. GENERAL LAYOUT The general layout of a stope is shown in Fig. 1 As with all the figures in this chapter, only the prin¬ciples of the method are emphasized and the details may not be valid for any particular site in the mine. In transverse stopes, the horizontal width of the ore determines the length of the stopes. The average length is approximately 50 m, the width of the stopes is 20 m, and there is a 10-m wide vertical pillar between. The distance between the main levels is 60 m. Usually a horizontal (sill) pillar 10 m thick is left below the mucking level. Each stope is first undercut along the entire stope length. This undercut is made with 50.8-mm (2-in.) diam drill holes in a fan pattern. At the moment this is being done by a trackless Atlas Copco Promec jumbo equipped with two Cop 130 pneumatic rock drills. The drill drift is 4 X 4 m, and the crater is 20 m high. This leaves a 36-m stope height to mine with large blastholes. To achieve a fair portion of vertical drilling withlarge blastholes, there are two parallel drill drifts at the top of each stope. These drifts are 4 m high and 9 m wide, separated by a 4-m wide pillar in the middle of the stope. To ease the drilling of perimeter holes, the drifts extend 1 m into the vertical pillar on each side of the stope. The pillar will thus be only 8 m wide on the drill drift level, but 10 m wide down the rest of the stope.

Jan 1, 1982

A panel discussion on carbon-in-pulp processing of gold and silver ores was one of the highlights of the 1981 AIME Annual Meeting in Chicago. The session generated considerable interest and discussion among panelists and the audience. For those unable to attend the panel, the program was recorded and the discussion appears in this two-part report. The panel was co-chaired by Robert S. Shoemaker, vice president of San Francisco Mining Associates, and Laurence D. Hartzog, principal engineer with Bechtel Civil and Minerals Inc., San Francisco, CA. Panel members included: George Potter, consultant, Tucson, AZ; Kenneth B. Hall, metallurgical superintendent, Homestake Mining Co., Lead, SD; and Donald M. Duncan, resident manager, Pinson Mining Co., Winnemucca, NV. There have been several types of carbon-in-pulp adsorption vessels in use, such as the Dorr-type agitator, the simple propeller agitated tank, the Pachuca tank, and, lately, the draft tube-type of agitated tank. Which one of these is best and why? Potter: On the selection of the most appropriate adsorption vessel, there is a considerable difference of opinion that stems largely from the fact that every ore is different. The mesh of grind, the pulp solids, and the apparent viscosity of the pulp as caused by its clay content and chemical conditions all differ. The traditional Dorr agitator with the slow speed center sweep and either peripheral or center column air lifts has worked quite well on the finer grinds of all minus 65 and probably 70-80% minus 200 mesh. Pachuca tanks have also been successful and they may be capable of handling a coarser feed than the traditional Dorr tank. One uranium mill, for example, handles minus 28 mesh sandstone ore in a Pachuca. The most recent development and one that commands consideration is the draft tube agitator in which there is a turbine which closely fits inside a draft tube. Velocity in the tube is carefully calculated to avoid undue shear and thus abrasion of the carbon. The objective in all cases, of course, is to mix the granular carbon, which is typically 6 x 16 mesh, very gently but thoroughly in a slurry with minimum carbon abrasion. I do not know that there is one outstanding choice just yet and I have been unable so far to get information on C-I-P service for ore grinds as coarse as 35 mesh. Duncan: The major advantage of draft tube-type is the ease of startup. At the Pinson plant we installed draft tube-type agitators but it is too early to quote experience with them. We also haven't operated long enough to determine just what our carbon loss is. The advantage, of course, in the draft tube design is that it requires only about one-third the horsepower input of a conventional agitator. If it has no other advantages, it has that. Hall: The type of vessel most suitable for C-I-P adsorption depends on the type of ore treated and the prevailing operating conditions. In most cases, a deep tank with turbine-type mechanical agitator and low speed tip velocities would be satisfactory. Turbine-type impellers give a positive type of agitation which assures optimum ion contact and reduces short circuiting. Thorough aeration is possible with an air sparge properly located. A mechanical agitator can easily be started up after an outage, but it is usually necessary to drain and wash out a Dorr rake-type or Pachuca before restarting. The turbine-type impeller requires more power, but maintenance costs are negligible if rubber covered impellers are used. Carbon losses are minimal. In smaller plants, Pachuca type agitators provide adequate aeration and agitation. Bob Polak, Occidental Minerals (from the floor): Mr. Hall, you favor the mechanical-type agitator with the preface that the proper design is critical. Could you elaborate on this for us? Hall: The most important thing is low tip speed to prevent carbon attrition. I think the impeller should be sized so that you get a good sweeping action toward the bottom of the tank, across its bottom, up the sides, and back to the center. The advantage of the draft tube is that you get a more positive agitation and you probably get improved aeration too. Polak: Do you know of anyone using an upflow mechanical agitator rather than a downflow unit? Hall: I think that Bob Wilson at Custom Equipment has designed one where he has located the sparge directly beneath the impeller. The air bubbles are broken up by the impeller as the slurry passes through it. The slurry flows upward, outward, down the sides, and, again, back to the center of the tank. I don't know that any tanks of this design are in commercial use. Hans Von Michaelis, Randol International (from the floor): A question to any of the panelists on flat bottom versus conical bottom Pachucas. Hall: I would say the conical bottom would be most suitable in most cases. Some plants have experienced problems with flat bottom Pachucas in that they have a tendency to sand in on the sides of the tank so only the center of the tank is active. Duncan: With our draft tube tank we placed an inverted cone in the center of the tank bottom and blanked off the bottom comer around the circumference of the tank to facilitate movement of the pulp. Other than that, as far as conical bottom tanks are concerned, I'm generally opposed to them because of the cost and height differential. Larry Kramer, Kennecott Minerals Co. (from the floor): Mr. Duncan, you mentioned that the draft tube-type could be agitated with about one-third the horse-power applied to a conventional propellor agitated tank. I have trouble trying to pin down that kind of number. Could you give a bit of rationale as to why that mechanism has a lower horsepower requirement? Duncan: I think it has to do with the fact that below the propeller you have straightening vanes. The pulp, which is flowing vertically downward, is turned and flows upward again without any recirculation. You have an inherently less horse-power requirement in that type of tank. The one-third horsepower requirement is a number I obtained from Lightnin, which supplied us with details of the draft tube. There is probably much more recirculation with a conventional impeller type that is wasted motion. Potter: Last week I saw some agitators in South Africa with short draft tubes and actual turbine-type impellers. This plant was handling 200 kt of ore per month. The tanks were flat bottomed, and the agitation appeared to be quite satisfactory.

Jan 8, 1981

By N. Parra, M. N. Herrera, T. Vargas, B. Escobar, C. Gonzalez

originally published May 1998, Vol. 15, No. 2, pp. 15-1 9 Discussion by G. Rossi, University of Cagliari Reply by T. Vargas, University of Chili

Jan 1, 2000

By John D. Rockaway

This report and accompanying engineering geologic map describe the generalized engineering geology of St. Louis County, Missouri for use in determining the most efficient, economical, and environmentally sound program for land development and urban growth. It includes an evaluation of such factors as the physical and engineering properties of soil and bedrock, surface stability, and subsurface geologic structure. The map and text are designed primarily for property owners, engineers, architects, real estate developers, and city planners involved in land-use planning. The geologic information and engineering interpretation must be considered as a general guide and not as a substitute for a complete on site investigation prior to initiation of any land-use project. The map scale, 1:63,360 (1 inch to a mile) is too small to include the detailed information normally provided on the 7 1/2-minute quadrangle series (1:24,000). The smaller scale is most useful at regional planning levels and was chosen to present information on the more significant engineering factors of the geologic environment for county-wide planning and development. A conventional geologic map deals primarily with the distribution and variation of bedrock formations, and a pedologic map presents information on the soil or natural medium for plant growth. Basic geologic maps and pedological maps are valuable tools that can be of assistance in studying the geologic environment, however, these maps do not provide all the information required for land-use planning. The engineering geologic maps are designed to provide this data. The engineering geologic map is related to both of these basic maps but emphasizes the engineering properties of the rock formations and soils rather than their age or origin. Some of the unit boundaries shown on the county map will closely follow some geologic formation and/or pedologic boundaries; others will transgress these conventional boundaries. Earth materials, bedrock and soil, that have similar engineering characteristics have been mapped as one unit without regard to their geologic or pedologic classifications. This is to assure that geological factors are recognized, adequately interpreted, and presented for use in engineering practice.

Jan 1, 1970

By Edmund W. Murphy

At the first workshop on Surface Subsidence Due to Underground Mining (1 981 ) , Yarbrough delivered a paper (1 ) describing the mine subsidence insurance program in Illinois operated as the Illinois Mine Subsidence Insurance Fund (IMSIF) . In this paper we will present claims data and describe changes that have occurred in the Illinois Insurance Code during the first six years the IMSIF has been in existence. The change in the original Illinois Public Act 80-1 41 3 that most affected the public was an amendment increasing the maximum amount of mine subsidence insurance avail- able on a structure from $50,000 to $100,000. Seventy-two counties in Illinois have been undermined for minerals, primarily bituminous coal. In Franklin County (.in the southern part of the state) 36 percent of the land area has been undermined for coal extraction. In St. Clair and Madison Counties , two counties that include the St. Louis Metropolitan area along the western border of the state, 10 percent and 7 percent of the land surface have been undermined. Most of the undermined area in the two counties is beneath urbanized land. In 34 counties, approximately 1 percent or more of the land surface has been undermined (2). The first shaft mine in Illinois was sunk in the city of Belleville (east of St. Louis) in 1842, and the mines of early vintage have initiated the majority of the mine subsidence claims. Data compiled from the first six years reveals that most Illinois mine subsidence events that affect structures occur over abandoned mines where early methods of high- extraction mining practices were used. Maps of these mines operated in the 1800s reveal irregular development of room and pillar areas and poorly defined production areas (panels). Rooms in the mines varied from 15 to 40 feet in width and the pillars varied in width from 5 to 40 feet. The pillars were often long and narrow. Recovery rates in mines that operated prior to 1900 generally averaged 50 per- cent, and extraction rates as high as 80 percent have been documented (2). Some pillar robbing occurred.

Jan 1, 1986

By Ingo Kaundinya

"INTRODUCTION Tunnels are important elements of the German federal road network which have to cope with an increasing traffic load especially of heavy goods vehicles. In order to ensure a high availability and safety level for these key infrastructures, big efforts are required by owners and operators in order to keep all infrastructures up to date. This mainly relates to current European regulations regarding tunnel safety (Directive 2004/54/EG) [1] and their national implementation, which must be fulfilled by many existing tunnels. In this context the German Federal Government spends more than 1.1 billion € for the structural and operational retrofitting of more than 120 existing road tunnels. As many of these projects have recently been finished or are still under progress, experiences and best practices could be reported. After the description of current requirements in national standards regarding structural fire protection, the paper gives an example for an upgrade of a tunnel’s fire protection by installing an additional fire protection system, which is a quite costly measure. Another possibility to fulfill current requirements regarding structural fire protection could be the numerical proof of sufficient bearing capacity of a structure. The influence of different fire loads and fire durations on the bearing capacity of typical road tunnel structures have been investigated in a research project funded by the German Federal Ministry of Transport (BMVI) and coordinated by the Federal Highway Research Institute (BASt). The results of this research project have been used to define a procedure how to investigate existing tunnels regarding structural fire protection. Finally, the research results have also been the basis for a current update of national standards for tunnel construction. REQUIREMENTS REGARDING STRUCTURAL FIRE PROTECTION The structural fire protection design of road tunnels is usually based on temperature-time curves of exposure to fire, for simplicity called ""fire curves"". The fire development after fire growth is illustrated with fire curves. Fire curves can be subdivided into a heating-up phase, a fully developed fire phase, and if applicable, into a cooling-off phase (cf. Figure 1). During the fully developed fire, the fire remains at a certain temperature level."

Jan 1, 2016

By Omar A. Muhtadi

8.1 INTRODUCTION Several methods currently exist by which gold may be recovered from a pregnant leach liquor. Two of these methods are discussed in this chapter, the Merrill-Crowe (zinc cementation) process and the carbon adsorption process. These two methods represent the most frequently used techniques for gold extraction and are, -in fact, used almost exclusively in U.S. heap leaching operations. Because of the efficiency and ease of use of these two methods, it is extremely likely they will continue to be quite prevalent within the gold mining industry in years to come. The remainder of this chapter includes information on the history of the two processes, descriptions of the actual processes, criteria to consider in selecting a recovery process, and commercial design and construction considerations for both of the gold recovery methods. As with other chapters in this book, some fairly technical information is presented for the benefit of those readers needing a greater level of detail. However, the overall intent of the chapter is to provide an overview of the processes likely to be used in gold extraction, thus giving the reader a firm basis for more technical research into the topic, should this be needed. 8.1.1 History of Zinc Cementation During the late 18901s, zinc cementation was introduced for the precipitation of gold and silver from cyanide solutions. This occurred at about the same time as the introduction of the cyanidation process, the history of which is recapped in Chapter 1. The initial cementation process involved introduction of a gold-bearing cyanide solution onto a bed of zinc shavings. It proved to be quite inefficient because the reaction rate was very slow. The zinc quickly became "passive", inhibiting further gold deposition. Shortly after this first introduction, zinc precipitation was improved by adding a lead salt (usually lead nitrate) to the zinc. This allowed a zinc-lead couple to form on the surface of the shavings, eliminating passivation of zinc surfaces and thereby allowing continued deposition of gold. Further improvements were shortly forthcoming. The first of these involved the use of zinc dust rather than zinc shavings. This provided a much

Jan 1, 1988

By Gerald M. Ellis

Introduction Gold exploration in the Dominican Republic has not received much publicity. The fact is, however, that a continuous program has been underway for some time. For the past two years, the Restauracion area of the western Dominican Republic, near the Haitian border, has been the scene of mineral exploration by the Dominican Department of Mines geologists, and later by geologists of Rosario Dominicana. The latter is a state-owned gold mining company and operator of the Pueblo Viejo open-pit gold mine. The exploration is for mineral deposits with the goal of finding another Pueblo Viejo-type deposit. Geology Exploration is being carried out in the western central Cordillera, the backbone of the Island of Hispaniola. Geology consists of quartz veins and veinlets and ramifying stringers carrying pyrite, barite, minor amounts of copper, and significant amounts of gold in places. They are associated with regional faulting in a complex of Cretaceous beds of andesites, dacites, argillites, siltstones, and shales. They are intruded by tonalites, diorites, and gabbros. Lesser amounts of zinc-lead mineralization occur in some locations. The showings appear as disseminations along the faults. To date, the work has been mostly geologic mapping - 1 to 50,000 with details on a scale of 1 to 20,000 and 1 to 10,000-stream and soil lithogeochemistry, and geophysics. Trenching and sampling were used to follow possible vein extensions. Particular emphasis has been placed on the Los Candelones district that was taken over by Rosario Dominicana. Here is where drilling has been initiated in the past months. The first hole drilled showed assays of 21 g/t (0.6 oz per st). The second hole intersected an interesting oxidized zone that reminded observers of the original Pueblo Viejo oxidized zone. A third drill hole went deeper than 75 m (246 ft), the depth to which the first two holes had been scheduled. An assay of 50 g/t (1.46 oz per st) was cut during drilling. The presence of an oxidized zone of any vertical and horizontal extent is important. This gives rise to continued exploration, as the

Jan 2, 1986

By Dan Martz, Stuart Lipofsky, Dean Workman

"The Indianapolis Deep Rock Tunnel Connector (DRTC)—first in a vast network of storm water storage tunnels below Indiana, USA—was a wildly successful endeavor. Crews for the Shea/Kiewit JV drove a 6.2 m Robbins Main Beam TBM to world record rates. The machine achieved 124.9 m/day, 515.1 m/week, and 1,754 m/month in limestone and dolomite rock. The advance rates can be attributed to many factors including ground conditions and knowledgeable crew, but continuous conveyors are also of key importance. The novel conveyor system, manufactured by The Robbins Company, enabled continuous tunneling in a difficult layout that included two 90-degree curves and two S-curves. Spanning 11,777 m in its longest iteration, the system included nine booster drives plus a main drive. A vertical belt moved muck up the 76 m deep shaft to a radial stacker for temporary storage. The system, one of the most complex in North America and the first to operate in 90-degree curves, made swift tunneling possible. This paper will examine the world-class tunneling done at the Indianapolis DRTC and the role of continuous conveyance in reaching high advance rates. The logistics of the system will also be examined as it could apply to future tunneling projects with similarly complex layouts. INTRODUCTION The Indianapolis Deep Rock Tunnel Connector (DRTC) is the first phase of a nearly 28-mile long network of deep rock tunnels being built 76 m beneath the City of Indianapolis, Indiana by Citizens Energy Group. The Deep Rock Tunnel Connector (DRTC) Project is a 12.2 km, 5.5 m diameter tunnel constructed by S-K JV, a joint venture between J.F. Shea Construction and Kiewit Infrastructure Co. The tunnel is a $179 million project and is the first segment and southernmost portion of a deep, large diameter conveyance and storage tunnel system to provide overflow relief during wet weather events. The project is estimated to be completed in 2017 and will initially deliver 204,000 cubic meters of the planned 1 million cubic meters of CSO storage for treatment, as well as provide future connections to other tunnels in the overall system (see Figure 1)."

Jan 1, 2016

By Don Suttie, Frank Roybal, John Wright

The Roybal Raise incorporates the simplicity and low cost of a short bald-headed raise while overcoming the usual impediments to driving a high baldheaded raise. The concept provides for speedy excavation and maximizing the use of the excavated opening, using the rock for structural support. The design is simple: two baldheaded raises are driven parallel to one another, interconnected by dogholes about 8 m (25 ft) apart. In this fashion, the raise can be driven safely and efficiently to 60 m (200 ft) or more. When the raise is complete, the excavation requires no timber for support or partitioning. Rather, the rock itself supports the construction. As a test, a two-compartment Roybal Raise was attempted at the Independence mine in the Willow Creek District of Alaska. One side was driven 2 x 2 m (6 x 6 ft), intended for a manway and equipment skip. The other was driven 1.5 x 1.5 m (5 x 5 ft) for a muck compartment. Both sides were driven at a 65° angle from horizontal and 1 x 1 m (4 x 4 ft) dogholes spanned the 4.6-m (15-ft) pillar between the two compartments. After topping out at 50 m (160 ft), the raise was found to meet all expectations of efficiency, low cost, and practical simplicity. Technique Preparing the Roybal Raise for driving involves no elaborate or costly site prep except assembling materials and providing for muck removal. Normally, the raise muck falls directly onto the sill or track and is picked up either by an overshot mucker or a front-end loader. Raise mining proceeds initially by taking up both compartments simultaneously to the first doghole. It progresses by alternating one side as a clear manway for access with the other side, the active side, as a muck compartment (Fig. 1A). This stage proceeds quickly for no mucking or staging is required to cycle the first three rounds. On the fifth round, in addition to the normal raise round, rib holes are winged out toward the adjacent raise, thus starting the dogholes. Finally, as the sixth raise rounds are drilled and shot, the first doghole is also completed from both sides. The second stage commences by establishing one side as a clear manway for access while the other side, used as a muck compartment, is prepared for active mining. Figure 1B illustrates this step in the succession of raising with the active side advancing to the second doghole level. During this stage, services are maintained through the manway side and equipment is stowed in the doghole during blasting. As the active side advances, muck may be allowed to fill the bottom portion of that side since it is unnecessary to clear it for access. The next stage of raising involves cleaning out the former muck compartment and preparing it for use as a manway. At the same time, the former manway is prepared for advance and use as a muck compartment (Fig. 1C). As in the previous step, the same doghole is used for access and equipment storage. The active side advances until the second doghole is completed, at which point equipment and access are moved to the new doghole. From there, the active side is driven past the second doghole 8 m (25 ft) more to the third doghole level. Again, half of the doghole and the next raise round are mined. The fourth stage involves switch ing back the manway side for the muck side (Fig. ID). The new muck side is driven to the third doghole where equipment storage and access are reestablished. From there, the active side is advanced to the fourth doghole. In the same fashion, the raise is advanced to the top, alternating the manway side with the active side, maintaining access through the inactive manway, and storing equipment in the uppermost doghole. The cycle time for driving each compartment, regardless of the raise height, is greatly improved over that for normal bald-headed raises of comparable height. By storing equipment in the upper-

Jan 1, 1984

Brace yourself-what used to be a 1 billion ton deposit is now a 1 Gt deposit. And that 1520-hp ball mill is now a 1.134-MW ball mill. Beginning with this issue, all engineering units and statistics in MINING ENGINEERING will be reported in SI units. The exception to this is the Technical Papers section, where units may still be reported by dual notation. For the past four years, ME has reported all units of measurement first in SI, followed in parentheses by the equivalent English value. This metrication policy, adopted by the SME-AIME Board of Directors in September 1975, was designed to aid magazine readers through a transitional period from English unit reporting to SI unit reporting. In October 1979, the SME-AIME Publications Board completed the English-to-SI transition by requesting the editorial staff, beginning February 1980, to report exclusively in SI. The Publications Board also requested the editors to prepare a chart of SI conversions useful to ME readers; this information follows on pages 156-157. The What and Wherefore of SI In 1960 the General Conference on Weights and Measures met in Sevres, France, and recommended that a new metric system-one that could uniformly cross all international borders-be adopted to replace all other metric systems. The outcome of this recommendation was the SI system of units, an abbreviation for le Systeme International d'Unites (the International System of Units). It replaces the more familiar cgs (centimeter-gram-second) system and the m/ksA (meter-kilogram-second-Ampere) system. SI consists of seven base units, two supplementary units, and a number of derived units. Other units, called allowable units, may be used even though they are not part of SI. The seven base units are: meter (length); second (time); kilogram (mass); mole (amount of substance); kelvin (temperature); ampere (electric current); and candela (luminous intensity). The two supplementary units are radian, a measurement of plane angle, and steradian, a measurement of solid angle. Supplementary units can be used in combination with SI units, e.g., radiant intensity is expressed in W/sr, watts per steradian. Derived units are combinations of base and supplementary units. Velocity is expressed as meters per second (m/s), and acceleration as m/s'. Using the relation established by Newton's second law of motion, force is expressed as kg-m/s2, more commonly known as the newton (N). Energy (work) is force times distance, expressed as kg-m/s2 and known as the joule (J). Power is work per unit time, expressed as kg, m2/s3 and known as the watt (W). Allowable units are the exception to the rule-they are not taken from base or supplementary units but, nevertheless, are used in the SI system. Volume units are normally expressed as the cubic measure of a length, such as cm 3. But a fluid volume of a cubic decimeter (dm3) is called a liter (L); the "L" is capitalized in another special SI rule to keep the lowercase "1" from being confused with a numeric "1". Other allowable units are degrees Celsius (°C) for temperature; and the minute (min), hour (h), day (d), and year (a) for time measurements. The prefixes in the table on the following page are used with SI units to express the magnitude of a measured quantity. A mill that processes 1000 tons of ore per day would be designated a 1 kt/d mill; a mine producing 10 million tons of ore per year would be called a 10 Mt/a mine. The rule here is to use prefixes that keep the value in a convenient range from 0.1 to 1000. In other words, it would be incorrect to write 25 000 kg-the prefix should be changed to indicate 25 Mt. One of the distinct advantages of SI is that it retains the decimal relation between multiples and submultiples of the base units for each physical quantity except for time. One more rule of SI syntax-numbers are separated into groups of three with a space between each group rather than a comma, e.g., 3,839,174 becomes 3 839 174. By the same token, a decimal number is also broken by spaces, e.g. 0.738153 is written as 0.738 153.

Jan 2, 1980

By D. C. Sisti

A paper was presented during the 1970 SME Fall Meeting on the Stinson Preparation Facilities by representatives of Beaver Creek Consolidated Cool Company and Heyl & Patterson, Inc. This paper summarized a combined approach to development of a flow diagram and selection of cleaning equipment required to produce a uniform metallurgical coal at maximum recovery with special emphasis on sulphur reduction. These facilities were located in the Elkhorn. Coal Fields of eastern Kentucky and, in particular, the area of Wayland, Knott County, Kentucky, where approximately forty-five- million tons of Elkhorn No. 3 coal could he recovered. The facilities incorporated a Bradford Breaker for reduction of the ROM coal to minus 1-1/2 x 0, 6000 ton-raw coal storage, preparation plant, water clarification and thermal drying facilities, clean coal storage and a 4000 TPH unit train loading facility.

Jan 1, 1973

By Dennis Arthur Fenton

The hydrodynamic bearing was first described by, Reynolds in 1886, and has since been widely applied in motors, engines, compressors, hydroelectric turbines, and other machinery. For various reasons, the hydrostatic bearing currently predominates in the large, horizontal rotating machinery field (e.g. grinding mills). However, the current industry trend towards increasing size and energy efficiency has resulted in renewed interest in the advantages of hydrodynamic bearings. These advantages include structural improvements, reduced bearing friction, improved bearing life, improved transient characteristics, and reduced lubricant pump size and power requirements. In particular, the higher energy efficiency results in substantial savings on large machines. This paper describes some aspects of the design of large hydrodynamic bearings based on thirty years, of experience in the design of such bearings on grinding mills ranging in journal size from 2.1 m to 6.4 m, see Table 1.

Jan 1, 1985

By Robert S. Lehto

Zinc concentrates received in La Oroya zinc refinery are marmatitic in nature, contain about 11% iron and traces of other metal values, Figure 1 shows the zinc circuit. The calcine from the roasted concentrates is leached in spent zinc electrolyte. Major part of the impurities is iron which has been oxidized to its trivalent form, This is the origin of the zinc leach residue for which the ZILERET process was developed and so named. The impure liquor from the leach is sent for purification and thence to electrolysis- zinc deposits on aluminum cathodes from which: it is stripped, melted and cast into slabs Typical assays of zinc calcine and leach residue are given in Table 1. Metal values contained in the leach residue, without counting any value of iron, amount to roughly $130 per short ton.

Jan 1, 1969

By William P. Walsh

In December 1983 the author undertook an investigation in east Texas which was organized to learn the feasibility of applying geophysical techniques over shallow lignite deposits. The work embraced experimental high resolution seismic reflection and offset Wenner resisitivity sounds over three 1/4 to 1/2 mile test sections. This site also was sponsored by Woodward-Clyde Consultants (WCC) and a consortium headed by North American Coal Corporation. The principal objective of this investigation was to learn whether the aforementioned geophysical techniques could be applied over shallow lignite deposits (i.e. located at approximately 100' to 300' in depth) and consequently provide structural and stratigraphic information useful for resource and reserve characterization. A secondary objective was to determine if these same techniques could disclose the nature of the overburden and the distribution of ground water therein.

Jan 1, 1986

By C. R. Zegarra

Extraction of manganese, gold, and silver from a low-grade manganiferous-silver ore was investigated by thiourea leaching followed by standard cyanidation. The levels of reagent concentrations in this investigation were 0.0 -1.0 M thiourea and 0.0 -1.0 M sulfuric acid. Temperature and agitation speed were maintained at Z5 °c and 300 rpm, respectively. The dissolution of gold, and silver was shown to be dependent mainly on sulfuric acid and thiourea concentrations. In general, 50.0 % Ag, 98.0 % Au, and 75 % Mn were recovered from a Mn-Ag ore by thiourea leaching. Overall silver extraction of 98.0 % was obtained by cyanidation after thiourea leaching. Thiourea leaching of sodium bisulfite leach residues was also studied, and the total extraction of gold and silver were 97 % and 40 %, respectively. Finally, a two-stage thiourea leach was also investigated.

Jan 1, 1988