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By Terry Quinlan

The financial evaluation of an orebody uses estimates for the grade of ore produced during each year for the life of the mine. Assay data from the Dolphin Scheelite Mine is used to estimate grades by the Polygonal, Ellipsoidal and Disjunctive Kriging Methods. The results are compared with simulated production data, to show the existence of a bias associated with each method. The affect of bias on the financial evaluation of a theoretical model ia assessed.

Jan 1, 1979

By Euclid P. Worden

In 1977, an article in World Mining (1) reported a total of 2726 rotary blast hole drills in service. This figure included only those machines produced by three of the manufacturers building drills of this type. The total, therefore, is much higher: Rotary drilling in the mining industry is big business to the manufacturers and suppliers involved, and a significant cost item, to mine operators. The time has come to take a hard look at rotary drilling technology with an eye to re¬ducing costs by increasing drilling efficiency. The efficiency of any rotary drilling operation is measured in terms of the rate of penetration achieved by the bit. In blast hole drilling, five factors affect the rate of penetration. They are: 1. Bit selection 2. Weight on the bit 3. Rotary speed 4. Hole cleaning 5. Formation properties Of these factors, all but one may be control¬led by the operator. Looking at any one in detail may reveal possible improvements in the system. Cleaning the hole is a good place to start. In order for a bit to perform at its maximum capability, each tooth or insert should strike fresh, clean rock as it falls into its drilling position. If the tooth or insert lands on previously drilled cuttings, a major portion of the energy delivered to the bit is absorbed in re drilling this broken material. A major portion of the interest, therefore, should be directed to the proper cleaning of the bottom of the hole.

Jan 1, 1979

By G. Norman Owen, Roger E. Scholl, Tor L. Brekke

INTRODUCTION Underground structures are generally regarded to be safer than surface structures during strong seismic motion, except when crossing a causative fault. Recent studies of observed earthquake damage to tunnels generally support this view, although instances of severe damage and, occasionally, collapse are reported. For proposed underground construction of sensitive structures such as nuclear power plants, liquefied natural gas facilities, and nuclear waste repositories, a verification of the seismic response of underground structures will be required.. This paper discusses the stabilization recommendations used in a study of additional costs required to protect tunnels against ground motion caused by earthquakes and underground nuclear explosions (1). The study is a part of the site investigations for possible construction of a nuclear waste repository. The additional costs estimated in the study are not presented here; only the stabilization measures are presented in the present paper. The tunnels, which would constitute the underground facilities of a possible repository, are assumed to be 5.5 m (18 ft) in diameter and to be located at a depth of 600 m (2,000 ft). Three candidate geological materials are considered: argillite (shale), granite, and tuff. Several peak ground accelerations are used: 0.3g, 0.5g, and 0.7g for earthquakes, and 0.5g, 0.7g, and 1.0g for underground nuclear explosions. Potential damage due only to shaking is considered here. This paper does not discuss problems that result from creep or sudden rupture of intersected faults or from ground failures at portals.

Jan 1, 1979

By R. B. Tippin

The material balance computer program (MAT-BAL) developed at the Minerals Resource Research Center of the University of Minnesota, has been used in conjunction with routine process profile samples taken in the potash compaction plant at Great Salt Lake Minerals and Chemicals Corporation, Ogden, Utah. In this manner, a uniform data base is being established for the process, and selected key indicators are used for decisions in plant operation and maintenance. With present emphasis on improving product quality and increasing production rates, the standard data base system provides a means of measuring results from test programs to optimize the circuit. The development and application of this process control method is discussed. The compaction plant computer-assisted data base has been in existence for only two years, but it has made significant contributions to the knowledge of the process and continues to assist in monitoring the plant performance. It is anticipated that the material balance program will be adapted to the GSL flotation plant in 1979, and to other circuits in the potassium sulfate chemical plant.

Jan 1, 1979

By B. W. Stunkard

Hibbing Taconite Company is located by the cities of Hibbing and chisholm near the center of the Mesabi Range, approximately 169 km (105 miles) from the Canadian Border. The plant site is equidistant from these two communities, about 6.4 km (4 miles) from each. A 169 km (105 mile) railroad connects us with the port of Superior, Wisconsin, where the pellet production is loaded into lake carriers for lower lakes destinations. Immediately to the south of the mine area is the world famous Hull- Rust-Mahoning Mine which produced more than 700 000 000 million tons of red ore between the years 1893 and 1973.

Jan 1, 1979

Stratabound uranium deposits consist generally of two comingled geologic and statistical populations. Barren zones, usually having radiometric assays too low for recording on digitized logs, comprise a large portion of a mineralized stratum. In between are mineralized zones; where the grades are distributed essentially lognormally. Estimation of open-pit ore reserves requires the prediction of the tonnages of selective mining units (the smallest mined volume which may be segregated as ore or waste) and their grades within panels of convenient size for open pit planning and economic analysis. The proportionate tonnage of mineralized selective mining units within each panel can be predicted by kriging an an indicator function assigned to each composite (1 if mineralized, 0 if barren). Next, using logarithms of mineralized assays, the conditional lognormal distribution of grades of mineralized selective mining units can be found. Using this distribution and the krigedestimate of the probability a selective mining unit is mineralized, the grade and tonnage of selective- mining units above any cutoff can be predicted. A case study illustrating this two- step procedure is presented for the Imouraren uranium deposit located in the Republic of Niger. Using cross validation, the estimation procedures were checked and were shown to be adequately accurate for mine planning purposes.

Jan 1, 1979

By Joseph Cervik

Common methods of controlling gob gas in U.S. nines are by means of ventilation of gob areas and gas drainage through surf ace boreholes. Costs of drilling surface gob holes increase as deeper coalbeds are mined. Where gob holes cannot be drilled because of topography, ventilation is the only means of control. Even then, sufficient air may not be available to dilute methane in the bleeder entries to acceptable levels. Thus, new methods of gob gas control need to be developed that are independent of surface topography, and of the mine ventilation system, and less costly than present systems. Except for slight variations in procedures and equipment, Great Britain, West Germany, France, Belguim, Poland, and Czechoslovakia use a common methane control method that is different from the system generally used in the United States, This Bureau of Mines paper reviews European mining methods, mining conditions, and methane control technology and discusses their applicability to U.S. mines.

Jan 1, 1979

By L. A. Beckett

Load-Haul-Dump machines (LHDs) have contributed significantly to high productivity in hard-rock mining in recent years and are now being introduced into the coal mining industry (Rao, Haycocks and Lucas, 1975; Mining Magazine, 1975). The superiority of the LHD when compared to loader-plus- shuttle-cars has been demonstrated for several operations (Rao et al., 1975), and the number of manufacturers now in the coal mining LHD market (Mining Magazine, 1975) indicates that the concept will be of considerable future importance in the U.S. coal industry. Currently more than 90 percent of U.S. underground coal production is obtained by room-and-pillar methods (Coal Age, 1976). This system presents certain complexities with regard to haulage analysis; among these are constriction of travel routes, variation of haulage distances as different faces are worked, and change in the overall physical layout as mining progresses. Furthermore, current experiments in the use of more than one continuous miner in a section (Mining Magazine, 1975; Coal Age, 1976) suggest the possibility of highly intensive mining operations, with loading proceeding from more than one point simultaneously, in which the overall patterns for machine movements must be carefully arranged to minimize delays.

Jan 1, 1979

By J. T. Crawford

Computer techniques for open-pit limit analysis were introduced into Kennecott Copper Corp. (KCC) during the early 1960's. The first computerized approach used block matrix mineral and economic models, and a removal increment mining simulator. A detailed description of the approach was published in the May issue of Mining Engineering (Carlson, et al., 1966) and also in Chapter 30.3 of the SME Mining Engineering Handbook (Pana and Davey, 1973). Since then, many refinements have been made to improve the system and the resulting pit limit design. The fundamentals, however, generally remain unchanged. Computerized open-pit limit analysis is presently used by Kennecott's open-pit copper properties and by the Quebec Iron and Titanium Corp. The system and the fundamental evaluation techniques it represents are cornerstones of all open-pit mine planning activities within KCC. This case study describes a typical example of open-pit limit analysis within KCC. The case involves a copper sulfide porphyry deposit with additional minor amounts of molybdenum, gold, and silver. The ore-grade mineralization within the pit design will be mill feed for a sulfide ore concentrator, and subore grade mineralization will be classed as waste and placed on dumps. The contributions to production and profits from dump leaching are insignificant, having minimal effect on the pit economics and design, because the material responds poorly to leaching solutions.

Jan 1, 1979

By Eero Pokki

The Päijänne Tunnel is being constructed to conduct raw water from Lake Päijänne to the Helsinki Metropolitan Area. The length of the tunnel will be 120 km and the cross-sectional area is 15.5 sq m. Excavation started in 1973 and will be completed in 1982. 94 km of the tunnel have been constructed from Lake Päijänne towards Helsinki. The construction has advanced according to plan, and due to Finland's hard bedrock there have not been any great surprises. When completed, the tunnel will meet the water requirements of the Helsinki Metropolitan Area for a long time in the future because, with a free fall of 36 m, 9.6 cu m/s of water can be conducted through the tunnel from Lake Päijänne to the Helsinki Area. With pumping it will be possible to obtain a flow of 15 cu m/s, although at first the discharge will only be 5 cu m/s. The tunnel was partly taken into service as early as March 1979 when it was possible to conduct water to the municipalities north of the Helsinki Area. It is already possible now to conduct water to Helsinki, too, but from the Korpimäki pumping station onwards it must be done through the River Vantaa, which is polluted. When the tunnel is completed in 1982, good quality water can be conducted from Lake Päijänne to the Helsinki Metropolitan Area through a continuous rock tunnel. The cost estimate in 1973 was $ 88 million while it is now $ 90 million notwithstanding high inflation. AREAS NEEDED FOR CONSTRUCTION The areas needed for construction were bought through voluntary deals. Areas of 4 - 5 hectares were bought along the tunnel line at about 5 km intervals. The excavation of access tunnels was started from these places, and the crushed rock from the main tunnel was also transported there. However, it was not necessary to buy the area

Jan 1, 1979

By L. A. Romo, J. A. Ferrer

INTRODUCTION The principle line of the Caracas Metro is located generally along the most important east-west traffic corridor of the city from the yards and shops in Propatria at the western extremity to Petare, and eventually beyond to Palo Verde at the eastern end, a route distance of 21 kilometers. There will be 23 stations along the alignment, underground, at grade and elevated, and the line is forecast to handle 1,209,100 passengers daily by the year 2000 with a travel time of 29 minutes between Propatria and Petare. This principle line was divided into three sections for design purposes: Propatria-Fuerzas Armadas, Section I, Fuerzas Armadas-Chacaito, Section II and Chacaito-Petare, Section III. These sections were further subdivided into subsections for construction purposes. The Venezuelan Government, through its Caracas Metro Office in the Ministry of Public Works, awarded design contracts in sections to PARSONS BRINCKERHOFF-TUDOR-BECHTEL for Propatria-Fuerzas Armadas, Section I, OCOIDESA-GRID-APM for Fuerzas Armadas-Chacaito, Section II and OCOIDESA GRID-APM for Chacaito-Petare, Section III. Construction contracts were awarded by international bidding and work is now underway, by joint venture organizations of Venezuelan and foreign firms. The first two sections are under the administration of a company of Engineer/Inspectors in each section. The construction supervision of contracts in the first section is under the administration of the Venezuelan firm, Gómez-Celis-Herrera, (GCH) Engineers, acting under contract to the new C.A. Metro de Caracas that replaced ' the former Office of the Metro. The Propatria-Fuerzas Armadas section, approximately seven kilometers long, consists in part of the 2,951 linear meters of twin-bore

Jan 1, 1979

By Robin Adams

After nearly five years of economic hardship, the world copper industry is, in our opinion, about to enter a period of considerable strength. This opinion is based on three main propositions, namely: 1) there are a number of reasons, both general-economic and industry-specific why copper demand is likely to perform significantly, more strongly in the 1980's than in the 1970's; 2) the virtual cessation of investment in new facilities, especially at the mine level, over the past years makes it physically impossible for capacity to expand as fast as demand in the mid-1980's; and 3) there is much less "excess" inventory available than most analysts believe. If we are correct in these assumptions, then the implication is that copper prices will climb up to, and possibly beyond, the cost-justified price level. Combined with general inflation, this suggests a price in excess of $2/pound by 1985 or roughly treble the 1978 average.

Jan 1, 1979

By Timothy P. Smirnoff, Arthur P. Chase, Herman J. Smith

INTRODUCTION The Newtown Creek Pollution Control Project (P.W. 53) is a major system of facilities designed to collect wastewater from large areas of the Boroughs of Manhattan, Brooklyn and Queens, then channel it to a treatment plant where the wastewater is processed and discharged. The North Branch interceptor collects wastewater along the east side of Manhattan and transports it downstream to a Junction Chamber at Avenue "D" near East 13th Street. The South Branch interceptor similarly collects wastewater from the southern part of Manhattan Island on both east and west sides of Manhattan and transports it to the same Junction Chamber from where both flows are delivered to the adjacent Manhattan Pumping Station which pumps the wastewater under the East River to a treatment plant in Brooklyn. This system is shown in Figure 1. Construction of the Manhattan Pumping Station began in the Summer of 1966 and was delayed by a number of mishaps. In the vicinity of the North and South Interceptors, approximately 3,060 cubic meters (4,000 cubic yards) of soil flowed into the excavation: 765 cubic meters (1,000 cubic yards) at the North Branch Interceptor, and approximately 2,300 cubic meters (3,000 cubic yards) at the South Interceptor. These movements were accompanied by surface subsidence, collapse, or both. On June 4, 1973, while pumping station construction was underway, approximately five to seven cubic yards of sand entered the junction chamber through a temporary bulkhead valve in the North Branch Interceptor.

Jan 1, 1979

By Robert P. King

In April, 1975 the Colorado School of Mines began monitoring rock mass response associated with a long- wall shield demonstration project near Raton, New Mexico. This project was sponsored by the U. S. Bureau of Mines (DOE) and performed at Kaiser Steel Corporation's York Canyon Mine. Two of the results of the intensive instrumentation effort are the empirical subsidence and horizontal strain models presented herein. These models are based on data gathered during mining three adjacent longwall panels (4N, 5N and 6N). A total of 251 permanent surveying monuments were monitored over the three-panel area. Lines of monuments were oriented along panel centerlines, perpendicular to centerlines and diagonal (45O) to centerlines. Active monuments were monitored twice a week, with less frequent monitoring of stable or slow moving monuments. Primary subsidence monitoring equipment consisted of an AGA Geometer Model 75 and a Wild T-2 Theodolite with -a Wild-Sercel DI-3S distomat.

Jan 1, 1979

By C. B. Manula

This paper presents the results of an application of the PSU/USBM Open Pit Materials Handling Simulator (1) to a large stripping mine in the Southwestern Coalfield. The specific objective of the study was to demonstrate the application of the simulator to a complex multi-pit operation, with emphasis on simulating short term and long term production plans recommended for the mine by Theodore Barry Associates (2). Mine Description The mine is located in the Southwestern Coalfield, which encompasses parts of the states of Utah, Arizona and New Mexico. The lease area presently covers 121500 x 103 m2 (30000 acres) in a strip 37.007 x 103 m (23 miles) from north to south and up to 1609 m (1 mile) in width, located at an elevation above 1525 m (5000 ft). Three gently dipping coal seams are present, of which the upper seam (#3) averages 0-5.5 m (0-18 ft); the intermediate seam (#2) averages 0-1.2 m (0-4 ft); and the lower seam (#1) averages 0-2.4 m (0-8 ft). Overburden is mixed sand, gravel, calcareous clay, caliche and sandstone, with calcareous shale and sandstone burden between seams. Where the seams are covered by less than 6.1 m (20 ft) of overburden, they are generally weathered and oxidized and are not extracted. Stripping ratios have ranged from 3.5 to 6.0 with an average of 5.4. All future plans were based on a stripping ratio of 5.3.

Jan 1, 1979

By C. F. Lee, Bezalel C. Haimson

A 300 m deep test hole at Ontario Hydro's Darlington nuclear plant site, 65 km east of Toronto, was used to determine the state of stress in the Paleozoic limestones (O-22O m depth) and the Precambrian granitic gneiss (220-300 m). The results of the hydrofracturing measurements indicated a state of high horizontal stress along the depth of boring and a consistent orientation of principal stresses in each geologic unit. These results were confirmed by a borehole TV camera survey, and by overcoring tests. The magnitudes and orientations of the measured in situ stresses were used in a feasibility study of siting an underground nuclear plant, and in particular in the geotechnical design of proposed reactor caverns in Precambrian rocks. Their implications on cavern stability under operational and accident conditions were analytically studied, taking into consideration the potential superposition of thermomechanical stresses under loss-of-coolant condition, and the possibility of thermal spalling in the host rock.

Jan 1, 1979

By Oscar Letelier Gonzalez

The goal of mine planning is to determine "the best" strategy to exploit an ore deposit during its workable life. Such strategy must be consistent with the objectives of the company and must be subject to the contraints imposed by economical, physical and geological conditions. Mine planning process may be broken down into three categories, according to the planning horizon: Long range, Short range and Operational. The shorter the scope of planning, the higher the level of details and the lower the degree of freedom associated to each plan. A long range plan defines the economic limit of mining at the mining life end of the deposit; it also can be used as an aid in the evaluation of the economical potential of a mineral deposit and as a guide for short range plans.

Jan 1, 1979

By W. C. Foshee

The large volume of water used in coal preparation plants makes water recycling a necessity. Economical cleaning of dirty water usually requires flocculation with an efficient organic polyelectrolyte. This paper discusses identification of optimum flocculant systems. Supporting laboratory data with on-site tests and full scale field trials supplies information necessary for matching the polymer to the plant.

Jan 1, 1979

By Franklin B. Woodbury

As part of its goal of helping maintain an adequate supply of minerals to meet national economic and strategic needs, the U.S. Department of the Interior's Bureau of Mines is evaluating alternative technologies to treat oxidized taconites. Since process control is an essential element in the application of these process technologies, research was performed with a prototype monitoring system. A Californium-252 neutron source and a thermal neutron capture- gamma ray spectra analysis method was utilized to measure the amount of iron and the percent solids in the process slurries. The concentrate and tailing streams in a 410-kilogram-per-hour (900- pound-per-hour) pilot plant employing a selective flocculation and cationic flotation process were monitored. The results obtained showed that the thermal neutron capture-gamma ray reaction can be used to provide process control parameters.

Jan 1, 1979

By Ernst Langer

INTRODUCTION Morro do Niquel is located near the small town of Pratdáolis in the southwest part of the State of Minas Gerais, Brazil, about 300 km in a straight line north of São Paulo and 300 krn west of Belo Horizonte. The deposit is known since 1935. Systematic exploration started, however, only in 1957 and production with one furnace in 1962. Since 1969, after the installation of a second furnace, the mine produces about 2 500 tons of nickel, contained in ferronickel, per year. This production is now consumed totally by the Brazilian steel industry. The "Morro do Niquel" is a serpentinite stock of 900 by 400 m which rises 200 m high above the surrounding metamorphic rocks of Algokian age (see Figure 1). On top of the hill there is a leached barren zone of about 200 by 400 m, generally 20-30 m deep, which consists practically only of a boxwork of quartz colored brown by limonite. Surrounding this zone on the sides of the hill is weathered, decomposed serpentinite which contains about 213 of the economic nickel mineralization. Roughly 113 of the ore reserves lie below the leached overburden representing ore of a different type. The geological map and cross-section, Figure 2, give a clear picture of the occurrence.

Jan 1, 1979