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By Drupad B. Desai, Alexander Rosenbaum

In recent years, there has been an increasing use of two-pass tunnel lining system consisting of segmental unbolted expandable precast concrete initial lining and cast-in¬place concrete final lining in the United States. Used so far in soft ground tunneling performed under free air, it often includes a waterproofing or gasproofing membrane placed between the initial and final linings. This paper describes the utilization of this two-pass system for an extension of the Baltimore Metro, constructed in part under compressed air. It offers a description of the extension project, known as Section C; addresses the tunnel lining selection process and the design of the lining elements; examines the effect of the adopted lining system and tunneling technique on the design of the project's major structural components; and describes design modification instructions in order to accommodate constantly changing field conditions.

Jan 1, 1993

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By W. Jack Wilkes, Gilbert L. Butler

TRANSPORTATION TUNNELING The use of tunnels for transportation, namely rapid rail transit systems and highways is on the increase. How much that increase might be through the 1980's is the question that is discussed in this paper. In 1975, there were 218 route miles of subway tunnels for rapid rail transit and 154 lane-miles of highway tunnel. From the viewpoint of many transportation planners and designers, the tunnel is one of the tools which may be used to by-pass many types of obstacles: a mountain, a body of water, or a Central Business District (CBD). Therefore, the transportation tunnel is a rural as well as an urban solution with major emphasis on the latter because of the current transportation needs of our cities. In the case of rural locations, tunnels are usually constructed through mountains in order to eliminate miles of steep grades and dangerous curves, such as the Eisenhower Tunnel which cuts through the Continental Divide in Colorado along 1-70 and saves 10 miles and 40 minutes of driving time. Tunnels also provide an excellent transportation link across bodies of water in order not to interfere with shipping lanes. The Chesapeake Bay Tunnel and Bay Tubes of the BART system are excellent examples of this type of tunnel.

Jan 1, 1976

By T. N. Moore, E. M. De Souza

Vibratory hand-tools can produce Hand Arm Vibration (HAV) Syndrome in their operators. This disease, often called Vibration White Finger, can cause varying levels of hand-arm discomfort and, in severe cases, impairment. In addition to the human suffering, there are also significant economic repercussions for industry. These include reduced operator efficiency, lost time through increased absenteeism and increasing compensation awards. In the mining industry, jackleg rock drills are a particularly troublesome source of hand vibration exposure. As a result, a significant amount of work has focused on redesigning rock drill handles to minimize hand-transmitted vibration. This paper details the results of a study undertaken to determine the effectiveness of using elastomer covered ("cushioned") handles in place of the standard steel design. The results indicate that, for a particular cushioned design, the overall acceleration level experienced at the hand-handle interface can be reduced by as much as a factor of three, relative to the steel handle. It is also shown that this reduction takes place predominantly in the frequency range above 500 Hz, with performance improving with increasing frequency. Qualitative evaluation of the cushioned handles by the drill operator indicates a significant increase in comfort over the standard steel handle.

Jan 1, 1992

By Ted Budd, Vladimir Goubanov

The East Central Interceptor Sewer, commonly known as the ECIS Project, is the largest public works construction project awarded by the City of Los Angeles. The complexity of this $240 million project, together with the challenge created by the mandated 1,000 calendar days for construction completion, make this project a major interest for the tunneling industry. This paper will describe the ECIS tunnel project, currently under construction by the Joint Venture of Kenny (Sponsor) / Shea / Traylor / Frontier-Kemper. Mining of the18.5 km (about 11 miles) long and 4.7 m (15ft-4in) diameter tunnel is underway with four Tunnel Boring Machines mining at the same time. Details on the mining machines, muck handling systems, innovative construction sequence, together with mining progress will be given.

Jan 1, 2003

By Barrie W. Sullivan

The New Celebration Gold Mine near Kalgoorlie in Western Australia was commissioned in late 1986 and the CIL treatment plant currently processes 1.7 million tonnes of gold ore annually. The 1988 geological resource was 13.2 million tonnes grading 2.37 g/t Au in several open pittable deposits with the bulk of gold mineralisation occurring as a stockwork in Archaean quartz feldspar porphyry intrusives along a major regional shear zone. Mine planning, scheduling and operating techniques have focused on minimising open pit dilution whilst maintaining production requirements. This paper describes the successful achievement of this aim in extremely variable ground conditions and orebody widths.

Jan 1, 1989

By Mark J. Logsdon, Barbara Basse

Mississippi Valley Type Pb-Zn deposits are widespread in the midcontinent U.S. and have been mined from the 1800's to present. Some tens of thousands of acres of mined land in the Tri-State District have been designated as Superfund sites as a result of surficial mine wastes and ground waters with elevated levels of metals. EPA considers the surficial mine wastes to be hazardous and a major source of acid mine drainage fAMD) that degrades water resources and threatens human health and the environment. To mitigate potential risks at one site near Galena, Kansas, a group of mining companies who had been identified by EPA as Potentially Responsible Parties IPRPs) proposed subgrade disposal of the mining wastes in shafts and subsidence features. However, this proposal was challenged by the Agency on grounds related to AMD and the potential for degradation of ground water quality. The PRP team developed a program of materials characterization, laboratory-scale leaching, and finally large-scale, pilot testing in the field. Static and dynamic leaching tests at both laboratory scale fusing artificial rainwater under both variably and fully saturated conditions) and field scale (using pH 4.5 ground water) showed that the surficial wastes do not produce acidic leachate and are not a long-term source of metals. Laboratory chemical and mineralogical tests show that the mine wastes contain substantial carbonate, but very limited iron sulfides, consistent with the leach test results. EPA "characterization" tests fEP Toxicity and TCLP) compare very poorly with tests simulating the expected leaching conditions. While these tests would classify the surficia1 wastes as "hazardous, they do not characterize the behavior of the high-volume, low-concentration, coarse-grained materials with respect to environments and disposal options of interest. To be meaningful, characterization of mining wastes must be performed in a manner that simulates the environmental conditions that are actually of interest with respect to a specific decision, such as subgrade disposal in the saturated zone.

Jan 1, 1991

By L. S. McNickle

[Commonly accepted maintenance methods are being outdated by the rapid acceptance and use of more complicated hydraulic circuitry on modern mining equipment. It is no longer adequate to use the maintenance program to merely repair or sustain equipment operation. Ever increasing operating costs place a huge premium on lost production time. It has, therefore, placed the maintenance programs in the position of being required to alleviate equipment malfunctions on production shifts and still attain the maximum unit life from each component part.]

Jan 1, 1961

By M. Menegaki

Restoration of abandoned mine sites is still one of the weak points of mining practice. The latter stands especially for the case of the surface coal and lignite exploitations, where the extended size of the disturbed land and the drastic changes to the geomorphologic features do not facilitate the implementation of a cost effective and environmentally sound reclamation plan. The paper examines the parameters involved towards the legal requirements for the reclamation of the first lignite exploitation in Greece, namely the ?Plakes? mine, located at Euboea, due to the particularities of the site. Then an integrated redesign procedure is presented, aiming at minimizing aesthetic degradation, confronting technical difficulties and preparing the site to accept new uses.

Jan 1, 2005

By J. H. Rowland

Four different types of fire suppression systems were evaluated in large-scale fire experiments to study the effect of air velocity on the effectiveness to extinguish conveyor belt fires. The four different types of suppression systems were water sprinkler, deluge type water spray, and two different dry chemical fire suppression systems. The large-scale fire tests were conducted with 72-in-wide fire resistant rubber belt that met the Mine Safety and Health Administration flame resistant requirement specified in Title 30, Code of Federal Regulations, Part 18, section 18.65 (also known as the 2G Flame Test). Each fire suppression system was tested at air velocities of 500-550 feet per minute (fpm) and 1350-1500 fpm. Two tests were conducted at both air velocities for each fire suppression system. Both of the water-based fire suppression systems were able to suppress the fire to the point that a miner could extinguish it with a fire extinguisher. However, the amount of water needed to suppress the fire to the point where a miner could walk up to extinguish any smoldering belting was greater than the current MSHA regulations require. MSHA regulations only require 10 minutes of water supply to the suppression system. In the test set up neither of these systems would have suppressed the fire had the water been turned off to the system after 10 minutes. The dry chemical suppression systems produced mixed results. Dry chemical suppression system A did not suppress the conveyor belt fire at either air velocity. Dry chemical system B suppressed the fire at 500 fpm air velocity, but produced mixed results at the higher air velocity. Details are presented on the large-scale fire test set-up, arrangement of the fire suppression systems, and conclusions regarding the effect of air velocity on fire suppression system design and performance. Disclaimer: The findings and conclusions in this report are those of the authors and do not necessarily represent the views of NIOSH.

Jan 1, 2009

By Martha McJunkin, Lawrence Pietrzak

This paper describes and demonstrates the application of systems analysis and mathematical modeling techniques in the planning of actual tunnel projects and research and development to improve tunneling techniques. A recently completed tunnel project located in Utah is used as an example. Although cost estimating computer programs are used in tunnel project planning, the introduction of performance modeling, simulating the time to excavate, has added advantages that are explored here. The performance/cost tunneling model (TM) described herein mathematically simulates on a high-speed digital computer the important activities of a number of tunneling techniques accounting for activity interactions, variable external environments, mechanical failure, and other uncertainties. A cost accounting segment is included to facilitate cost-effectiveness comparisons. It is shown that, when combined with informed judgment and appreciation of limitations, the TM can be a useful tool that may help improve tunneling operations and cost estimation, for example, by achieving more balanced system designs or bounding the performance/cost implications of geologic and other uncertainties. Ultimately, this might lead to more competitive bids and greater profits. The need for systems analysis and modeling in tunneling research and development (R&D) is well recognized and documented.(1-5) In R&D one must seek broadly useful solutions to the basic problems of face fragmentation, materials handling, ground support, and environmental control. These solutions must take account of the interdependence of these elements and, for a variety of ground conditions, the limitations or constraints that any innovation in one element may impose on total system performance and cost. In this regard the structure and logic of the TM has been kept as simple and modular as possible so

Jan 1, 1974

By Toshio Fujiwara, Sumikichi Ito, Masao Kitamura

The Tennohji-Benten Giant Trunk Sewer Project has been under construction since 1974 in the eastern part of Osaka, the second largest city in Japan. The entire project is divided into 7 blocks, some of which have already been completed. This 8.8-km sewer tunnel, being driven through a densely populated area, was planned to be constructed by the shield tunneling method. This paper described shield tunneling performances and ground behaviors during the construction of Blocks 2 and 3. An earth pressure shield (dia., 6,750 mm) was employed in Block 3 tunneling. The ground was generally soft and the cover over the tunnel varied from 13 m to only 4.5 m, so that a considerable amount of chemicals was injected to stabilize the ground. A hand-digging type shield (dia., 7,250 mm) was used in Block 2 tunneling under compressed air. For 85 m of the total length of 1.6 km, the soil at the face was extremely soft, and therefore, the ground freezing method was additionally employed.

Jan 1, 1981

By J. R. Johnson, C. G. Stewart, D. K. Myers

INTRODUCTION Radon-222 is an inert, radioactive gas formed by the radioactive decay of radium-226, a long-lived member of the uranium-238 decay chain. Radium is present in varying amounts in virtually all soils and, on the average, about 36,000 pCi (1330 Bq) of radon per square meter of soil diffuse into the atmosphere each day (UN 1977). Radon decays with a half-life of 3.6 days through four short-lived daughters to lead210 and it is these short-lived daughters[ [Ra A, (218Po, t, = 3.0 min) , Ra B, (214 Pb, t;, = 27 min), Ra C, (214Bi, t] = 20 min) and Ra C1 (214Po, t] = 2.5 x 10-6 min)] ]which cause the major health hazard associated with radon (Bale 1951). Atoms of these daughters, either unattached or attached to the ever-present particles in air, are deposited on the surfaces of the respiratory tract; alpha particles emitted in their decay can result in large doses to the cells of the bronchial epithelium lining the respiratory tract. These daughters will be present in air in varying relative concentrations depending on the "age" of the air (time since radon emanated into it) and on the amount of mixing of radon and radon daughter contaminated air with clean air. [The practical unit developed to quantify the amount of radon daughters in air is the Working Level (WL). This unit was historically related to the equilibrium concentration of 100 pCi (3.7 Bq) of the short-lived daughters of radon in one liter of air (cf. Holaday 1969) and is defined as any mixture of the short-lived daughters in a liter of air that have a-potential alpha energy of 1.3 x 105 Mev (2.08 x 10-5J) in their decay to lead-210. The working level month (WLM) was developed along with the WL, and was defined as an exposure to one WL for a working month (170 h). This is equivalent to 2.2 x 107 MeV•h•L-1 (3.54 x 10-3 J.h.m-3). If the average breathing rate is taken as 1.2 m3.h•1 (ICRP 1975), then one WLM is equivalent to inhalation of 4.24 mJ of potential alpha energy.] It is now generally agreed that the inhalation of radon daughters is the major potential radiation hazard in uranium mining, and contributes a substantial fraction to the natural radiation exposure of the general population due to the accumulation of radon and radon daughters from natural sources in buildings. Radon daughter concentrations in modern mines are controlled by ventilation, and by blocking off old working areas (cf. Simpson, 1959). However, before the hazard from radon daughters was recognized, considerably higher concentrations of radon daughters were present in some uranium and non-uranium mines. These high radon daughter concentrations resulted in an increase in lung cancers in the mining population, and it is these results that are our main source of information on the risk of inhaling radon daughters. HISTORICAL REVIEW Many excellent reviews of the history of understanding the health effects of inhalation of radon daughters are available (see, for example, Hueper 1942, Lorenz 1944, Sikl 1950, Stewart 1964, Holaday 1969, Lundin 1971, Cross 1979) but a brief summary of some of the highlights in this area may be of interest. [a] 1556: Agricola describes an unusual and fatal chest disease occurring among underground miners in the region of Schneeberg and Joachimsthal (Jachymov) in the Erz mountains in Central Europe. (It is of some historical interest to note that Agricola's book was translated from Latin into English by a mining engineer and his wife; the engineer later became President of the U.S.A.) [b] 1879: Haerting and Hesse indicated that the majority of deaths among Schneeberg miners were due to lung cancer; the lung cancers in these miners (who were incidentally not cigarette smokers) were observed twenty to fifty years after they began working in the mines. [c] 1896: Discovery of natural radioactivity by Becquerel, followed by discovery of radon by Dorn in 1900. [d] 1924: Ludewig and Lorenser report high concentrations (400 - 15000 pCi or 15 - 570 Bq per liter) of radon in the air in the Schneeberg mines and suggest that radon could be responsible for the high rate of lung cancer among miners. However, the reason why radon should cause lung cancer specifically was still not really understood up to twenty years later (Lorenz 1944). High concentrations of radon in the air in the Joachimsthal mines were reported by Behounek in 1927 and a high incidence of lung cancer among Joachimsthal miners (similar to that among miners at Schneeberg, which is only some 30 km distant but in a different political district) was noted at about the same time (Sikl 1930; cf. Sikl 1950). It is estimated that about half of the Joachimsthal miners died from lung cancer and about half from silicosis and tuberculosis (Sikl 1950). [e] 1930's and 1940's: Radioactive ores are deliberately mined in the U.S.A., Canada and other countries primarily as a source of radium for medical purposes and for luminescent dials; other radioactive ores are mined as a source of several non-radioactive minerals, while extensive uranium mining did not begin until the late 1940's. [f] 1940: Based on crude epidemiology and dose calculations, Evans and Goodman propose 10 pCi L-1 as a maximum permissible concentration of radon in continued human exposure. This is the first known recommended maximum permissible concentration for radon. This recommendation was adopted by the U.S. National Bureau of Standards in 1941 and reconfirmed in 1953. [g] 1945: Mitchell identifies the short-lived daughters in radon as a likely cause of the increased lung cancers in the Schneeberg-Joachimsthal miners

Jan 1, 1981

Jan 1, 1999

By S. R. Rao, J. A. Finch, E. El-Ammouri, K. Ngoviky, P. A. Disten

Sludges produced by lime treatment of acid mine drainage (AMD) can contain significant quantities of base metals which represent both a potential hazard and a loss of values. A new technique for metal recovery from sludge is proposed by combining selective acid leaching and adsorption of metal ions by activated silica sol. The process is demonstrated on two AMD sludges, one containing primarily nickel and the second zinc. They were first treated with sulphuric acid (pH 3.5) to selectively leach the target metals and then activated silica sol was added to the leach solution to adsorb the nickel and zinc. Thickening and centrifuging gave about 90 percent rejection of the input volume to effluent. The adsorbed metals were redissolved by strong acid into a concentrated solution relative to the feed and the silica sol was recycled.

Jan 1, 2000

By Changhong Huang, Y. J. Wang

This paper describes the formulation of mine ventilation network optimization as a standard nonlinear programming problem and the solution of the problem using the generalized reduced gradient (GRG) method. In general, the GRG method is regarded as one of the most effective and efficient methods for solving large-scale nonlinear programming problems. The variables considered in the network optimization include air quantity, regulator pressure loss, and fan pressure for each branch. Each of the variables is constrained by lower and upper bounds. This facilitates proper assignments of ventilation requirements and provides a great flexibility to include various situations encountered in mine ventilation. The objective of the network optimization is to minimize the overall air power consumed by or supplied to the network; whereas the constraints are, in addition to the variable bounds, the two fundamental network laws: Kirchhoffs voltage law and Kirchhoffs current law. A computer program has been developed based on the GRG method for the network optimization. An example of network with 18 branches, 8 regulators and 3 fans is included for illustration.

Jan 1, 1993

By J. Girard-Dwyer, J. Johnson, J. Whyatt

The choice of fragmentation method is one of the most fundamental aspects of mine design, and often has a major impact on the safety and economic viability of an operation. Almost exclusively, underground mining operations rely on either mechanical excavation or drill and blast techniques to extract rock. Each of these fragmentation methods has distinct advantages and disadvantages, and each is appropriate, and economic, for different mining conditions. However, the choice of fragmentation method can have a pronounced effect on the amount of damage created in the rock mass surrounding the opening. This paper reviews a number of case studies where both mechanical excavation and traditional drill and blast were used in underground mines. Ground support requirements in the mechanically excavated areas are compared to the support requirements in the blasted sections in order to demonstrate the potential safety and cost benefits of the various fragmentation methods.

Jan 1, 2003

By E. Suoninen, K. Laajalehto, L. Kartio

Efficient separation of pyrite from other sulfides is of great importance for the metallurgical performance of flotation processes. In this work Fourier Transform Infrared Spectroscopy (FTIR) and X-ray Photoelectron Spectroscopy (XPS) were used to study the surface chemistry of pyrite in depression, activation and xanthate adsorption under conditions of controlled potential. Modifications of pyrite surfaces after treatment with depressants (lime, sulfur dioxide, sodium cyanide) and after activation with metal ions (Cu, Pb) were studied. The principal adsorption product identified on pyrite was dixanthogen whose formation started at about +0.15 V and +0.25 V (vs, SHE) in ethyl and amyl xanthate solutions, respectively. Copper xanthate was formed on copper(II) activated pyrite. Activation mechanism of pyrite by copper(II) salts is likely to be electrochemical where copper occurs as copper(I) on the surface of pyrite. Effective depression is achieved by sulfur dioxide and sodium cyanide. Depression at high pH is due to formation of iron(III) hydroxides. Calcium ions do not affect the electrochemistry but adsorb on pyrite and reduce the surface sites for dixanthogen adsorption.

Jan 1, 1995

By Howard J. Brown

The Lucerne Valley limestone mining district is in the San Bernardino Mountains, one of the major ranges in the Transverse Range Province of southern California. Several large limestone mining operations are present along the north range front in Lucerne Valley, and the area is one of the premier limestone mining districts in the country. Resources are contained in Paleozoic carbonate rocks that are part of a thick late Precambrian and Paleozoic sequence that uncomformably overlies crystalline basement. The district has been affected by multiple episodes of metamorphism and deformation, as well as by a complex sequence of late Cenozoic structural events. Detailed studies, however, have demonstrated that these rocks can be divided into formations correlatable with late Precambrian to Permian strata of the Cordilleran miogeocline. The district is the leading producer of white, high-purity calcium-carbonate products in western North America. Pluess-Staufer (California) Inc. mines and processes high-purity white limestone from extensive deposits, and is one of the largest producers of high-quality limestone fillers and extenders in the western United States. Major factors influencing the genesis of economic carbonate deposits in the district include depositional environment, post-depositional processes such as diagenesis, dolomitization and/or silicification, regional and/or contact metamorphism and associated intrusive activity, faulting, folding, uplift and erosion, and preservation from erosion. Regional geologic studies, detailed geologic mapping, intimate knowledge of stratigraphy, and recognition of complex structural relationships have proven to be invaluable tools in the discovery of high quality calcium carbonate deposits in the district.

Jan 1, 1987

By Deepak Malhotra, William F. Riggs, Edward C. Dowling

INTRODUCTION Mining companies are In the business of making an acceptable profit on their Investments, It became obvious in the recent recession that economic profitability/viability depends on being the lowest cost producer In order to compete on a global basis. The burden of achieving this objective falls on the mineral manager. He must strive to produce a salable product in an environmentally acceptable process while continuously Improving the profit margln.(1,2) Metal prices fluctuate on a regular basis. They directly Impact revenue and hence operating philosophy. The mineral engineer needs to develop an optimum operating strategy to meet his overall objectives, One would expect that, In the modem era of complex computer models, it would be easy to develop an economic model to maximize profits and/or cash flow. Discussions with mill managers Indicate that one mine/corporation's optimum operating strategy is not necessarily optimum for another. This paper approaches the problems from basics "Economics 101" and Illustrates the complexity of developing a universal model which can be applicable to all mines. Case examples are used to illustrate the need for better communication to avoid the basic economic fallacy “that optimization of segments leads to optimization of the whole system.” TECHNICAL MODELS Technical models are a prerequisite to the development of economic models because they provide the necessary Input data to these models, The flow of material from a mine through the sale of the finished product is shown In Figure 1. The ore Is mined and milled, and the concentrate can be stored for sale or smelted and refined In-house or at a toll facility. The finished product(s) are held In Inventory until sold either through existing contracts or in the spot market. [ ] Fairly reliable technical models have been developed for mining, milling and smelter operations. Mine plans are developed to provide a good Indication of the relationship between production and cut-off grade. It is easier to change surface mine plans than plans for underground mines. However, there are limits to maximum production that can be achieved with existing equipment. It Is possible that mine benches may not be developed to handle additional equipment, even If It was purchased to enhance production. More often than not, the cut-off grade Is lowered to Increase throughput goals.

Jan 1, 1993