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By Roe-Hoan Yoon

Froth flotation is widely used for separating particulate materials from one another in aqueous media. The basis of this important separation process is to control the hydrophobicity of particles. However, it is difficult to predict the kinetics of flotation based on contact angle (?), which is a thermodynamic property. A solution to this difficulty would be to develop a relationship between ? and the gradient of the hydrophobic forces across the solid-liquid and bubble-liquid interfaces. Such information can be obtained using the surface force apparatus (SFA), the atomic force microscope (AFM), and the thin film balance (TFB). The present communication reviews the results of the measurements conducted using these methods, and offers a methodology for deriving a rate equation for bubble-particle interactions

Jan 1, 1999

By Li Honggui, Liu Maosheng, Zhao Zhongwei, Sun Peimci, Li Yunjiao

Leaching of Shizhuyuan low-grade refractory tungsten ore was carried out in a mechanically activating reactor. The effect of experimental parameters-temperature, amount of Na2C03, addition of NaOH, ratio of liquid to solid and leaching time - on the recovery of WO3 was studied. The WO3 recovery up to 96.99% have been achieved when the tungsten ore was digested by soda not more than three times the stoichiometric amount at 185-195°C for 1 .O-2.0 hours. Iso-extraction countour diagrams were drawn to illustrate the effect of various parameters on leaching. The WO3 recovery could be expected to increase 10-15% compared to that of conventional soda autoclave process.

Jan 1, 1995

By G. C. Adams, J. H. Fowler

Canada's gypsum industry began more than 200 years ago in the Maritime provinces when farmers attempted to supplement their incomes by selling "land plaster" to traders bound for the United States. Rising from such humble beginnings, gypsum now contributes significantly to the economies of nine of Canada's 10 provinces. Thirteen open-pit and three underground mines produced a total of 8.4 Mt (9.3 million st) of crude gypsum, and 17 processing plants manufactured 2.5 Mt (2.8 million st) of finished products in 1985. Recent economic growth in North America has resulted in steady increases in gypsum production. However, competition from foreign producers and by-product gypsum pose a growing threat to Canada's gypsum mining industry.

Jan 1, 1987

By L. W. Pommier

In the field of minerals processing the concentration of valuable minerals normally takes place in suspensions of ore particles in liquids. The fluid resistance offered by Newtonian liquids to the movement of free settling spheres for the limiting cases of creep flow and turbulent flow is well described by Stokes' and Newton's Laws. However, these equations are not applicable to non-Newtonian fluids, since the relationship between applied shearing stress and resulting rate of shear is a non-linear function, and more than one viscosity parameter is necessary to describe their rheological properties. For this reason, the term "apparent viscosity," which has been commonly used to explain the flow resistance of the non-Newtonian systems including suspensions, may be misleading. A rigorous analysis of the mechanisms of momentum transport of solid-liquid systems is complex indeed. There are many variables that affect the properties of such systems, and precise limitations and assumptions are necessary in order to produce quantitative relationships for predicting their behavior. Very dilute suspensions of spherical particles in Newtonian liquids of identical density as the solids have been examined in the past (12); however, the study of suspensions in which there is a considerable difference in density, between the liquid and the suspended solids has been neglected.

Jan 1, 1968

By K. Mulukhov

Introduction The depth of open-pit mines continues to increase. The typical designed depth for large quarries is now in the 300 to 600 m (1,000 to 2,000 ft) range, with some quarries even reaching depths of 700 to 800 m (2,300 to 2,600 ft). The cost of handling material from such deep mines can be 60% to 70% of the total cost. The capacity of using a cyclic means of transport, such as trucks or skip hoists, continues to drop as the depth of the mine increases. In addition to the problem of maintaining relatively low slope angles for haul roads, increases in environment pollution and increased fuel consumption make the use of trucks ineffective below 100 to 150 m (300 to 500 ft). In handling large quantities of bulky materials, i.e., more than 15 to 20 Mt/a (16 to 22 million stpy), under fixed terminal points, belt conveyors are preferable in almost in all cases. An exception is when the lump size of the load is too high for the conveyor to handle. Belt conveyors in mines can handle loads having a maximum lump size of no more than 250 to 300 mm (10 to 12 in.). The use of a movable crusher prior to the belt conveyor in an open-pit mine would make the process very expensive. Such an operation could not compete with truck haulage on a cost basis. Belt-car conveyor The main factor that restricts the lump size for conventional belt conveyors is collision impacts between the load-carrying belt and idlers. These impacts are completely eliminated in the belt-car conveyor, which was previously described by Mulukhov (1977). The unique feature of the conveyor is that the load-carrying belt is supported by moving cars. Each car consists of a troughed cross strap that supports the belt with the load and two rollers, on which the cars move along upper and lower rails. Cars are connected by two endless chains revolving over sprockets. The system uses conventional drive pulleys, tail pulleys and return idlers. There are no fasteners between the carrying belt and the car cross straps. The conveyor belt simply drives the cars along by the force of friction between the belt and the straps. A belt-car conveyor was first installed in 1970 at the Karatau mining operation in Kazakhstan. The system was an elevating conveyor with an inclination of 20° and was located on the final part of the truck haul road. The conveyor transferred blasting phosphate ore from the quarry to the crushing plant. A bunker was mounted between the trucks and the conveyor. The maximum lump size was typically 1,200 mm (48 in.). However, there were some lumps as large as 1,500mm (60 in.), which was greater than even the belt width (Fig. 1). Because the prototype belt-car conveyor was intended only for industrial trails, it had a relatively short length of about 50 m (160 ft). Nevertheless, the conveyor yielded a substantial profit over the cost of truck haulage. Naturally, the first question one might ask is why such a system is not now widely used. The reason was the further development of a huge transporting complex for the whole quarry at the base of the belt-car conveyor. After successful industrial trials from 1971 through 1973, which demonstrated the systems engineering feasibility, the State Planning Committee of the USSR included the development of the belt-car conveyor in its plan for the development of the national economy. However, mistakes were made. First, a single organization was not in charge of the development. Second, a plant that was not specialized in belt conveyors was put in charge of the design and manufacturing of the conveyor system. In addition, the crusher was installed before the stacker, which negated the major advantages of the belt-car conveyor. For loading the conveyor, a special blade feeder with pendulum suspension was developed, but it was found unreliable. The main conclusion was that, in-

Jan 1, 2003

By Rolf Pollak, Jan Tegtrneier, Gerhard Keller, Horst Gerhardt, Wolfgang Müller, Dieter Tetzner

INTRODUCTION Under certain geological and mining conditions mining near the surface results in an intensive destruction of the overlying rock strata. If the system of mine openings remains filled with air after having completed all mining activities because the flooding of the mines depending on the morphology of the mining area does not exceed a certain level, this circumstance might result in dangers for the public. This is particularly true where uncontrollable air- ways occur within the destroyed rock or if former uranium mining areas are concerned. In these areas radon emanation resulting in undue radon concentration should be taken into account. This circumstance can be counteracted by suitable measures of ventilation. In the following the example of a uranium deposit in the Erzgebirge Mountains (Saxon Ore Mountains) is used to illustrate how statements about the ground flow through overlying rock can be made on the basis of systematic investigations. Finally comparisons with experience gained in coal mining yield orientation data which are used as input quantities in ventilation network calculations and in project preparation of mine fans. Last not least it is proved that endangering the population by radon can be excluded by installing an appropriate fan. INITIAL SITUATION From 1946 to 1989 the uranium deposit of Schlemal-Alberode was mined starting from superficial zones down to depths of 1,800 m. The total output amounted to 80,000 t of uranium, i.e. Schlemal Alberode is one of the largest uranium deposits in the world. The mine openings comprise a cavity of 40 million m3. Al- though thin seams (0.2 to 0.3 m) and steep to semi-steep veins (60 - 70 degrees on average) were extracted using the method of overhand cut-and-fill there were also deposit sections with larger thicknesses and poor ore zones, respectively, where especially after the war until about 1955 large openings were formed due to missing filling. Due to this uncontrolled, relatively "wild" mining considerable subsidence of the surface and fractures in the border zone of the subsidence area occurred. The town of Oberschlema was destroyed and an area of 0.14 km2 was damaged by mining involving a maximum subsidence of 6 m in the centre - called deformation area. The rehabilitation work for the deposit area of Schlemal Alberode involves the flooding of the openings up to a level of the Markus-Semmler-Sohle as long-term preventive measure. Fig. 1 shows the flooding level of October 1996. The level of the -540 m drift will probably be reached in June 1997. From that point of time a general rearrangement of the ventilation has to be made because the central return air ventilation via shaft 373 will no longer be possible. This shaft has no connection with the superficia1 system of mine openings. A new return air ventilation system based on shaft 382 as return air shaft will be designed (Fig. 1). The capacity of the future mine fan system will e.g. be determined by the flow conditions in the deformation area mentioned above. For this reason investigations were carried out in a test drift 13a on the level of the Markus-Semmler-Sohle in the stretching zone of the subsidence trough for several days (Fig. 2 and 3). INFLUENCE OF THE OPERATING CONDITIONS OF THE MAIN MINE FAN ON THE RADIATION SITUATION ON SURFACE The beginning of flooding in 1991 was also the beginning of the investigations into the options of future ventilation due to the reduction of the mining area. The first analyses of radon emanation made in May 1991 showed that only about 20 % of the total radon emanation result from systematically ventilated mine workings. This means that despite of existing dams the air streams through the majority of mine openings of the deposit. The analysis of the individual radon inflows illustrates the special problematic nature of the deposit involving all possible effects on the surface. The shutdown of the main mine fan over the long period between 2.10. and 6.10.1991 showed the essential influence of the operating conditions of the fans on the whole radiation situation on surface. Fig. 4 shows the comparison of the radon concentration level and its daughters during a 'normal" weekend with the level during long-term shutdown in the old 'Casino" of the papermill of Niederschlema. This papermill as well as the 'Casino" are situated in the outcropping zone of the deposit section of Niederschlema. The -60 m level was driven as the upper level in this area, from which the panels were driven to about 20 m below the surface. Also below the "Casino" there is such a panel in the vein 'Glirck", whose roof is directly situated below the building in a depth of 39 m. Directly next to it there is the raise 2 of the panel 'Namenlos" in a depth of 27 m. The graphics clearly shows the immediate increase in radon concentration as well as in the daughter concentration after the shutdown of the fans. The elimination of the artificially generated negative pressure results in the reversal of the convection direction due to the natural ventilating pressure. The interface structure allows the mine airto pass through aero- dynamic connections to the surface. Neglecting the meteorological parameters the concentrations of radon and its daughters in- creased by about 350 % as compared to a 'normal" weekend during longterm shutdown (i.e. in 4 days). Furthermore it could be noticed that a stationary state between the radon and its daughters has not been established in the period

Jan 1, 1997

By George R. Powe

If Campbell County, Wyoming were an independent nation, it would rank fifth in the world among countries with known coal reserves. Only the United States, Russia, mainland China and West Germany would have greater volumes available. Not too many years ago that statement would not have seemed important, because oil and natural gas had become our nation's principal fuels. Western coal deposits, particularly those of the Powder River Basin in Montana and Wyoming, and which underlie virtually all of Campbell County, were largely unknown. But recently, in response to the three-pronged stimulus of vanishing gas reserves, rising petroleum prices, and clean environment statutes, these vast deposits have been delineated and have become, in the minds of many, indispensable to our goal of energy independence. Not only does western coal lie in thick seams near the surface where it can be mined in huge quantities at low cost, but also it possesses another coveted attribute in this age of environmentalism: low sulfur content. At Burlington Northern we count ourselves at the forefront of those who believe this coal is essential to any hope for the government's "Project Independence." But to that belief we add another article of faith. We are convinced the economics of unit train delivery of this incredibly immense resource make it the only transportation system that can help us achieve energy self-sufficiency in this country.

Jan 1, 1975

By R. C. Emmett

Gravity sedimentation is one of the most widely used unit operations in mineral processing and hydrometallurgy. Properly selected and operated sedimentation equipment can provide probably the most economical means of concentrating solids or clarifying solutions. Because of the variety of different approaches possible, including the use of modern flocculants, it is easy to overlook the optimum approach in selecting equipment. The different types of sedimentation systems that can be used in the minerals field and the sizing methods that are currently employed will be discussed here. Sedimentation applications usually fall into two areas, clarification or thickening. While the two operations are basically similar, the difference in emphasis (a well-clarified overflow as opposed to a highly concentrated underflow) results in significant differences in testing, equipment design, and operation. Clarifiers may be merely simple settling vessels with no special features, other than ample volume to provide the detention time needed for suspended solids to settle out of the feed slurry. Or, they can be designed to employ elaborate solids-contact techniques utilizing coagulants or flocculants to achieve more rapid clarification through the growth of particle agglomerates. Thickeners also are available in a wide variety of designs, with the main differences concentrated in the design of the raking mechanism. Thickeners operating in series (continuous countercurrent decantation or CCD) are used to recover valuable solution or purify the insoluble solids. Tray thickeners are used for recovering fine particulate solids such as alumina seed, but in other applications have been supplanted largely by unit thickeners or inclined plate units. Thickeners and clarifiers sometimes are used for storage of either solids or overflow liquor, or employed as precipitating reactors, such as in desilication applications. The use of flocculants and coagulants is much more common in modern processing plants, primarily due to the reduction in capital costs rendered possible through the use of these reagents. The terms flocculant and coagulant are often used interchangeably; generally, it is understood that flocculants refer to organic, high molecular weight synthetic or natural polymers which are useful in flocculating most suspended solids, but generally are ineffective on colloidal suspensions. Coagulants refer to chemicals such as those commonly used in water treatment, alum, ferric salts, lime, etc., which are usually far more effective on colloidal solids but cannot achieve the degree of agglomeration that is possible with the polymers. In some cases, both coagulants and flocculants are used in order to achieve optimum results.

Jan 1, 1986

By A. E. Koski

This paper describes the implementation of specialized software in mine planning at the Cleveland-Cliffs' managed Empire Mine on Michigan's Marquette Iron Range. The application puts together years of mining experience on the eastern end of the Marquette Range beginning in the early 1960's. The Empire Mine, in its thirty-fifth year of production, uses the software to model a complex and highly variable low grade magnetite ore body, capturing the economics of mining and processing to the final product. With special attention given to geostatistical methods, the software is used for all geologic modeling, ore reserve estimation, ore control, pit design and short and long range planning.

Jan 1, 1999

By John O. Marsden

The project development metallurgist is faced with an interesting dilemma: The use of overly conservative metallurgical projections and cost estimates can kill a project that has the potential to provide needed jobs and growth to a company's bottom line. On the other hand if the metallurgist uses a highly aggressive approach by trimming capital and operating cost estimates to the bone and by applying highly optimistic metallurgical projections, there is a risk of creating a project that can never live up to expectations. The metallurgist must walk a tightrope between "sandbagging" and "overoptimism," and this tightrope is fraught with political intrigue and difficulty. This paper takes a some- what lighthearted look at some of the political aspects associated with this issue, and other issues, and offers a few guidelines for walking the "political metallurgical" tightrope.

Jan 1, 2001

By R. V. Ramani, K. V. K. Prasad, M. Swaminathan

Engineering design involves, in addition to numerical calculations, considerable amount of expertise and judgement. The expertise and judgement is required to conceptualize and idealize the problem, select a suitable mathematical model for the problem and to interpret the model results in terms of stipulated design constraints. In recent years, knowledge based systems (KBS) have provided the capability to integrate the qualitative experiential and expert knowledge with the quantitative knowledge for improved analysis and design of systems. A possible KBS architecture for mine ventilation system design was discussed by Ramani and Prasad (1987). The expert system (ES) component of the KBS supports the analysis and interpretation phases of the ventilation design process with an algorithmic program used for the solution of alternate designs. The purpose of this paper is to discuss the data and knowledge needs for developing such a system and a formal means of their representation on the computer. A number of ES shells were evaluated to explore possible development and delivery environments for the system. A brief summary of this evaluation is also provided.

Jan 1, 1989

By Edward C. Spalding

The mineral estate of Arizona's State Trust lands, which are administered by the State Land Department for the benefit of the common schools and 14 other beneficiaries, totals more than 10.5 million acres, or almost 15 percent of the State's entire area. Trust lands exist in all 15 of the State's counties, with the most and least acreage found in Yavapai and Gila Counties, respectively. Revenues derived from the Trust mineral estate include royalties, rentals, and fees. In Fiscal Year 1983-84, these revenues amounted to $4.2 million, or 13.5 percent of all income received by the Trust during this period. Industrial minerals, led by sand and gravel, contributed $1.1 million from 82 active properties.

Jan 1, 1987

By Terry O. Arney

Ore Reserve Models are well established evaluation and mine planning tools. BY definition, these models are made up of regularly sized blocks, arranged in a three dimensional matrix. Two dimensions correspond to north and east directions, and the third, to the depth or elevation. The dimensions of the blocks need not be equal, but generally at least two of the dimensions are equal. The blocks are located in space by coordinates which are generally translated into index values for computer applications. Each block will have quantitative attributes, such as grades, densities, rock characteristics, etc. assigned to it. The process of assigning values to each block in the three dimensional matrix is the process of ore reserve model development. One of several interpolation procedures may be employed to predict the value of a block, based on surrounding data. Block value models have several uses. The utilization of block models in ore reserve computations is quite common. Here, ore reserves may be calculated by level or by cross-section for the entire deposit. For each ore reserve calculation, multiple cut-off grades may be analyzed to determine the ore reserves under different cut-off conditions. The other major use of block value models is in mine simulation. Through simulation, profitable studies that analyze the effect of alternate ore removal sequences can be undertaken to determine economic optimality. Other studies that are undertaken in mine simulation problems include the evaluation of mine layout, pit slope effects, and the impact of varying cut-off grades.

Jan 1, 1974

By Tim Q. Haldane

The ability to efficiently process low grade gold ores has allowed the Gold Fields Mesquite Mine in Imperial County, California to expand from a 2.1 million tonne per year rate in 1986 to its present 7.3 million tonne per year processing rate. Increased tonnage demands led to the 1989 construction and start-up of a dump leach pad to supplement the existing heap leach operation, resulting in an excellent opportunity to compare the two processes at the same operation. Additional operating costs associated with the crushed ore heaps are justified through enhanced recoveries, even at very low ore grades. An overview of current operating practices at Mesquite is presented along with metallurgical and economic comparisons between the crushed and run-of-mine ore heaps.

Jan 1, 1990

By Keith G. Wallace, Gary K. Rogers

As a supplement to an analysis involving methodologies for the estimation of airflow resistance in shafts used for both ventilation and hoisting, a detailed shaft characterization survey was distributed worldwide. The survey provided detailed information which enabled verification of recommended velocities and established limitations on fill factors and relative velocities. This paper evaluates the information obtained for 37 shafts and presents a statistical evaluation of the results. Apart from airflow rates, conveyance rates, and physical characteristics, qualitative information regarding means of conveyance stability and velocity control techniques at unloading and loading points are also presented. Calculations of fill coefficient, free area velocity, and most importantly, relative velocity between the airflow and hoist, are conducted. These results are presented by frequency diagrams which illustrate the trends, as well as provide limitations for these parameters. The data illustrates that, commonly, the fill coefficient for two or more conveyances, does not exceed 30 %. A limit of 50% is suggested for a single conveyance shaft. A recommended free area velocity of 10 m/s (1,970 ft/min) for combined ventilation and hoisting shafts, is sup- ported by these results. Relative velocities of 30 m/s (5,905 ft/min) are most frequent and 50 m/s (9,840 ft/min) being a practical maximum.

Jan 1, 1987

By Dao De Xu, Heechan Cho, Mark S. Klima

A study was conducted to determine the feasibility of using dense-medium centrifugation for separating fine (-150 µm) coal. A centrifuge provides the capability of varying the rotational speed and retention time independently, which offers the potential for a higher degree of control as compared to a dense-medium cyclone. Tests were conducted with a 140-mm diameter by 76-mm long Plexiglas separator, which was fitted with appropriate feed and discharge pipes to allow continuous separations to be made. A suspension of ultrafine (-10 µm) magnetite and water was used as the dense medium. The effects of operating conditions, including the magnitude of the 'g"-force and retention time, on centrifuge performance were investigated.

Jan 1, 1995

By C. I. Wilmot

FMC's Paradise Peak Mine was discovered in July 1983. The deposit at discovery contained 11,340 ,000 tonnes (12,500,000 tons) of ore grading 3.71 glt (0.108 oz/ton) Au and 120 glt (3.5 oz/ton) Ag. A feasibility study was completed by Davy McKee in May 1984, with detailed engineering starting in June 1984. Construction of the facility started on January 2, 1985 with plant start-up taking place on April 14, 1986, two and one half months ahead of schedule and more than 10% under budget. This paper describes the status of the operation approximately 1-1/2 years after start-up including revisions and additions to equipment and process.

Jan 1, 1988

By J. DeYoung

Weak demand for metals and consequent low metal prices resulted in continuation of the depressed state of US mineral exploration activity through 1983. Data collected by Resources for the Future Inc. and The Pennsylvania State University for a cooperative study on mineral exploration activity indicate that US exploration in 1983 fell by as much as 25% from 1982 levels. Precious metals remained the focus of most exploration efforts, as gold and silver prices remained attractive compared to base metals. Reports of 1983 domestic exploration and development activities in several industry journals indicates that about 85% to 90% of those projects were directed primarily at gold and silver. Gold exploration included work by several companies on Carlin- type disseminated gold deposits in Nevada, geological prospecting and drilling in the Wenatchee Gold Belt in Washington, and interest in lode and placer gold properties in California, especially by several small Canadian firms. Precious metals exploration projects were also reported in Alaska, Arizona, Colorado, Montana, New Mexico, Oregon, South Dakota, and Utah. There was also tome precious metal exploration activity in the eastern US. Development continued at the Ropes gold mine near Ishpeming, MI, and companies explored historic gold-producing areas in the Southeast. Some exploration for nonprecious metals took place last year. Stratabound copper-silver deposits were examined in western Montana, and US Borax continued exploring molybdenum resources in the Misty Fjords area in Alaska. Work on these projects during 1983 may have been stimulated by time limits on the availability of certain federal lands for exploration more than by projected improvements in metal markets. During 1983, several papers were published on mineral exploration strategies and the future of US exploration. David Watson of Raymond Kaiser Engineers, in a paper presented at the 1983 AMC Mining Convention, forecast increased exploration activity for gold and silver because of world economic conditions. He added that antimony, manganese, chromium, graphite, mercury, tantalum, tin, and barite are also likely targets based on their limited domestic reserve availability. He cited concern over reported subsidized mineral production by foreign producers and US regulatory and land-planning policies as retarding mineral exploration activity. Douglas Cook of Freeport Exploration Co., at the same meeting, presented a paper that showed that a 1950-75 period of vigorous exploration has been followed by a period where acquisition has replaced exploration as the predominant means of acquiring new mineral properties. At a December IIASA conference on "The Economics of Mineral Exploration," in Vienna, Austria, A. Rose and R. Eggert of The Pennsylvania State University pointed out that recent exploration decisions tend to rely more on evaluating production costs and comparing them with existing deposits, rather than on return of

Jan 5, 1984

By Frederick T. Graybeal

Silver Bell is located about 56 km (35 miles) northwest of Tucson, AZ. It is a classic porphyry copper deposit containing enriched disseminated chalcocite ore in altered porphyritic rocks with chalcopyrite in adjacent skarn zones. The discovery of Silver Bell was an evolutionary process that occurred over a 75-year period. Recorded mining activity began about 1865 on Boot Hill east of El Tiro pit (see Fig. 1) where outcrops of altered limestones were found to contain 425 to 482 g (15 to 17 oz) Ag mixed with copper oxides. By 1874, the Young America Copper Co. was mining shallow occurrences of 30% Cu along the north side of the present Oxide pit. The ore was shipped by wagon to Yuma, AZ, and then by boat to San Francisco for smelting. By 1881, the district supported several mining companies, two small smelters, and was being favorably compared to Bisbee in local newspapers. Most of the ore mined in the period 1865-1903 came from pods of massive sulfide or their oxidized products which were distributed in an unpredictable fashion through a large skarn zone in Paleozoic limestone east of El Tiro pit. Ore grades varied from 5 to 10% Cu.

Jan 1, 1991

By O. R. Swendseid

The Mountain Con mine is a relatively old, deep mine located in Butte, Montana. Operations at this property through the present shaft were started before the turn of the century. As mining progressed to depth, increasing demand for ventilation facilities and changing mining conditions made it convenient to convert some nearby shafts to airways. The fact that these openings had not been originally planned for use as airways made it necessary in some instances to utilize long and inefficient already-estab¬lished entries to service working areas. This situation existed in the Lexington fresh air system. The Lexington system consists of a vertical, 3-compartment shaft extending from surface to the 3200 level and a series of 2-compartment air raises and old horizontal-haulage openings extending into a mining area several thousand feet west of the Lexington shaft between the 3900 level and the 4200 level. Because of restrictions and high resistance in the portion of the system between the bottom of the Lexington shaft on the 3200 and the mining area on the 3900, the capacity of this system is only 85,000 cubic feet of air per minute. It was calculated that the capacity of the system could be in¬creased to meet the demands of the current mining program in this section of the mine by utilizing a single, straight, horizontal air course on the 3200 and by driving an inclined air raise 740 feet long between the 3900 and the 3200.

Jan 1, 1967