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By W. S. Row

This mine is operated at 4000 to 4500 tons daily through a single shaft, with one rock hoist and one senice hoist. Latest shaft construction is concrete with wooden dividers. Economics of drifters and air-leg drills are compared, as well as costs of shrinkage and sub-level stoping. Economy of large haulage equipment and underground crushing is shown. THE property of Kerr-Addison Gold Mines is located in the Larder Lake district of Ontario about 3 miles from the Quebec boundary on the highway between Kirkland Lake and Noranda. It is roughly 320 miles due north of Toronto and 365 miles north of Buffalo, N. Y. The property was staked in 1906 and after several development efforts finally came into continuous production at 500 tons per day in 1937. In 1939 the daily tonnage -was increased to 1200 and in 1941 to 2100 tons. At the end of the war, because of the shortage of labor, tonnage was down to 1100 tons per day, rising to 2200 in February 1947. A further plant expansion inaugurated in 1946 was completed by December 1948, bringing the capacity of the mine and mill to a production rate of 4000 tons per day. Current production, with 880 men employed, is 4400 tons per day. Changes in mining methods and equipment, required by the recently completed expansion program, have yielded interesting comparative results, some of which are described in this paper. Hoisting of ore and waste from the mine, as well as all service facilities for handling men and supplies underground, are provided for in a single vertical five-compartment shaft. Ore and waste are hoisted in 12½-ton skips in balance by a 14-ft diam hoist with drums 115 in. wide, grooved for 2?-in. wire ropes. This hoist, with a total rope pull of 85,000 lb, was designed to operate at a speed of 2600 ft per min to a depth of 4000 ft, hoisting 15 tons of ore in each skip. It is amplidyne-controlled and powered with dual drive dc motors of 2250-hp each, through flexible couplings to two pinion shafts, one on each side of the main gear. Each motor has a peak load rating of 4500-hp at point of greatest duty. Direct current power to the hoist motors is supplied from a motor generator set consisting of a 5000-hp synchronous motor driving two 1750-kw dc generators. The loading pockets used so far are at depths of 1500 and 2725 ft respectively. Consequently the present hoisting speed is 2000 ft per min and the skips capacity only 12½ tons. Now that the shaft has been deepened the new pockets at 3934 ft will be increasingly used so that the speed and the skip size will be increased up to original specifications as required. Some difficulty in meeting the required hoisting cycle is expected because of the shape of the skips. As now in use the cross-sectional dimensions of the inside of the skip are 4 ft 6 in. x 3 ft 8 in. To hold 121/2 tons, therefore, they are 14 ft 10 in. deep. Since the skips are the conventional overturning or Kim-berley type, considerable hoisting time is lost in retardation in entering the headframe dump and in acceleration when leaving. It is expected that to obtain the specified tonnage from this hoist when the skip size is increased to 15 tons it may be necessary to resort to bottom dump skips. This type of skip could reduce considerably the present excessive acceleration and retardation times. Analyses of the skip-hoisting operation are made monthly, and every effort is made to reduce controllable delays. For a nine-month period in 1951 such an analysis shows that the hoist was in operation 92.5 pct of the available time, which was allotted as shown in Table I. Overall hoisting costs, cage and skip, for the year 1951, averaged $0.168 per ton of ore milled. The mine is serviced entirely by means of a single cage, capacity 35 men or 15,700 lb, operating in con-

Jan 1, 1954

By C. Kremer Bain

DURING the past several years of diamond drilling in Washington County, Mo., the St. Joseph Lead Co. has discovered a concentration of commercial lead-zinc ore at four different points within an area of 36 square miles. The largest and to date the most promising group of these orebodies is located in the Indian Creek area. It was considered to be of sufficient value to warrant development. The Indian Creek area is in the center of the northern half of Washington County, 11 miles northwest of Potosi, Mo., the county seat. The location is readily accessible from Potosi by car or truck over the all-weather Missouri State Highway No. 114. The topography is one of rolling relief, varying from an elevation of 700 ft above sea level on Mineral Fork and Indian Creek to a maximum of 1176 ft at High Point Tower. It is divided in equal halves between the drainage to Mineral Fork and to Indian Creek. The shaft site at 1060 elevation is on the divide between the two drainage areas. The Indian Creek side is rough and heavily wooded, with only a few cultivated clearings. The slopes to Mineral Fork are more gentle and rounded, with much of the area cleared for farm land. Most of the surface is covered with Potosi and Eminence formations, which show little evidence of any structure. Ore structure was worked out by extensive diamond core drilling, with holes varying in depth from about 700 to 1300 ft. As of Oct. 1, 1950, 218 prospect holes had been drilled within the 36-square mile area. The ore is disseminated galena with considerable accompanying sphalerite in the Bonne Terre forma-tion in the shape of irregular blanket-type deposits. The main controlling structural feature of the outlined mineral zone is a buried pre-Cambrian rhy-olite-porphyry ridge extending partly through the Bonne Terre formation, which laps against the por phyry and also covers it. The mineralized zone parallels and forms an irregular halo around the porphyry ridge and is concentrated in the Bonne Terre formation where the La Motte sandstone is cut out by the porphyry ridge. This is a very common mineral occurrence in the southeastern Missouri lead belt district. The La Motte sandstone, which normally underlies the Bonne Terre formation, is thus cut out entirely. In this area the Bonne Terre is usually about 275 to 320 ft thick, the La Motte varying in thickness from zero up to 400 ft. Commercial concentrations of ore are indicated at several places, the most notable being at the shaft site where an orebody at least 4000 ft long with widths of 500 to 600 ft has been outlined. The total thickness of mineralized rock reaches 150 ft in some holes, although total ore thickness runs from 8 ft to about 60 ft, often in more than one layer. Preliminary Planning Ore reserves as outlined lend themselves to open-stope mining with trackless haulage and mobile loading machines and drilling equipment. This type of equipment has proved to be very efficient during recent years in the company's lead belt operations. After a study of bids for shaft sinking and an estimation of comparative costs, it was decided that the shaft would be sunk by St. Joseph Lead Co. personnel using company equipment. Since the plant site is about 26 miles from the company's nearest lead belt operations, it was necessary to plan a complete surface plant, including mill, shops, office, change room, and warehousing facilities. Immediately following the filing of an application for a certificate of necessity, a headframe was selected for which detailed drawings were already available. The order was placed for fabrication so that it could be delivered and erected in time to serve for sinking the shaft. This eliminated the expense of a temporary headframe. Drilling was started on two churn drill holes near shaft location, and two 1000-gpm deep well pumps

Jan 1, 1954

By J. J. Lillie

Sublevel stoping was first developed in the Michigan iron mines many years ago. Since that time this method, and modifications with long hole drilling, have been used in a number of non-ferrous mines and have been described in various papers and articles. With a few exceptions, the operations where sublevel methods have been applied are producers of large daily tonnages from ore bodies of considerable size. In selecting mining methods for small properties or relatively small detached ore bodies, sublevel methods often have been overlooked. The method has been used successfully in relatively small operations at Copper Canyon, Nevada, and, more recently, in mining small ore bodies at Darwin, Calif. As a result of this experience, we believe that a review of the general procedure, the advantages noted, and the results obtained will be of interest to many operators of relatively small mines. General Description Briefly, sublevel stoping consists of, first, driving sublevel drifts at predetermined vertical intervals through the length of the stope block; second, breaking an initial mining slot across the stope; third, driving crosscut benches from each sublevel by slabbing into the slot; and, fourth, breaking the ground between benches. Applicability Since this is an open stope method, it requires wall rock that will stand over reasonable stope lengths without caving or excessive slabbing. Ore must be moderately firm but need not be hard. The method is best adapted to steep dipping, tabular or lenticular ore bodies but may be applied to ore bodies of almost any shape if their vertical extent is appreciable. It can be used with dips that are too flat for successful shrinkage stopes and can be modified for use with slushers on dips below about 40". Initial mining at the Copper Canyon operation consisted of sublevel stopes for the wider ore sections and shrinkage stopes for the narrow sections where ore widths averaged about 15 ft. After some experience and a study of comparative results on completed stope blocks, sublevels were substituted for shrinkage stopes even in the narrow ore widths. Practical stope lengths depend upon the character of the walls and the continuity of values within the ore body. Where blocks of low grade or waste are included in the stoping limits, these areas can readily be left as pillars for intermediate wall support and longer stope lengths may be planned. Raises can be spotted in narrow or low grade areas and stope lengths adjusted accordingly. Stoping Details To illustrate stoping details, assume that we have an ore shoot 15 to 30 ft wide and about 200 ft long, with walls that would stand over a stope length of 80 to 90 ft (Fig1). PRELIMINARY DEVELOPMENT Preliminary development for sublevel stoping would be simple. A service raise is driven at about the center of the block and stope limits planned to leave a pillar around this raise. Sublevel drifts driven each way from this central raise will outline the

Jan 1, 1950

By E. M. Norris

Phosphate deposits are distributed widely over the earth's surface. Of the known areas of deposit, eight fields are of particular interest because of their vast reserves of high grade phosphatic material and the volume of their production that goes to the world's great consuming areas. These fields, in the order of their present economic importance, are situated respectively in Florida, French Morocco, Kola Peninsula of U.S.S.R., Tunisia and Algeria, Tennessee, Mon- tana-Idaho-Wyoming-Utah, Ocean and Naru Islands in the Gilbert group, and in Egypt. The Rocky Mountain deposits have a comparatively low ranking in the phosphate trade of today, but this field contains one of the largest known reserves of phosphatic material, associated with which are recoverable trace minerals, which are of increasing importance in our national economy. The first discoveries of phosphate rock in the western states were recorded in Cache1 and Rich² counties in Utah, respectively in 1889 and 1897. Mining operations were begun on these deposits near Montpelier, Idaho, in 1906. During the latter year geological studies of the phosphate beds were started by Weeks and Ferrier of the U. S. Geological Survey. These studies have been continued to the present day by a distinguished roster of the Survey's technicians. As a result of this systematic exploratory program, the Survey has published a series of bulletins, papers, and maps, which provide the mining public with a comprehensive study of the exposed

Jan 1, 1950

By Rudolph G. Wuerker

MUCH descriptive matter has appeared on the subject of suspension roof supports, or roof bolting, as it is more commonly called. The widespread introduction of roof bolting into coal mines and metal mines is truly phenomenal. Mine operators were quick to recognize the advantages of supporting wide openings without hindrance to machine maneuverability and ventilation. Although suspension roof support has long been installed at St. Joseph Lead Co. mines in southeast Missouri,'" its application to coal mining presented new problems, such as proper anchorage and bearing for the bolts, bolt diameter, and spacing of bolts. After continuous testing and experimenting at the mines, standard roof-bolting materials were determined.'!' The study reported in this paper is not concerned with such details as bolt diameter, which may be considered already solved in practice. In the tests discussed here, small models patterned on actual bolts were found to function in the same way and as satisfactorily as their prototypes. The aim of these tests was rather to investigate the influence of roof-bolting systems on the stress distribution around mine openings and to study the fracture patterns obtained in actual testing. Little was found about this in the literature, as testing of suspension roof methods and quantitative measurements are only now coming to the fore. Several suggestions and actual measurements have been made to evaluate critically the functioning of roof bolting systems, single roof bolts, and parts thereof. Outstanding among them is Bucky's outline of structural model tests.'" Since none of the suggested testing equipment was available, however, for the experiments discussed below, a different approach was chosen. The response of a mine roof under stress has often been compared to that of a beam. The slow coming down and bending through of beam or plate-like banks of shale, sandstone, or top coal is a familiar occurrence, extensively cited in the literature." It was felt that testing of roof-bolt systems installed in a concrete beam which was loaded in bending would be a fair approximation of the behavior of a mine roof underground. Another school of thought considers the roof behavior over an underground opening in connection with the stress distribution all around a circular or rectangular opening. This is more accurate, and leads to the concept of a dome-shaped zone of material destroyed under tensile stress. This is likewise a common sight in unsupported roadways where the continuous fall of roof results in what has been called the natural outline of roof fracture. This theory could not be tested and is treated separately in Appendix B. It is important to note that according to both assumptions the immediate roof fails in tension; the use of a beam in these tests, therefore, should give information valid for either of the two theories. With the testing equipment at hand it was possible to load concrete beams 6xlx0.5 ft under two-point loading, giving an equal bending moment over the center part in which the model bolts were installed. A comparison was made of the ultimate loads needed to break plain beams and beams in which roof bolts were installed. Arrangements were made with: 1—plain beams; 2—bolts with plate washers, some with holes drilled at 90" angles and others with holes drilled at 45" angles; 3—bolts with channel irons underneath; 4—bolts in holes filled afterward with cement; and 5—bolts anchored in a stronger stratum. The foregoing arrangement is made in order of increasing strength, as assumed from the theory of reinforced concrete. Likewise, laminated beams with wooden model bolts and with combinations of the foregoing set-ups were tested. All in all, 21 experiments were made out of the much greater number of combinations possible. There were, too, some trial tests. Enough observations from this limited number were made to interpret the behavior of mine roof, supported by various types of suspension bolts, at fracture. In present-day concepts, which have been proved by mathematical derivations and stress analyses, any opening driven underground will change the distribution and magnitude of the stresses existing around it. It does not matter whether the stresses become visible, as in rocks whose strength is less than the forces acting upon them, or whether they are invisible, as in the gangways lacking evidence of rock pressure. In this latter case the rocks can withstand changes in stress-distribution. To consider the mine roof as a beam, there are, with transversal loading, tensile stresses in the lower fiber and compressive stresses in the upper layers above the neutral axis of the beam. Beams of brittle material such as rock and concrete fail exactly as shown in Fig. 1. Nearly all model beams showed the same fracture pattern as that of a tension crack. The influence of support, by roof bolting or conventional

Jan 1, 1954

By George C. Lipsey

FOR many years the mining of coal and other minerals occurring in sedimentary formation has been planned according to information obtained by diamond drilling. Placer mines have used churn drills for sampling and estimating deposits, and some metal mines have used the method either wholly or in part. The diamond drill method, therefore, although not entirely new, has come into more prominence in the last few years. Several large-scale developments now being planned or in the production stage have used this method in varying degree. This paper will describe the operations of the Howe Sound Exploration Co., Ltd. at Snow Lake, Manitoba, where a gold mine was developed and equipped for mining and milling wholly on the basis of diamond drilling results. The property has been in operation nearly 3 years, so that results can be compared with the estimates and assumptions made from the diamond drilling. In developing the subject it is necessary to outline as briefly as possible the normal method of mining, from discovery to production stage. Table I sets forth these various stages, the methods of financing, the percentage of total cost, and the relative time interval required for each section of the work. The plan presented in this table cannot be applied to all mining developments, but it gives a basis on which to work. In Stages 2 and 3 the investment is moderate, and if the property does not respond favorably the options or interests can be terminated. Stage 4 represents a sizeable investment, and it is more difficult to drop the property after this part of the operation is completed. Stages 2, 3, and 4 may be repeated several times by diflerent groups or companies before Stage 5 is reached, but when Stage 5 has been started the total capital investment must be provided and there can be no turning back. It will be noted that there is a total of 116 units of investment in the overall cost if one company proceeds with Stages 2, 3, 4, and 5 in order. Stage 4, however, consumes time which is important when markets are being considered. Bringing a property into production when metal prices are low is much more difficult than when the prices are high, and if this stage can be set aside production is attained more quickly. When Stage 4 is eliminated the total becomes 86 units of cost for Stages 2, 3, and 5, but an adjustment of 14 units must be added to Stage 5. These 14 units represent the useable value of Stage 4. Thus Stage 4 represents an investment of 30 units if carried on before Stage 5, but 14 units if included with Stage 5. This is accounted for by the fact that development work can be accomplished more cheaply under Stage 5, since everything that is done is planned on a production basis with permanent services. Stage 4 must of necessity be governed by expenditures, and a great many temporary services, including housing and power plants, must be provided. The prospect shaft is a small one, and drifts,

Jan 1, 1953

By George C. Lipsey

FOR many years the mining of coal and other minerals occurring in sedimentary formation has been planned according to information obtained by diamond drilling. Placer mines have used churn drills for sampling and estimating deposits, and some metal mines have used the method either wholly or in part. The diamond drill method, therefore, although not entirely new, has come into more prominence in the last few years. Several large-scale developments now being planned or in the production stage have used this method in varying degree. This paper will describe the operations of the Howe Sound Exploration Co., Ltd. at Snow Lake, Manitoba, where a gold mine was developed and equipped for mining and milling wholly on the basis of diamond drilling results. The property has been in operation nearly 3 years, so that results can be compared with the estimates and assumptions made from the diamond drilling. In developing the subject it is necessary to outline as briefly as possible the normal method of mining, from discovery to production stage. Table I sets forth these various stages, the methods of financing, the percentage of total cost, and the relative time interval required for each section of the work. The plan presented in this table cannot be applied to all mining developments, but it gives a basis on which to work. In Stages 2 and 3 the investment is moderate, and if the property does not respond favorably the options or interests can be terminated. Stage 4 represents a sizeable investment, and it is more difficult to drop the property after this part of the operation is completed. Stages 2, 3, and 4 may be repeated several times by diflerent groups or companies before Stage 5 is reached, but when Stage 5 has been started the total capital investment must be provided and there can be no turning back. It will be noted that there is a total of 116 units of investment in the overall cost if one company proceeds with Stages 2, 3, 4, and 5 in order. Stage 4, however, consumes time which is important when markets are being considered. Bringing a property into production when metal prices are low is much more difficult than when the prices are high, and if this stage can be set aside production is attained more quickly. When Stage 4 is eliminated the total becomes 86 units of cost for Stages 2, 3, and 5, but an adjustment of 14 units must be added to Stage 5. These 14 units represent the useable value of Stage 4. Thus Stage 4 represents an investment of 30 units if carried on before Stage 5, but 14 units if included with Stage 5. This is accounted for by the fact that development work can be accomplished more cheaply under Stage 5, since everything that is done is planned on a production basis with permanent services. Stage 4 must of necessity be governed by expenditures, and a great many temporary services, including housing and power plants, must be provided. The prospect shaft is a small one, and drifts,

Jan 1, 1953

By Be Van Presley

There has been developed at Latouche a rather unusual system of mining which, for want of a better name, has been called the "Latouche system of mining." It is a modified form of shrinkage stoping applicable to orebodies of large or fairly large size where the ground is such that timbering would not be required in stoping. Mention has been made of this method of mining by Birchll who gave no details as it had been in use only a short time and the final outcome undetermined. The system, however, has proved to be entirely successful and it is thought that a paper covering it might be of general interest. Before proceeding with a detailed description of the Latouche system a brief description of the property and the various mining methods previously used will be given. Location The Beatson mine of the Kennecott Copper Corpn. is situated on the west coast of Latouche Island, in Prince William Sound, Southwestern Alaska, 45 miles east of Seward, the coast terminus of the U. S. Government Railroad. It is a regular port of call for ships on the Southwestern Alaska run. The island is about 12 miles long and 3 miles wide. A ridge of mountains runs the entire length with a maximum elevation of 2500 feet. Climate Latouche has to contend with possibly the greatest amount of precipitation on the North American continent. The annual average precipitation for the past 4 years has been 188 in., with 248 in. for the year 1926. A rainfall of 4 in. per day is not uncommon. In the winter the thermometer rarely falls below +20° F. and the snowfall is usually heavy during the months of March and April.

Jan 1, 1928

By Robert Y. Grant

This study outlines the history of metal mining in Japan and the characteristics of the industry as they existed from 1925 to 1945. Mining and milling operations are described together with the role of the Japanese Government. A description of post-hostilities conditions and the present state of the industry is included. MANY of Japan's larger and better known mines date back to the early centuries of the Christian Era. The first stimulus to mining in Japan came through the growth of Buddhism, A. D. 600 to 700, and the attendant need for copper and other metals for temple ornaments and statues. During the Feudal period, 1333 to 1568, mine operations were expanded to meet the demand for ornament metal and for financing of campaigns. After removal of the seat of government from Kyoto to Tokyo, about 1600, the Tokugawa Shogunate placed great emphasis on mining, and many new mines were opened, some of which are still operating. Toward the end of the Tokugawa period, mining had declined sharply from its earlier peak. The easily worked, high grade portions of the orebodies had been exhausted, and the mining techniques then available could not exploit the deeper, lower grade ores. Similarly, extraction of metal from the sulphide ores presented a major problem. With the Meiji Restoration in 1868 and the rapid opening of the country to western ideas came a rejuvenation of the mining industry. Mining, milling, and smelting processes were greatly improved, resulting in the extension of development work into hitherto inaccessible portions of orebodies. The demand for metal during World War I stimulated activity throughout the industry; copper mining, particularly, expanded to a point where Japan ranked second among world producers.* During the depression following the war, mining underwent a severe contraction in activity, recovering slowly during the late 1920's only to undergo a further setback in the worldwide depression which started in 1929. Mineral output declined in 1930, but the beginning of Japanese expansion, 1931 to 1932, brought about an increase in the output of mine products starting between 1932 to 1934. Production continued to rise, the peak output being attained 1939 to 1943 in virtually all commodities. The greatest production of iron ore, chromite, and manganese was not experienced until 1944 when blocked imports, coupled with strong demands for steel, provided the needed incentive. Through necessity, Japan produced much of her mineral needs from her own mines at high cost during World War 11. Taking the 1932 to 1936 period as an average, however, Japanese mines produced only 16 pct of the iron available to industry, 8 pct of the lead, 33 pct of the zinc, and 68 pct of the copper. Nearly all of Japan's output of gold, silver, copper, lead, and zinc between 1925 and 1945 came from 30 to 49 pct of the mines for which production data are available; for example, 66 of 174 gold-silver mines accounted for 98 pct of the output of this group. Further analysis of mine output indicates that a large part of Japan's metal output was derived from a few large mines; about 10 gold-silver, 15 copper, 3 lead-zinc, 5 manganese, and 2 iron mines account for most of the production. Many of these mines have been operated for 200 to 400 years. With the

Jan 1, 1952

By L. E. Snow, L. F. Pett

AT the Utah Copper Pit of Kennecott Copper Corp. a versatile mobile drill unit has been developed and tested. Through increased drilling speed and additional available drilling time, an improvement of 233 pct in drill performance has been realized. Operators using drill power can quickly and easily convert the unit from a vertical to a toe hole drill. Drilling is done with the machine in line with the unit or at right angles to it. Table I covers the 1950 drill record at the mine, showing the types of machines used and the footage obtained. Development of the new units was prompted by a desire for solution of the following problems: 1—The ever-expanding operations within the Utah Pit, requiring some 65 miles of transmission pipe lines radiating from a central compressed air plant, have now taxed this compressor plant beyond capacity. While 100 lb per sq in. pressure is maintained at compressors, transmission losses from innumerable small leaks result in low terminal pressure at the drills. Fig. 1 illustrates that while terminal pressure varies according to size of air hose and fittings used, the average drill is operating on air pressures within the 60 to 70-lb range. The compressor plant is operated primarily for air rock drills. All drilling is done on day shift, but it is necessary to run two of the three 3720 cu ft capacity compressors during night shift in order to prevent drastic loss of pressure through leaks. On day shift all three compressor units are operated at capacity, so that there is no reserve safeguard available for extensive repairs to any compressor. The calculated volume of compressed air required for operation of 25 drills is but 33 pct of three day-shift compressors. While compressed air is being used to a minor extent for purposes other than drilling, it is felt that satisfactory substitute facilities could be pro- vided in the event that air lines were no longer available. 2—Drill performance has increased beyond the capacity of present 3-in. drifter machines. 3—The installation, maintenance, and moving of air lines in routine mining is a growing problem. The 2-in. level lines are placed outside electrification towers and must be moved to new location upon completion of each shovel cut. In addition to these moves it is often necessary to dismantle level lines so that excessive shovel spillage, or material covering track following a blast, can be pushed in the clear with bulldozers. Where level room is limited, it is often necessary to remove air lines in giving bulldozers room to prepare grades for wagon drills. In spite of precautions taken, many level lines become buried, twisted, and constricted, with excessive air loss resulting. Excessive back break in blasting often leaves several sections of air line suspended in mid-air. Occasionally air lines come in contact with defective electric transmission towers, so that holes are burned through pipe. 4—Rail-handling of drilling equipment such as wagon drills from one location to another requires considerable locomotive crane crew time. Any new unit developed should travel under its own power. 5—Present practice requires accessory material including hose, water pressure tanks, water barrels, drill steel, and bits to be transported by means of push cars when drill locations are changed on any level. 6—In order to supply numerous drill locations with material, many stock piles of new drill rods, drill bits, and other essentials must be left at strategic points, usually in the vicinity of powder magazines. This practice makes for poor housekeeping and contributes to the loss of supplies. Planning the Unit It was felt that a mobile drill unit which would rectify the foregoing problems should meet certain demands: 1—It must furnish compressed air to operate a 4-in. rock drill at a pressure of 90 to 100 lb.

Jan 1, 1953

By L. E. Snow, L. F. Pett

AT the Utah Copper Pit of Kennecott Copper Corp. a versatile mobile drill unit has been developed and tested. Through increased drilling speed and additional available drilling time, an improvement of 233 pct in drill performance has been realized. Operators using drill power can quickly and easily convert the unit from a vertical to a toe hole drill. Drilling is done with the machine in line with the unit or at right angles to it. Table I covers the 1950 drill record at the mine, showing the types of machines used and the footage obtained. Development of the new units was prompted by a desire for solution of the following problems: 1—The ever-expanding operations within the Utah Pit, requiring some 65 miles of transmission pipe lines radiating from a central compressed air plant, have now taxed this compressor plant beyond capacity. While 100 lb per sq in. pressure is maintained at compressors, transmission losses from innumerable small leaks result in low terminal pressure at the drills. Fig. 1 illustrates that while terminal pressure varies according to size of air hose and fittings used, the average drill is operating on air pressures within the 60 to 70-lb range. The compressor plant is operated primarily for air rock drills. All drilling is done on day shift, but it is necessary to run two of the three 3720 cu ft capacity compressors during night shift in order to prevent drastic loss of pressure through leaks. On day shift all three compressor units are operated at capacity, so that there is no reserve safeguard available for extensive repairs to any compressor. The calculated volume of compressed air required for operation of 25 drills is but 33 pct of three day-shift compressors. While compressed air is being used to a minor extent for purposes other than drilling, it is felt that satisfactory substitute facilities could be pro- vided in the event that air lines were no longer available. 2—Drill performance has increased beyond the capacity of present 3-in. drifter machines. 3—The installation, maintenance, and moving of air lines in routine mining is a growing problem. The 2-in. level lines are placed outside electrification towers and must be moved to new location upon completion of each shovel cut. In addition to these moves it is often necessary to dismantle level lines so that excessive shovel spillage, or material covering track following a blast, can be pushed in the clear with bulldozers. Where level room is limited, it is often necessary to remove air lines in giving bulldozers room to prepare grades for wagon drills. In spite of precautions taken, many level lines become buried, twisted, and constricted, with excessive air loss resulting. Excessive back break in blasting often leaves several sections of air line suspended in mid-air. Occasionally air lines come in contact with defective electric transmission towers, so that holes are burned through pipe. 4—Rail-handling of drilling equipment such as wagon drills from one location to another requires considerable locomotive crane crew time. Any new unit developed should travel under its own power. 5—Present practice requires accessory material including hose, water pressure tanks, water barrels, drill steel, and bits to be transported by means of push cars when drill locations are changed on any level. 6—In order to supply numerous drill locations with material, many stock piles of new drill rods, drill bits, and other essentials must be left at strategic points, usually in the vicinity of powder magazines. This practice makes for poor housekeeping and contributes to the loss of supplies. Planning the Unit It was felt that a mobile drill unit which would rectify the foregoing problems should meet certain demands: 1—It must furnish compressed air to operate a 4-in. rock drill at a pressure of 90 to 100 lb.

Jan 1, 1953

By E. R. Borcherdt

IT is notable that the first large-scale mine operation equipped entirely with detachable bits was the Badger State mine of the Anaconda Copper Mining Co. in Butte, Montana, just 30 years ago. This mine in 1922 was producing approximately 1200 tons of ore per day. Much of the data presented in C. L. Berrien's article' describing the development and installation of the Hawkesworth detachable drill bit were obtained from these operations. As in any pioneering effort, no precedent existed and many difficult problems required solution, so that the changeover to detachable bits at all Butte hill mines was not completed for 6 years. There was widespread disbelief as to the probable efficiency of the new installation. Some attempts were made in 1931 by the owners of the Hawkesworth patents to interest Ontario gold mine operators in the bit. These efforts were not successful, but they undoubtedly stimulated thinking which resulted in the invention and patenting of several well-known Canadian detachable bits, one of which is now a widely used throwaway bit. The success of the Butte installation also led to the development of the threaded type of bit connections by several well-known manufacturers, and in 1935 these bits were introduced to the mining industry on a national scale. The original Hawkesworth bit was not provided with a water hole but, depended upon water passing through the clearance opening between the tongue in the bit and the groove in the rod to flush cuttings from the drill hole, see Fig. 1. In December 1935 it was found that this method of introducing drilling water to the bit face resulted in high dust counts. To correct this a water hole was drilled on the central axis of the bit, passing through the tongue. Unfortunately, quenching water would rise through the small water hole, spot-hardening the tongue to cause breakage, never completely eliminated. In the fall of 1936 large-scale tests indicated that savings would be effected by use of a threaded type of bit, which was therefore adopted as standard for all Butte mines. This type of bit was used until 1947, when it was superseded by a one-use slip-on type. Since the first use of the Hawkesworth bit every detachable bit of importance has been investigated, and where advantages which might reduce costs or increase efficiency were indicated, substantial tests of the bit were carried on in the Butte mines. When tests demonstrated the advisability of changing from one kind of detachable bit to another the change was made at one level or in one area each day until the new rod and bit equipment was used throughout the mine. This involved a minimum of cost and disruption of drilling. Intelligent selection of a detachable bit to obtain optimum results requires careful consideration to achieve a balance between the three principal types of equipment used in the drilling process: 1—drill bits, 2—drill steel, and 3—drilling machines. Optimum results imply maximum output and minimum cost per unit of output. Since every rock type differs in drillability and it is generally impractical to provide equipment for more than one or two types of rock which may occur in one operation, selection of equipment must encompass average drilling conditions. However, on exceptional occasions several widely differing conditions may make it mandatory to provide equipment best suited to each condition. The choice of rock-drilling equipment is a most controversial subject and one that is further complicated by unreliable and frequently misleading performance claims. Small operators without the means for making accurate evaluations of equipment frequently suffer from these over-enthusiastic claims. It is apparent from experience in rock drilling throughout the world that rock drillability is not alike in any two places, and that selection of proper equipment can only be made after conducting thorough trials of various types of equipment. Some recent drilling tests in tactite and hornstone at the Darwin, California mine of the Anaconda Co. present some interesting clues on rock drillability. Microscopic examination of thin sections of these rocks reveals that mineral composition and rock texture are equally important in governing drillability. The Darwin hornstone is at times so abrasive that the carbide bit cutting edges become flattened to 3/32 in. in 2 to 4 ft of drilling, and some carbide bits were dulled to this point after 9 to 10 in. of drilling. This wear was determined to be the proper point for resharpening to eliminate carbide insert breakage or breakage of the steel rod when drilling with 1½ to 1?-in. bits, with a drifter of 2 3/4-in. diam and 90 to 100 psi air pressure, see Supplement A. Before considering the merits of various bit designs it may be well to review the mechanics of drilling rock with percussion drills. A sharp bit cuts by penetration and chipping. The amount of penetration governs the amount of chipping and depends upon the contact area of the cutting edge, the foot-

Jan 1, 1954

By Rudolph G. Wuerker

The progress made in testing the strength of rocks and minerals as they are encountered in mine operation is reviewed. An attempt is made to correlate these physical measurements with abrasive hardness, grindability, and behavior in comminution on one hand and fracture of rocks in pillars and roof control on the other. THIS paper reviews the progress made in testing the strength of rocks, ores, coal, salts, and other minerals as they are encountered in mine operations. It attempts to correlate the results of these physical measurements with technological properties more useful to the mining engineer: abrasive hardness, grindability, and behavior in comminution on one hand, and roof control, fracture of rocks in pillars, and mining methods with controlled caving on the other. In the following pages, the materials discussed will be referred to as rocks. Basic to rock mechanics and comminution are the problems of strength, elastic behavior, and failure, common to all brittle materials. A distinction will be drawn as to theoretical and applied research, and discussion of the progress made in each field will include test data obtained by the U.S. Bureau of Standards,1-1 the U.S. Bureau of Mines," the Iowa Engineering Experiment Station,"." the Committee on Geophysical Research at Harvard University,'" Basic Industries Research of the Allis-Chalmers Manufacturing Co.,11,12 by Philipps,13 and by Mueller," to name only a few. With refinements of testing methods and increased standardization. more useful and more comparable results have been achieved. This is especially important in testing a material like rock, as the inherent heterogeneity demands careful and exacting procedures. New measuring procedures that appear to supersede well known standard methods have contributed to faster and less costly testing yet have introduced new concepts, with implications as to comparability of results which must be watched. Reference is made to the sonic method for determining elastic properties," to be discussed in detail below. Basic Investigations Historically, all work in the field has started with the simplest determinations such as those for crushing strength, abrasive hardness, and grindability. These serve the limited objectives in the researcher's field of specialization: building construction, road ballast, roof control in mines, comminution, and seismic prospecting. Occasionally, fundamental properties like the modulus of elasticity E and Poisson's ratio v have been determined with the idea that they might have some bearing on the technological properties of the material under investigation. But it was not until the work of Philipps,13 of Harvard University,'" and of the U.S. Bureau of Mines"' that sufficient basic data were collected to allow researchers to go beyond the technological test and find the fundamental laws behind the behavior of rocks in mine and mill operations. The properties to be looked for are those that describe the elastic behavior of any material, the modulus of elasticity E and Poisson's ratio v being the ones determinable with least difficulties. Only two such properties are required to compute any other property such as the shear modulus, the modulus of rigidity, and the bulk modulus, all of which are related to each other according to well known equations of the theory of elasticity." In spite of their heterogeneous character, all rocks tested have possessed elastic properties. This does not mean that rocks of the same type always have the same modulus of elasticity, which varies exactly as the crushing strength or any other physical property of a rock can spread over a wide range. This has been explained by imperfections of the material always found in rocks, but to some extent this scattering of data is caused by inaccuracies inherent in the testing methods. Modulus of Resilience, a Criterion of Failure Increased availability of E values should allow us to test the validity of the quantity of strain energy theory which has been used in the solution of roof control problems by Philipps13 and by Holland.'" Recently Bond and Wang12 have applied this theory to explain the failure of an elastic material in comminution. Actually it is a very old theory, proposed as far back as 1885 by Beltrami.16 By its assumption the condition of yielding is determined by the term S2 M, = — x volume. Here M, is the modulus of 2E resilience, and its dimension is inch-pounds per cubic inch, that is, work per unit volume. Its numerical value is equal to the area under the stress-strain diagram. In the foregoing equation S is the yield stress (in psi) in tension or compression, whatever the case may be. E, the modulus of elasticity, is in psi. The great appeal of Beltrami's concept of stored energy lies in the fact that the two properties which seem to influence failure most, strength and elasticity, occur in the formula for the modulus of resilience. As an illustration of this, the moduli of resilience in compression of some typical materials tested by the U.S. Bureau of Mines" have been plotted in Fig. 1. The sample of concrete of conventional mix is shown only for the sake of comparison. Its determination was made in the Department of Mining and Metallurgical Engineering, University of Illinois. The values of the moduli of resilience of the various specimens in the plot are:

Jan 1, 1954

By Gerald C. Mathis

FERGUS County, Mont., shown in Fig. 1, is known for its once famous gold mines near the old towns of Gilt Edge, Maiden, and Kendall. But at Heath, a small farming community near the foothills of the Snowy Mountains, thousands of tons of gypsum have been mined since 1916. The value of the gold taken from the mines of Fergus County runs into millions of dollars; the value of the products made from the gypsum of the Heath mine is also in the millions. The gypsum mine will continue to produce for many years, while the gold mines are closed today. The Heath gypsum mine was developed by the Northwest Gypsum Co. in 1916. The products of the small mill built by this company were plaster, agricultural gypsum, and raw gypsum for cement manufacturing. The United States Gypsum Co. prospected the adjacent property in 1916 and 1917 by means of two adits and some exploration drill holes. During World War I, the company abandoned operations at Heath, but again became active in this area in 1928 hen it purchased all the property of the Northwest Gypsum Co. A long range program of expansion was begun with the building of a block plant and the purchase of additional land and mineral rights. A modern wall board plant was built in 1936. When the old wooden mill purchased from the Northwest Gypsum Co. burned in 1937, an all-metal calcining plant located at Blue Rapids, Kansas, was dismantled, moved to Heath, and erected on the site of the mill. In August 1951 a new mine was opened 3000 ft north of the mill. Rock is conveyed from the mine to the mill by an 18-in. belt conveyor. The Heath gypsum deposit lies at the base of the Ellis formation of Jurassic Age. The Ellis at Heath lies unconformably on the black fissle shale of the Heath formation of the Big Snowy Group, which is Mississippian in age. The strike of the gypsum bed is N45°E and the dip approximately 3" NW; the regional strike and dip of the area. The gypsum lies in a series of six layers separated from each other by thin seams of shale. The lower layer is composed of' gypsum and shale intermingled and is not of commercial grade. The upper five layers are high in purity, but only the middle four layers are mined on the advance, as the top layer is left for roof support. This top layer is mined on retreating operations at the same time the pillars are robbed. Total thickness of the gypsum layers varies from 12 to 14 ft. Fig. 2 shows the average thickness and arrangement of the layers. Mining Methods The mine was originally developed by driving an adit into the gypsum horizon. The Northwest Gypsum Co. mined in the top four layers, the top layer being split, leaving about 12 in. for roof support. A modified room and pillar method was used: entries were 15 ft wide and rooms varied between 20 and 30 ft and were turned on 90" from the entries. Crosscuts between the rooms left pillars from 10 to 20 ft square. The rock was loaded by hand into 1-ton mine cars and pulled to a central gathering point by mules; from the gathering point, mules or horses were used to pull the string of cars to the mill tipple. This method was continued by the United States Gypsum Co. from 1928 to 1937.

Jan 1, 1954

By Carville E. Sparks, Rollin Farmin

TRIAL installations of rock bolts, of the slit-rod-and-wedge type, were under way at several units of Day Mines, Inc., when Korean hostilities interrupted the already slow deliveries of steel bars to the Coeur d'Alene district. Factory-made bolts had not yet been put on the local market, so the program was halted for lack of supplies. Interest was revived by a visitor's description of wooden roof bolts. These were said to have been used briefly with apparent success in a coal mine, until apprehension voiced by the U. S. Bureau of Mines caused the practice to be suspended. To make wooden bolts for trial in ground support, Day Mines acquired a second-hand doweling machine equipped with two cutting heads, one to turn out the desired round rods of 2-in. diam, the other to turn out 1-in. rods to be used as powder-tamping sticks. This machine was installed in the all-weather sawmill of the Hercules mine unit at Burke, Idaho, where fabrication of the wooden bolts commenced early in 1951. Most of the mining in the Coeur d'Alene district is along steeply dipping veins in shaly quartzite and argillite of Algonkian age. Ground support commonly is required in zones where the rocks have been sheared, brecciated, and hydrothermally altered. Pressure from the sidewalls is more troublesome than weight overhead, but both increase with the size of the mine opening. Caving may come from a progressive sloughing of irregular rock fragments or from an exfoliation and buckling of the layered wall rocks. The disintegration is thought to develop from an initial elastic expansion of the rock toward the newly-created mine opening, followed by the dilation of many tiny partings in the rock by absorption of water. As the partings widen, masses of rock develop weight and become free to fall. The function of rock bolts is to prevent or retard widening of partings in the rock supported. Wooden Bolts, Wedges and Headboards Bolt assembly used by Day Mines consists of a bolt 4 or 6 ft long, two wedges 16 in. long, and a headboard 30 in. long, Fig. 1. All four pieces are made of local red (Douglas) fir, either green or well-soaked in the mill pond before it enters the sawmill. Bolts are fabricated from cants, 2 1/4 in. sq, cut from relatively straight-grained timber with a minimum of knots and trimmed to 4- and 6-ft lengths. The bolt then is turned in the doweling machine from 21/4 in. sq to 2 in. diam round, except for a 4-in. length at one end which is left full square to provide the striking head and the shoulder that holds the headboard in position for wedging. The foot end of the bolt is slit with a thin saw for a length of about 16 in., thereby making a slot to receive the wedge against which the bolt is driven for anchorage at the bottom of the rock hole. A similar slit, 12 in. long, is made in the opposite (head) end of the bolt to receive the second wedge, which crowds the headboard against the ground at the collar of the rock hole and puts the bolt in tension. The second slot is aligned 90" from the plane of the first slot to avoid Longitudinal splitting and is notched out slightly to allow easier insertion of the collar wedge after the bolt has been driven to bottom. To prevent splitting the headboard by spreading action of the head wedge, this slot is oriented at 90" to the grain of the headboard when the pieces are assembled, Fig. 2. The wedges are similar to standard mine wedges, but more slender; they are cut 1 7/8 in. wide and 1 in. thick at the heel and taper out in 16 in. of length. The headboard, or bearing plate, is not necessary for some types of ground but generally is desirable because it helps the bolt to support an area of loose, friable rock and reduces the tendency for the rock at the collar of the hole to split away from the wedged head by distributing the pressure over a wider rock surface. The headboard may be a 24-to 30-in. length of 3-in. plank, 8 to 12 in. wide, but a similar length of rounded sawmill slab serves equally well at 20 pct of the cost. A hole of 2-in. diam is bored or punched through the center of the headboard, either at 90" or at various high angles to its surface. The bolt is inserted to its shoulder through this hole, then driven into the rock hole. Bolts, wedges, and headboards are given a full timber preservative treatment to inhibit rot. Bundles of each are immersed in a warm saturated solution of Osmose salts in water for 48 hr, removed, dripped dry, and stored in a relatively humid underground depot to cure. Most wooden rock bolts used by Day Mines are 4 ft long. Holes to receive them, about 42 in deep and 2 1/8 in. in diam, are drilled into the rock' to be supported, nearly normal to the periphery of the mine opening. The type of drill used is dictated by convenience: stoper, jackleg, or jumbo-mounted drifter. Correct depth of the hole is assured by use of a measuring stick that has been cut to the proper distance from drill chuck to the ground at the collar of the hole when a standard length drill rod is at the bottom. The bolt is seated to the shoulder through the hole in the headboard, the foot-end wedge is placed in its slot, and the assembly is inserted into the rock hole. Then the bolt is driven until it is seated solidly on the wedge against the bottom of the rock hole. Driving may be by hand with a sledge, or

Jan 1, 1954

By A. E. Lillstrom

IN the development of iron mines and production of iron ore from the Marquette range, drilling blast-holes is an important phase of the mining cycle. The ground drilled in ore production can be classified into two main categories, soft hematite and hard hematite or magnetite. Within these categories the material exhibits a wide range of penetrability by percussion drills. Development work encounters various types of rock. Slate and altered basic intrusives constitute the softer types commonly encountered. Harder materials are represented mainly by greywacke, quartzite, iron formation, and diorite. Prior to the first tungsten carbide trials in late 1947 and early 1948, hard-rock and ore drilling was done with steel jackbits starting at 21/4-in. diam. These were reconditioned by hot milling. Automatic or handcrank 31/2-in. drifters were employed, mounted on Jumbos, posts and arms, or tripods, depending upon the working place. With the exception of shaft sinking jobs where 55-lb sinker machines were and still are used with 1-in. quarter octagon steel, the other production and development mining utilized 11/4-in. round and Leyner-lugged steel. The following properties have been selected as typical examples wherein carbide bit applications have proved economical. The Mather mine "A" and "B" shafts and Cleveland-Cliffs Iron Co. mines are soft ore mines where insert bits are used in rock development only. The Greenwood mine, Inland Steel Co., Champion mine, North Range Mining Co., and Cliffs shaft mine, Cleveland-Cliffs Iron Co., are hard ore mines where all drilling is done with tungsten carbide bits. Mother Mine "A" Shaft In the Mather mine "A" shaft and other soft ore properties where only rock development work is done with the tungsten carbide bits, several types and makes of bits have been tried since early 1948. The greatest proportion of failures have been at the connection end, although the early trials with the 13 Series Carset 11/2-in. bit used in conjunction with 31/2 -in. automatic-feed drifters, showed an equal amount of shattered inserts. To combat this shattering, the 31/2 -in. drifters were replaced by 3-in. drifters, thus eliminating, for the most part, insert failures. However, the attachment end of the rod continued to be the main source of trouble. The greatest amount of failure was in the stud or at the upset section approximately 2 in. behind the drive shoulder of the rod. Heat treatment was changed several times as well as the composition of the alloy studs. Since this failed to correct the trouble, a decision was made to change to a heavier attachment section. Timken 11/2-in., type M, bits were then employed and showed an exceptional improvement. The rods are discarded when the thread contour shows sharpening or wear on the shoulder. It was also learned that the Timken insert did not show as rapid gage and cutting edge wear as did competitive makes, and footage per use increased by approximately 50 pct. Prior to the Timken trials the average life per bit at the Mather mine "A" shaft on 6-ft change chain-feed drifters was 500 ft, and the rod life at the connection end was 50 ft. The Timken bit with chrome-plated thread averaged 1200 ft, and rod life increased to as much as 500 ft. However, the life of the connection end was much better on shorter length drill rods or in places where machines with 34-in. change were used. The bit thread continued to be the point of ultimate failure with thread strippage, constituting the cause for discard of bits. In one of the new development headings, harder rock was encountered for approximately 800 ft, dropping the life per bit to a low of 90 ft with shank and thread life of rods dropping to approximately 125 ft average. The stripped bits were then welded to the rods, increasing the life per bit by 75 to 100 pct. The rod transportation for main level development was not a problem so intraset rods were tried. Intraset rods have tungsten carbide inserts set into the rods proper by the manufacturer and can be obtained with chisel or four point bits. This type of rod eliminates the need for any connection and the steel being a special alloy will show more feet drilled per rod. The first trial was made with eight rods, and final results averaged 350 ft per rod, six of the rods worked the life of the bit end, and two broke shanks at less than 50 ft. The preceding example showed a considerable improvement, so additional steel of the same type was purchased, but its use has been limited to main level drifting only, because of the handling problem involved in transportation of the complete rod to mine shops for resharpening. Further trials are being made on improving the life per detachable bit by chrome plating. To date, the chrome plating shows an improvement of approximately 100 pct. However, final results will not be known until the present long term trials have been completed. Mother Mine "B" Shaft In November 1947, tungsten carbide bits were first tried at the Mather mine "B" shaft. The use of 1%-in. Carset 13 Series bits, for drilling the 72-hole, 7-ft shaft round, decreased the drilling time from an average of 41/2 hr per round required with steel bits, to 2 hr with insert bits. The best drilling time for

Jan 1, 1952

By B. O. Pickard

With more than a score of shafts and numerous stope openings to the surface, all inter-connected underground; with underground temperatures high, often exceeding 100' wet bulb; with an ore presenting a dust health-hazard; with certain rocks giving off inert gases; with an altitude of 6000 ft. above the sea level and a dry desert climate, the Tonopah mines offered an interesting problem to the student of metal-line ventilation. Tonopah is not an old mining district, relatively speaking. It was discovered in 1900 and Tonopah soon was a boom camp. Ry 1910 the mines of this district had paid many millions of dollars in dividends. At the time of the 1921 ventilation study, mining had passed its peak of production and the operators were striving to reduce costs, to discover new orebodies, and to attract desirable workmen. In 1925, the camp was operating on a still less prosperous basis, but the several managements had not relaxed in their efforts to keep the mines on a paying basis, neither had they lost their faith in the ability of the mines and the price of the silver to "come back." Early in the fall of 1921 the Bureau of Mines was invited by the mine operators at Tonopah, Nev., to make a detailed study of the underground air conditions in the Tonopah silver district. D. Harrington, chief engineer of the Safety Service Division of the Bureau, who was then directly supervising the ventilation studies, detailed district engineer B. 0. Pickard, mining engineer E. D. Gardner, surgeon C. E. Kindall, and petrographer IT. Insley to carry on the study. G. S. Rice, as chief mining engineer of the Bureau of Mines, was in general charge of the work. Upon the completion of the study each operator was furnished a report on conditions in his mine. In 1925, a supplemental study at Tonopah was made by the district engineer at the request of several of the operators, and each operator advised on the changes in his respective air conditions and recommendations for improvements. Recently, several of the operators gave the

Jan 1, 1927

By W. F. Netzeband

Jan 1, 1927

By W. J. Taylor, H. J. Mutz, A. F. Brock

THE International Kickel Co. of Canada Ltd. operates five underground mines and an open pit. Four of the mines, the Frood-Stobie, Creighton, Murray, and Garson, are on the south range of the Sudbury Basin, map above, within a radius of 10 miles of the city of Sudbury, Ontario. The fifth mine, the Levack, is on the north range of the Sudbury Basin, 30 miles to the northwest of the city. The open pit is operating on surface ores of the Frood-Stobie orebody. Two pits were opened at the Frood-Stobie mine; the Frood section was started in 1938 and the Stobie section in 1942. To meet the heavy demand for nickel during World War 11, open-pit production was accelerated until it constituted over 40 pct of Inco's total tonnage. As a consequence open-pit ore was depleted considerably alhead of schedule; the Stobie pit was completed in the summer of 1951, and it is expected that mining at the Frood pit will be completed in 1953. The major program launched 10 years ago to develop additional underground production as a replacement of open-pit tonnage thus had to be greatly speeded up. This conversion to all-underground mining will be completed within the next two years, and the underground mines will then have an annual hoisting capacity of 13,000,000 tons. One of the new mines in the underground expansion program, the Murray, was brought into production in 1950. Ore production from the new Stobie section of the Frood-Stobie mine was commenced on a limited scale in 1951 and will be increased as the program progresses. Delivery of ore to the new Creighton mill was started in the summer of 1951 from the caving project launched at Creighton mine. When mining was started in the Sudbury district 65 years ago the ore deposits were opened on a small scale and mining methods conformed to a general pattern of small open cast pits, followed by open stopes as mining was extended below the pits. The open stoping methods used were heading-and-bench and shrinkage. In certain cases a type of cut-and-fill mining was used with drywall tramways and dry-wall chutes in the fill. As the mines were deepened,

Jan 1, 1954

By Borje Hjortzberg-Nordlund

THE great Swedish export iron ore mines are Kiruna and Malmberget in the north above the Arctic circle and Grängesberg in Central Sweden, see Figs. 1 to 3. These mines exported in 1951 about 13 million metric tons of high grade iron ore and concentrates. The origin of these three ore fields seems to be the same. They are of pre-Cambrian age and are regarded as typical instances of primary eruptive ore that has thrust up as glowing masses of lava flow between two somewhat older beds of porphyry. Apparently the ore and the porphyry became segre. gated from common parent magmas by a process of differentiation. Despite the fact that the Kiruna orebody belongs to the earliest pre-Cambrian formation, it has been influenced to an unusually small extent by later metamorphosisms, apart from its having been set on edge through folding and wrenching off by a number of later faults. The Kiruna ore occurs along a stretch of more than 4 miles as a series of slabs or lenses sloping steeply toward the east and sandwiched between two beds of porphyry. The main orebody has a length of about 3 miles and an average width of 300 ft with a maximum of 600 ft. The dip varies from 65" to 45" toward the east and the known depth is 3000 ft, although magnetic surveys indicate it may be at least 6000 ft. Malmberget has undergone a more thorough transformation, having been pressed and crumpled into a meandering band of disconnected layers and lenses. The largest of the more connected bodies of ore, the Great Ore Layer, outcrops on a low mountain ridge and has a total length of about 4 miles and a maximum width of about 300 ft. The dip varies from 30" to 70" towards the north. The main orebody in Grangesberg, the so-called

Jan 1, 1953