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By M. Van Weyenbergh, G. Van Esbroeck

The copper mines of the Katanga region in the Belgian Congo lie along the same mineralized belt as those of Northern Rhodesia. There are two distinct types of deposits in that belt, the dolomitic and the quartzitic. In the dolomitic type the metallic minerals are found in a dolomitic limestone country rock. Through oxidation and enrichment, large deposits of high-grade oxidized copper ores were formed, some of which are being worked by the Union Minière du Haut Katanga. Such deposits are all in Katanga. The quartzitic deposits are found in Katanga and Northern Rhodesia. Here the primary mineralization occurred in sandstones and quartzites, in horizons stratigraphically lower than the dolomites. These rocks are not favorable to secondary enrichment and the oxidized zone is poorer than the sulphide zone, which alone contains the ore. There are also in Katanga a few ore-bearing veins. One of them is being worked in the Prince Leopold mine at Kipushi. Prince Leopold Mine The Kipushi ore body fills a fault vein and invades the wall rocks more or less extensively by metasomatic replacement. The vein strikes north-northeast and south-southwest and dips to the west. The variable dip averages 60' to a depth of 300 ft., and more uniformly 80' below 400 ft. In the upper section the average width is 200 ft. and below the width is about 150 ft. The length of the ore body is 2300 feet. The eastern side of the vein—the footwall—is dolomitic and contains irregular masses of ore sometimes extending 100 ft. or more from the main

Jan 1, 1940

[United Engineering Trustees, Inc. W. D. B. MOTTER, JR., '40 A. L. QUENEAU, '41 ALBERT ROBERTS, '41 The Engineering Foundation GEORGE D. BARRON, '44 F. F. COLCORD, '42 A. L. QUENEAU, '41 Engineering Societies Library Board J. W. LAIST, '40 A. B. PARSONS, Ex-officio T. T. READ, '42 F. F. SHARPLESS, '41]

Jan 1, 1940

By H. S. Auvil

[The shrinkage of coke after formation is one of the variables that affects the net expansion of the change in both industrial and test ovens. Various test results have been interpreted as indicating that cokes from certain coals shrink to a marked extent and that other cokes show very little shrinkage; out no comprehensive experimental study of the shrinkage of coke after its formation and of the factors influencing this shrinkage has come to the attention of the authors. While investigating the reactivity of cokes made by the BM-AGA method,6/ Reynolds7/ found that granular low-temperature cokes shrink considerably then reheated to higher temperatures. This shrinkage was large enough to influence the behavior of the charge (net expansion or contraction) in a coke oven to a marked extent. Since that time the Bureau of nines has investigated the shrinkage of coke different methods, all of thick had one common objective - the elimination of the effect of other factors, such as swelling in the plastic range, which could either partly or completely mask the shrinkage of the coke. Only parts of these studies have been published8/9/ in abbreviated form in]

Jan 1, 1940

By E. R. Maize

This report is the third of a series on convergence of roof and floor in coal mines and subsidence of surface that accompanies it. The first papery dealt with extraction in a mechanized mine (Montour 10) having relatively short rib lines that retreat rapidly. The second 6/ dealt with extraction in a mine (Crucible) in which coal was loaded by hand and in which the rib lines were long and slow-moving. The present report deals with a mine (Gibson) in which conditions approximated those in the Crucible mine, except that the distance between room centers was considerably less and the pillars were narrower in consequence. Moreover, the average cover was less and the studies were made under the steep hillside of a stream valley compared with a gently rolling surface over the portions of the second mine studied. Gibson mine lies in the east-central part of Washington County, Pa., approximately midway between the Montour 10 and Crucible mines on a north-south line. A cooperative agreement covering the work was made with the Hillman Coal & Coke Co. through W. L. Affelder, vice president. This paper describes the methods used in the investigation, gives excerpts from the data obtained, and offers such generalisations as appear to be justified.

Jan 1, 1940

By A. S. Rhea

SALT-WATER disposal in the East Texas field by means of injection into the Woodbine sand is a relatively new procedure. The purpose of this paper is to present what has been accomplished to date in this field, with-out detailed discussion of other means of disposal, such as surface impound-ing for evaporation or release into surface streams. Subsurface disposal of salt water is still in an experimental stage in the East Texas field, and, though water can be injected into the Woodbine sand, the economic problem is not as yet solved. As past experience has been that fields producing from the Woodbine formation produce large quantities of salt water, it was assumed early in development of East Texas that it would be necessary to dispose of a tremendous quantity of salt water in the later periods of production. There has been a gradual increase in the amount of salt water pro-duced in the field, until at present it is about 200,000 bbl. per day from approximately 4400 wells. It is only natural to expect an increase in the number of wells producing water, as well as the amount of water produced in the future, and it has been estimated that if present practices con-tinue the production of salt water may ultimately be as much as 700,-000 bbl. per day. The Woodbine sand is found in the East Texas field at an average depth of 3650 ft. The oil-bearing portion of the sand lies between the depths of 3050 ft. below sea level, on the east edge of the field, and 3320 ft. below sea level on the west, which depth was the estimated original average oil-water contact. The Woodbine sand under the field has the form of a broad tilted wedge pinching out on the eastern side of the field and dipping gently into the East Texas Basin toward the west. The Woodbine sand in the East Texas field has an average porosity of approximately 25 per cent and is highly permeable, the average being about 1300 millidarcys.

Jan 1, 1940

By J. W. Dougherty

THE Hollinger gold mine is in the Porcupine mining division, District of Cochrane, in the northern part of the Province of Ontario. The adjacent town of Timmins, with a population of 25,119, lies 490 miles by railway north of the city of Toronto. GEOLOGY The property is situated on the northwest limb of a major syncline, which pitches northeast at approximately 50°. The principal rocks of the Hollinger area are of pre-Cambrian age and consist of altered, schistose, Keewatin lava flows intruded by stocks of schistose, quartz-feldspar porphyry generally assigned to the Algoman period. The structure of the Keewatin series is extremely complex as a result of close folding and some faulting. The schistosity, trending north 70° east and dipping steeply to the southeast, is most highly developed along a subsidiary anticline in which the major porphyry bodies occur. The ore bodies, which exhibit three principal structures, consist of quartz veins and associated pyritized wall rock. Some lodes are com-posed of en echelon veins, others exhibit a main quartz body with branch-ing stringers and a third consists of a single definite lead. Gold, in a coarse visible form, may be confined to quartz in some ore bodies but in others it is finely divided and is found not only in quartz but also asso-ciated with pyrite in the adjacent wall rock. In many veins the dis-tribution of gold is such that there are several ore sections, separated by areas of low-grade material, in the same vein although its physical appearance remains unchanged.

Jan 1, 1940

By G. E. Steudel

The study of the possibility of effecting a lower cost in the manufacture of pig iron reveals the importance of the ever present question of slag chemistry and volume. Factors that determine slag chemistry and volume are the impurities in the ore available, the ash-sulphur content of the coke, and a flux material necessary to promote removal of sulphur. Of these, the ores present by far the major source of impurities and what may prove to be the lowest practical slag volume with desired chemistry in one locality may be considered low or high in some other center of iron manufacture. The Chicago district is not only blessed with ores of high quality, but with excellent coke and flux materials as well. It is the intent of this paper to discuss slag problems at the South Works only, where burdens consist of Mesabi ores, Michigan stone, and coke ranging from 7 to 9 per cent ash and containing an average of 0.57 per cent sulphur. Early in the development of modern blast-furnace practice, the consideration of slags was secondary to the many adjustments of mechanical difficulties, such as furnace lines, stock distribution, stove design for obtaining higher blast temperatures, improved chemical segregation and size of raw materials. Furnace operators, after making these improvements, turned their attention to the study of other refinements in practice, of which slag chemistry appeared to be the most important. It is only within the last two decades that operators have become familiar with the various aspects of slag compositions. The industry, while rather slow in its recognition, is indebted to such investigators as Rankin,1 Wright,1 Feild,2 Royster,2 McCaffery,3 Kinney,4 Holbrook,5 and Joseph,5 whose researches concerning the elimination of sulphur within the blast furnace and the effect of slag composition upon melting point, viscosity, and desulphurizing power have established criteria for the selection of slags of the requisite quality. Large slag volumes result in higher coke consumption and cost of disposal. Any attempt to lower slag quantity must be accomplished

Jan 1, 1940

By E. R. Maize

[In report of Investigations 3452,5/ baling with roof movement in the Crucible mine, reference was made do to the fact tart timber decay was being studied at the time of writing. This roper, written after completion of that study, describes s the work and summarizes the data obtained. When the study of roof movement as begun at Crucible two ideas were advanced on decay Of timber by some of the officials: (1) That timber decay vas unusually rapid in the Crucible mine, and (2) that timber coated with rock dust decayed more rapidly than those teat were not so coated. There yore no records by which these ideas could be tested, and it appeared wise to make a limited; investigation particularly of the effect of rock dusting on decay, as rock dusting is an important means of preventing coal-dust explosions.]

Jan 1, 1940

By C. A. Heiland

Measurements of gravity with gravimeters have come into increased use in this country and abroad in the past five years. Probably 100 to 125 gravimeter parties are working in the United States alone. In oil exploration, gravimeter methods have relegated the gravity pendulum definitely to second place, if not third, and now closely rival and possibly surpass the torsion balance in many applications. Gravimeters have lately been applied in mining, although it is doubtful whether they will ever replace electrical or magnetic methods in this work. Notwithstanding the widespread application of gravimeters, there has been no article covering the entire field in the domestic or foreign literature. The purpose of this paper is to meet the demand for comprehensive engineering information on the subject. The working principles of a number of gravimeters will be illustrated and treated in relation to the action of corresponding seismometers and astatization methods. The mathematical discussion of theory has been confined to the most necessary fundamentals and been given in as elementary form as possible. A description of calibration methods constitutes the last part of the paper. The symbols shown at the top of page 197 will be used in equations throughout the paper. Dynamic and Static Gravity Methods Methods of measuring gravity fall into two groups: dynamic and static methods. The first group involves the measurement of time. Examples are: determinations of the period of oscillation of a pendulum, measurements involving the free fall of a body, and comparisons of gravity with the centrifugal force of a rotating body. Measurements in the second group of gravity methods are concerned with deflections of certain indicators when the force of gravity is in equilibrium with another comparison force.

Jan 1, 1940

By R. A. Bowen

The Montreal mine, operated by the Montreal Mining Co., is four miles west of Ironwood, Mich., in Iron County, Wisconsin. It is the westernmost producing property on the Gogebic Iron Range of the Lake Superior district. Its production is 4000 tons of hematite per day from an average depth of 2300 ft. This tonnage is hoisted through a five-compartment vertical shaft. Several mining methods are in use but in the larger orebodies the method is sublevel caving. ORE Occurrence and Chaeacteristics The main orebodies of the Montreal mine lie in eastward-pitching troughs formed by the intersections of intrusive dikes with impervious members of the iron formation. The formation has an average width of 425 ft., strikes east and west and dips north at approximately 62. The dikes dip south at 45° and bear north 20' from the strike of the formation, making the intersections with various members pitch east at 16°. The dikes are not continuous across the formation but are interrupted by a longitudinal bedding fault, which has a throw of 380 ft. vertically and 885 ft. horizontally. North of the fault zone the formation is called "hanging," south it is called "footwall" formation. The faulting has created two ore-bearing horizons; viz., at the intersection of banging dikes with the slate members and the foot dikes with the quartzite footwall. In each formation the main ore-bearing member consists of wavy beds of hard chert interlaid with hematite. In the most productive part of the mine, where the sublevel caving method is employed, a hanging-formation dike lies in juxtaposition with a footwall dike and the iron formation carries ore in certain sections from the quartzite footwall to the hanging dike. Here a horizontal section of the orebody may be 300 to 400 ft. wide for a length of 150 ft. or more and tail out lengthwise in both the foot and hanging formations. The footwall ore section is much longer than the hanging wall because of the orebody's greater height above the "pitching dike."

Jan 1, 1940

By Leonard Larson

BEFORE direct or wet smelting of copper concentrate was adopted at the McGill smelter, in November 1932, actual furnace smelting tests had indicated the possibility of smelting between 400 and 500 dry tons of wet concentrate per furnace day. Under calcine smelting practice, solid charge up to 1165 tons per furnace day had been smelted. It was realized that in order to smelt wet tonnage in equivalent amount per furnace day, a higher combustion rate and higher temperatures would be neces-sary within the furnace. The reverberatory furnace considered in the following history of wet-smelting de-velopments is a unit 127 ft. long and 28 ft. 8 in. wide, inside dimensions (Fig. r). This furnace was in operation through all the years from 1934 to date. During 1933, the first complete year of direct or wet smelting, the plant was on a curtailed copper-production basis and the available concentrate tonnage to be smelted was limited. The average coal-firing rate in the reverberatory for that year was 124 tons per furnace day. This rate was lower than the average for years immediately preced-ing under calcine smelting practice. With this coal-firing rate, an average of 391 tons of new metal-bearing material (N.M.B.M.) was smelted per furnace day. The following year, 1934, the coal-firing rate was in-creased to 139 tons and the smelting rate increased to 448 tons N.M.B.M. per fur-nace day. In succeeding years production required that either more tonnage be smelted in the one furnace or an additional furnace be placed in operation. With this situation confronting the smelting depart-ment, it was determined to continue the program of increasing the coal-firing rate and also to intensify the temperature in the smelting zone of the furnace. In order to accomplish this, various improvements were made to aid combustion within the furnace. The burners were enlarged, in-creased primary air was supplied to the burners, and changes were made in the furnace to enlarge and streamline the out-let, thus increasing the furnace draft.

Jan 1, 1940

By C. A. Herbert

In summer, when hot, moisture-laden air enters a coal mine in the region east of the Mississippi River, the moisture is likely to condense on the roof and ribs because the incoming air is cooled below its dew point upon contact with the roof, ribs, and floor of the mine passageways. When the roof and ribs become wet owing to this condensation they are said to "sweat." In many mines, as soon as this sweating begins, shale or slate roof that appears solid and strong in winter begins to cut and fall. For a long time most mining men believed (many still believe) that the roof trouble was due to this sweating. However, from observation of conditions in one mine the operators concluded that the trouble was caused by expansion and contraction of the roof due to changes of temperature. This opinion was confirmed by experiments at several mines in cooling the air as it entered the mines. Methods used in these mines for cooling the air are described in this circular. MINE A At mine A the roof gave considerable trouble in the summer, and it was observed that where a spray of water from a leaking dam kept the roof wet and at a uniform temperature no difficulty was experienced with the roof, although it had cut and fallen on either side of that point. From this fact the company officials concluded that changes of temperature in the air rather than moisture caused the trouble and that if the incoming air were cooled to approximately normal strata temperature much of the additional roof trouble experienced during the summer would be overcome.

Jan 1, 1940

By O. C. Blade

The oil fields of Arkansas, all situated in the southern part of the State, have played an important part in the development of petroleum production in the northern coastal plain region of the Mid-Continent area. From the first discovery of oil in southwest Ouachita County in 1920 to December 31, 1930, a total of 480,167,5604/ barrels of crude petroleum has been produced from' this region. Development of the Arkansas fields proceeded rapidly after 1920. The El Dorado field was discovered in Union County in 1921. Drilling in Ouachita County resulted in the discovery of the Smackover field in 1922 and subsequently in the development of the whole southern area. To November 31, 1938, 6,4595/ producing wells had been drilled in 25 fields, aggregating 53,7356/ acres of proved territory in 6 counties. Peak annual production for the State was reached in 1925, when 80,823,7007/ barrels was produced, resulting largely from flush production from the Smackover field, which rose td an average of 444,9508/ barrel daily during that year. During 1924, 1925, and 1926 Arkansas ranked fourth9/ among the oil-producing

Jan 1, 1940

Sec. 1. This Institute is incorporated under the Membership Corporations law of the State of New York; its corporate name is American Institute of Mining and Metallurgical Engineers. Incorporated; and its abbreviated title A. I. M. E. Sec. 2. The objects of this Institute, as set forth in its certificate of incorporation, are: "To promote the arts and sciences connected with the economic production of the useful minerals and metals, and the welfare of those employed in these industries by all lawful means; to hold meetings for social intercourse and the reading and discussion of professional papers, and to circulate by means of publications among its members the information thus obtained, and to establish and maintain a place for meeting of its members, and a hall for the reading of papers and delivery of addresses, and a library of books relating to subjects cognate to the sciences and arts of mining and metallurgy."

Jan 1, 1940

By William W. Adams

Accidents to men employed in the quarrying and related industries of the United States were more numerous, absolutely and relatively, in 1937 than in 1936. Fortunately, the entire increase was in the category of nonfatal injuries, as accidents causing death declined in number and in proportion to the number of man-hours worked. Employment increased. More men were employed, more man-days and man-hours were worked, and the average employee had more hours of work than in 1936. Reports covering all commercial operations showed that 84,094 men were employed an increase of 5.1 percent over 1936. The number of man-hours of work performed by the industry was, 158 millions, an increase of 7.6 percent. Seventy-seven men were killed by accidents, a, reduction of 14 from the year before, and 6,348 men were injured and disabled for more than the remainder of the day on which the accident occurred, 11 percent more than in 1936. The fatality rate of 0.49 per million man-hours of exposure in 1937 compared favorably with the rate of 0.62 for the previous year, whereas the non-fatal-injury rate of 40.10 was 3.2 percent higher than the rate of 38.87 for 1936.

Jan 1, 1940

So commonplace that they are seldom noticed, mining-machine bits have a defi-nite and important bearing on the cost of coal production. At the average mine many thousands of bits are used during the year. The actual cost represented by the pur-chase of the bits used, as well as that of sharpening and handling them, represents a considerable sum of money. From the point of view of indirect costs they are im-portant, as the lack of good bits at a mining machine can seriously impair production. Good bits tend to ensure an adequate supply of coal.

Jan 1, 1940

By Lawrence B. Wright, Lester G. Morrell

On the steep slopes of the Kootenays, 15 to 20 miles south of Nelson, B. C., are a number of ore occurrences (Fig. 1). One that has many features of mining and geological interest is the Ymir Yankee Girl Gold Mines Limited. Discovery of this property was made in 1899 at an outcrop on the hillside, some 2500 ft. above the village of Ymir. This proved to be a branch vein 20 ft. long leading to two main veins, the Yankee Girl and the Spur (Fig. 5). The ore pattern as now delimited shows a remarkable increase in ore area with depth below the outcrop. This recent development is of interest particularly in view of the fact that the property was nearly condemned, professionally. Another interesting feature, which has a bearing upon its history, is the persistent excess of tonnage produced from period to period over visible reserves. This is not ascribed to lagging mine development, but to detailed vein characteristics, which we will endeavor to make clear by giving a description of the veins and their surroundings. HISTORY AND PRODUCTION Little interest_was_evidenced-in this-distriet until the building of the railway in 1893 provided easy access and the discovery of the Ymir mine in 1896 attracted attention. As a result, the Ymir Yankee Girl group of claims was staked in October 1899. In 1907, an American syndicate took over the property and subsequently extended the 50-ft. drift and opened the present 400 and 540-ft. levels. In 1908, the Syndicate, without recording any shipments in the meantime, dropped its lease, which was taken up by H. L. Rogers, who bonded the Yankee Girl group. Rogers shipped 2622 tons of ore, mainly from the uppermost stope in the mine, which is reported to have yielded

Jan 1, 1940

By A. J. M. Ross

Backfilling of stopes and other underground openings in the Homestake mine with sand tailings was undertaken primarily to reduce surface subsidence, which was wrecking much of the surface plant and a substantial part of the town, and to make possible maintenance of openings into the top of pillar stopes as mining progressed. The progress of this type of filling has been rapid and the results very satisfactory. Sand is now used in filling timbered stopes, shrinkage stopes, and abandoned drifts and crosscuts. A permanent pipe-line system for transporting sand tailings into the mine by gravity has been installed. The close filling of all stopes has resulted in important benefits other than the two major ones. The ore bodies in the Homestake mine occur as replacements in one distinctive layer of the metamorphosed pre-Cambrian sediments along a group of minor folds that together form a complex anticline that pitches at 35" to 40' to the southeast. The ore bodies are extraordinary for their magnitude. Widths of over 300 ft. have been mined and continuous ore has been followed for over 1200 ft. along the strike on some of the upper levels. Mining has been by shrinkage stoping with square-set mining of crowns and pillars. In the standard plan, shrinkage stopes 60 ft. in the direction of strike and the full width of the ore have been taken up to within 25 ft. of the level above. After filling, the crown has been mined by square-set methods, as have also the 40-ft. pillars between shrinkage stopes. In former days, when the supply of ore seemed inexhaustible, the mining of pillars was allowed to lag far behind because of the high cost. Many pillars were rather badly fractured on the level and in some places between levels. Coarse waste was used for backfilling all stopes, but in pillars it became increasingly difficult to maintain haulageways for the transportation of this fill over the top of the stopes as mining progressed. The cost of maintaining these crosscuts was excessive and in many pillars practically impossible. Therefore it became the practice to fill pillar stopes by tapping the back and sides of the stopes.

Jan 1, 1940

By H. H. Stout

IN the early fall of 1925, the writer was conducting, in the Ledoux and Co. labora-tory, New York, experiments directed to-ward ascertaining the effect on its impurity content when cathode copper was subjected to a current of various gases at an elevated temperature below the melting point, The apparatus consisted of a vertical 1-in. dia. silica tube about 12 in. long, heated on the outside, The cathodes were broken to pass a8-in. mesh, and a current of various differ-ent gases was passed up through a 4-in, long column of these loose cathode particles in the vertical tube furnace. Temperature was maintained to i500° to 1600°F. and the gas treatments lasted from 3 to 6 hr, The writer observed that whenever reducing gas (H2 or CO) was used, these cathode par-ticles stuck together at points of contact with each other. The cohesion was so marked and tenacious that an examination was made of the fractured surfaces, when cathode particles were broken from the cluster. Crystal growth was always found to have taken place across the surfaces where separate particles were in contact, but only after all oxides, sulphates, etc, had been completely removed. The idea of a possible new process was then conceived and was disclosed to A. M. Smoot (V. P. Ledoux and Co.) with intent to obtain his views. The first conception as given Smoot was "to gas-clean broken cathodes, put them in an extrusion press, solidify and extrude them at around 1500°F," Smoot's reaction, after a few moments considera-tion, was, "It looks perfectly all right to me."

Jan 1, 1940

By E. M. Wise

IT is known qualitatively that the recrystallization behavior of platinum is dependent upon the amount and particularly the nature of impurities present, the amount of prior cold-work and the annealing time and temperature, but published data are limited. A study of the effects of cold-work and annealing temperature on the hardness of a sample of very pure platinum was undertaken to supply part of this information. CONCLUSIONS 1. The platinum used in the experiments had a probable platinum content of 99.99+ per cent, and had the following mechanical properties in the completely softened condition: Vickers hardness number, 39.4; ultimate tensile strength, 19,340 lb. per sq. in.; proportional limit, very low; elongation, 37.5 per cent in 2 inches. 2. The effect of cold-rolling on the hardness is shown in Fig. 1. The hardness increases rapidly up to 10 to 15 per cent reduction and more slowly up to about 75 per cent reduction, beyond which it increases more rapidly. 3. The recrystallization temperatures for a 15-min. anneal, after cold-rolling 19, 39.5, 50.8, 66, 80.5 and 89.5 per cent reduction in thickness, are approximately 765°, 700°, 635°, 600°, 545° and 425°C., respectively. 4. The recrystallized grain size of this platinum is a function of the percentage of cold reduction, being larger the lower the percentage of cold-work. 5. Little or no grain growth resulted from subjecting recrystallized samples to higher temperatures or annealing for longer times.

Jan 1, 1940