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By Benjamin Gildersleeve

Uniquely colored, thin-bedded quartzite is quarried between Crossville and Crab Orchard in Cumberland County, Tenn. It is produced in all sizes up to the limits of transportation from beds usually ranging in thickness from less than Vi in. to 6 in. The stone is very resistant to wear and weathering and is marketed throughout the country for many different purposes in exterior and interior construction. The annual production exceeds 36,000 tons, valued at nearly $1,000,000. THE building stone of the Crab Orchard District is a fine-grained, thin-bedded sandstone which is so strongly cemented by silica that it may be classed as quartzite. As used in this paper the term building stone includes dimension stone, rubble, ashlar, flagstone, and rough broken stone. The Crab Orchard District is in the central portion of Cumberland County, Tenn. It is named from the town of Crab Orchard near which the stone was quarried first for industrial use. The heart of the district is an area of some 1200 acres from 1 to 3 miles east of Crossville which is 70 miles west of . Knoxville and 125 miles east of Nashville (fig. 1). At the present time, there are about 15 quarries worked by companies and individuals. Geology: The Crab Orchard District is on the Cumberland Plateau, or Cumberland Mountain as it is called frequently. Most of the area is characterized by a comparatively even surface from 1800 to 2000 ft in elevation and is drained by streams flowing into the Tennessee River. Included in the district are two quarries located 7 miles west of Cross-ville and this area is drained by tributaries of the Cumberland River. Most of the quarries are located near the 1800-ft contour which is just below the general upland surface. The surface rock over most of the district is the Rockcastle sandstone which has a thickness of about 260 ft.1 It is the topmost of the six formations of

Jan 1, 1951

By H. C. Fowler

No exact demarcation can be made between producfion engineering and engineering research projects which interest production engineers because the results of today's engineering research make possible many of the applied production engineering methods of tomorrow. A corollary of the development of fundamental research is that so-called practical research becomes necessary. Assuming engineering research to include all subjects related to oil and gas production on which knowledge is being increased through experimentation and through analyses of data obtained from both the laboratory and the field, a review of the work of 1932 shows that the fundamental concepts of at least three major problems have become better understood—bottom-hole pressures, solubility of gas in oil, and fluid movement—and that the established data on the relationships of oil and gas have been applied to obtain more efficient and more economic production of both. Engineers no longer are concerned primarily with the initial development of devices for obtaining the pressures at the bottom of wells. Advancement in the technique of obtaining reliable pressure readings in wells has been rapid. One of the first public suggestions regarding apparatus for obtaining precise information at the bottom of oil wells was made before the Petroleum Division of the A.I.M.E. in 1928. Even during the early development stages in East Texas, reservoir pressures were estimated from well-head observations. Soon, however, a number of proved devices and methods were available, and today bottom-hole pressures in many fields are being recorded and studied as an essential part of the production programs. The research now turns to interpretations of data, obtained with these devices, in terms of reservoir energy and to the development of methods whereby the oil and gas will be withdrawn in accordance with rational and equitable programs based on oil-gas-energy relationships. The term ((bottom-hole pressures" should not be interpreted as indicating the full scope of all research on this problem. By using pressure (and temperature) data obtained in related vertical-flow studies,

Jan 1, 1933

By F. E. Steigert, A. H. Geisler

THE permanent magnet alloy comol, which contains 17 pct Mo, 12 pct Co, balance Fe, exhibits a precipitation reaction analogous to that at the iron end of the binary Fe-Mo system. The equilibrium precipitate, (P,), has a structure isomorphous with Fe,Mo, which is rhombohedra1 with parameters a = 8.97A and a = 3038.6&apos; according to the work of Arnfelt and Westgren.&apos; In the ternary alloy the parameters are a = 8.87A and a = 31°0.5&apos;. Oscillating crystal patterns of a single matrix crystal, M, in which P, was precipitated by slow cooling from 1300°C, showed that the orientation relationship of the two lattices is as follows: (110)P3 // (110)* (111)p3 // (110)m There are crystallites of P, in 12 different orientations in a single crystal of the parent solid solution. Microscopic examination shows that the crystallites are generally plate-like with their lateral dimensions parallel to (110) matrix planes. When precipitation in a supersaturated alloy quenched from 1300°C progressed isothermally at temperatures between 600" to 800°C, the evolution of the P, structure could be traced. Diffuse diffraction effects in Laue patterns provided early evidence of the precipitation process. Analyses of these by a previously described technique2 showed that they corresponded to rod-like elements parallel to <110> matrix directions in reciprocal space. With further aging to promote growth of the precipitate, the rod-like elements contracted to points in reciprocal space for normal Bragg reflection from the precipitate as shown by the left diagram in Fig. 1. Thus the diffuse scattering was attributed to the small thickness of the plate-like particles which was inadequate for sharp diffraction. When the precipitate particles were first thick enough to give normal diffraction their structure was different from that of the equilibrium precipitate, P,. The first structure, P1, presumably is a coherent transition struc-

Jan 1, 1952

By S. F. Radtke, J. A. Snyder, R. M. Scriver

DISCUSSION, J. R. Long presiding R. I. Jaffee (Battelle Memorial Institute, Columbus, Ohio)—The authors have written a fine and important paper, and are to be congratulated. I do not entirely follow their discussion of the relationship between the ductility of titanium and its crystal structure. Since high purity titanium is extremely ductile, being capable of cold reductions of ovei- 99 pct, its ductility cannot be unduly restricted by its hexagonal crystal structure. It is true that interstitial solid solutions of nitrogen and oxygen do have a restricting effect on ductility, but this would appear to be an alloying effect rather than a crystal structure effect. The point made by the authors concerning the relative ductility between forged and rolled titanium does not appear to be conclusive, either. It would appear that in the same condition forged material would have equivalent or better ductility than sheet material. The difference between the ductility reported for forged bar and annealed sheet in Table I may be the result of a greater degree of cold work in the as-forged condition. S. F. Radtke, R. M. Scriver, and J. A. Snyder (authors&apos; reply)—We appreciate the discussion by Dr. Jaffee. His point is well taken that the hexagonal structure of titanium does not, in itself, restrict the ductility of iodide material and, therefore, should not be limiting in this case. It is true, however, that the ductility of different grades of titanium is not necessarily related to their hexagonal structure but more likely to their orientation and composition. It is also true that the presence of small amounts of foreign elements in hexagonal metals may have more pronounced influence on ductility than a corresponding quantity in cubic metals. Average values from a large number of tests might show relatively better ductility for forged titanium than is recorded in Table I. The forged bars used in this work were tested in the annealed condition; little is known concerning the effect of forging practice on ductility in titanium metal.

Jan 1, 1952

By Benjamin Gildersleeve

Uniquely colored, thin-bedded quartzite is quarried between Crossville and Crab Orchard in Cumberland County, Tenn. It is produced in all sizes up to the limits of transportation from beds usually ranging in thickness from less than Vi in. to 6 in. The stone is very resistant to wear and weathering and is marketed throughout the country for many different purposes in exterior and interior construction. The annual production exceeds 36,000 tons, valued at nearly $1,000,000. THE building stone of the Crab Orchard District is a fine-grained, thin-bedded sandstone which is so strongly cemented by silica that it may be classed as quartzite. As used in this paper the term building stone includes dimension stone, rubble, ashlar, flagstone, and rough broken stone. The Crab Orchard District is in the central portion of Cumberland County, Tenn. It is named from the town of Crab Orchard near which the stone was quarried first for industrial use. The heart of the district is an area of some 1200 acres from 1 to 3 miles east of Crossville which is 70 miles west of . Knoxville and 125 miles east of Nashville (fig. 1). At the present time, there are about 15 quarries worked by companies and individuals. Geology: The Crab Orchard District is on the Cumberland Plateau, or Cumberland Mountain as it is called frequently. Most of the area is characterized by a comparatively even surface from 1800 to 2000 ft in elevation and is drained by streams flowing into the Tennessee River. Included in the district are two quarries located 7 miles west of Cross-ville and this area is drained by tributaries of the Cumberland River. Most of the quarries are located near the 1800-ft contour which is just below the general upland surface. The surface rock over most of the district is the Rockcastle sandstone which has a thickness of about 260 ft.1 It is the topmost of the six formations of

Jan 1, 1951

By A. W. Hann

THE importance, to geologists and metallurgists, of knowing the mineralogical forms in which the silver occurs in ores has increased during the past few years, and will continue to increase as long as working hypotheses are converted into proved theories and rule of thumb methods are supplanted by scientific data. The methods used for the determination of these silver minerals have been: (1) careful study of polished ore sections with the microscope, and more recently, (2) metallurgical separation of the ground ore by flotation, tabling and panning, supplemented by microscopic examination of these separated minerals, and chemical analyses. During the course of some experimental work carried on by the testing department of the International Smelting Co. and myself, it became important to determine, in a general way, the distribution of the silver values in the minerals, as they exist in Tintic Standard and North Lilly siliceous oxidized lead-silver ores. These mines are in the East Tintic Mining District, State of Utah. In this work the second method was used. Differential flotation tests were run, separating the sulfide from the oxidized minerals, and the products were assayed. Various concentrates were carefully panned and the separated products assayed; special samples of ore were selected and assayed. The microscope was used only for the purpose of determin; ing whether the individual minerals were in the so-called "free" state. Analysis of Tintic Ores These chemical and mineralogical determinations showed the silver to be present as native silver, argento jarosite, argentite, and combined or inherent in the tetrahedrite, chalcopyrite, galena, pyrite, cerussite, anglesite, mimetite, malachite, iron oxide and occluded in the quartz. Careful microscopic work would probably show other forms in small amounts, such as the halide silver minerals and some of the antimony sulfide combinations. The tabulation of the analyses of the various products (Table 1) shows no evidences of fixed ratios of silver to any particular base metal.

Jan 1, 1929

By A. E. Wells

In this paper are described various comparative tests made on cracked and straight run gasolines. The results of these tests may be briefly summarized as follows: Samples of 54 to 55° Bé. gravity cracked gasolines, with end point of about 442° F., and with a somewhat larger proportion of the less volatile fractions than conformable to government specifications, were used in a single cylinder variable compression engine under the same conditions as were samples of straight run gasoline of 58 to 59° Bé. gravity, and of which one had a greater proportion of the more volatile fractions than called for in government specifications. The cracked gasolines showed decidedly less tendency to knock, hence could be used at higher compression ratios and thereby developed from 10 to 20 per cent. more power than did the straight run gasolines. These figures may be taken as one measure of the superiority of the cracked gasolines under test with the straight run gasolines. It is not claimed that all cracked gasolines are superior to all straight run gasolines to the extent indicated by the above tests; cracked gasolines vary as do the straight run products. Cross&apos;s work and that of others have indicated that certain cracked products may have a much greater superiority over straight run than that determined above. From various sources and by different methods of testing data are being obtained to show the definite and generally quite considerable superiority, as an anti-Cetonating fuel of cracked gasolines over straight run gasolines. It is now becoming generally recognized that by increasing the production of cracked gasoline motor fuel from those heavier petroleum fractions, which otherwise would be used for burning purposes, the petroleum industry is not only working in the interest of the conservation of the world&apos;s supply of petroleum, but is at the same time making available to the gasoline-consuming public a product distinctly superior to the straight run gasolines.

Jan 1, 1927

By R. S. Shoemaker, F. W. McQuiston

MAGMA COPPER COMPANY San Manuel Division The San Manuel Division of Magma Copper Company is located about 40 miles north of Tucson. Tons of ore milled in 1973 .were approximately 19 million at a rate of 63,000 tons per day. From this, about 42,600 tons of copper concentrate was produced as saying approximately 1.070 MoS2 and 28.5% Cu. As with many copper ores, the San Manuel ore contains a small amount of gold, but in this particular case, much of the gold follows the molybdenite. Separation of the molybdenite from the bulk sulfide concentrate by flotation produces a moly concentrate containing 9170 MoS2, just under 1.070 Cu, and 0.44 oz Au per ton. Cyanidation of this material for gold recovery is the only practice of its kind known to the authors of this Monograph. After thickening, the final moly concentrate is filtered for subsequent repulping with cyanide and lime. Gold dissolution takes place in three agitators in series with the pulp from the agitators being filtered in two stages in disc filters. Between filtration stages, the cake is repulped with barren solution. Filtrate from both filters is the pregnant solution. The Merrill-Crowe recovery unit is comprised of two solution tanks and the usual clarifying filter, vacuum tower, zinc dust feeder and Merrill precipitate press. Because of the presence of finely divided sooty chalcocite in the molybdenum concentrate, the cyanide consumption of 55 pounds per ton is extremely high. A compensating factor, however, is the low copper, moly concentrate produced. Precipitation of the gold from the highly copper-contaminated solution requires 0.56 pounds of zinc dust and 0.28 pounds of lead acetate per ton, high figures when compared to normal cyanidation practice. The Merrill- Crowe precipitate assays approximately 28.5% Cu, 8% Zn, 156 oz Au and 197 oz Ag per ton. The San Manuel cyanidation plant is an outstanding and innovative development.

Jan 1, 1975

By E. A. Stephenson

In spite of tremendous handicaps notable advances in production engineering have been achieved during 1932. The technique of deep well drilling has continued to progress, notably in California and at Big Lake, Texas. In the latter field a method has been devised by which two grades of oil in contact with each other are produced without intermingling. In rig design many improvements have been made whose objective is to facilitate deep drilling with reduced power losses and increased life of the equipment. Flush joint casing and drill pipe have been improved further, careful tests indicating that the drilling time can be reduced by from 5 to 20 per cent where flush-joint pipe is used. Slow-speed pumping has come into wider use, with further increases in over-all efficiencies. A number of improved well pumps also have been developed and are being tested. One of the most important of these is a positive displacement pump of the rotary type, directly connected to an electric motor placed in the bottom of the hole, and adaptable to various sizes of tubing and casing. In some localities the trend is towards the replacement of electric motors and large-cylinder engines by multiple-cylinder gas engines. Experiments have been conducted and critical analyses made of flow conditions in wells equipped with various styles of tubing which give promise that tubing may be designed to fit individual well conditions. In those wells producing at less than maximum capacity, the gas-oil ratio has been shown to depend on the rate rather than on the method of production, or the design of equipment. Important factors in the design of screened pipe have been carefully analyzed and made the basis of a method for determining the required mesh. Remarkable progress has been made in the use of subsurface pressure data. In East Texas particularly, use of such data is providing a means for the control and equalization of water encroachment and is also pointing the way to a sound means of allocating to individual tracts an

Jan 1, 1933

By Advisory Editorial Board

FOR many years there has been no book that adequately represented the present state of the art of coal preparation-an art that has been rapidly changing during the passing years, and particularly during the last ten or fifteen years. In fact, changes of preparation have been so rapid that no book could keep abreast of their progress, and even the Proceedings of the technical societies have not been able to do so. It has been the intention of this committee and of the editor to make this book as nearly up to date as possible and to make it one that would appeal primarily to operating men, while at the same time adequately showing the theory back of the many processes used. The start of the war and the increased calls upon the time of the authors selected for the various chapters have slightly dimmed the brightness of the ideal intended, but the result undoubtedly brings the practice of preparation and its literature closer together than they ever have been before. It is the hope of the committee that this book will be useful both to those whose daily work is concerned with preparation and to consumers who use the product produced at the many fine and up-to-date coal plants throughout the country, which are now producing, as they always have done, all the coal needed by this country in its greatest crisis. While the committee has exercised a general supervision over the preparation of the various chapters, all of the correspondence and edi¬torial work have been done by Prof. David R. Mitchell, and it has been his untiring interest and devotion that have made the book possible-the credit is all his. Advisory Editorial Board, B. M. Bird Cadwallader Evans, Jr. Thomas Fraser W. H. Lesser William C. Mcciilloch J. B. Morrow H. S. Salmon Raymond E. Salvati Kenneth A. Spencer H. F. Yancey Howard N. Eavenson, Chairman.

Jan 1, 1943

By Sir John Cadnan

The production from the Maidan-i-Naftun area for the year ending March, 1928, was 5,340,000 tons (40,210,200 bbl.), an increase of nearly 13 per cent. over the production for the year ending in March, 1927. The production by years since the opening of the field in 1912 is given in the following table: [1 ton = 7.53 bbl.] Year Ended MaRch Tons Year Ended March Tons 1912 (7 months)............. 43,084 1920....................... 1,354,631 1913........................ 80,800 1921....................... 1,767,070 1914........................ 273,635 1922....................... 2,310,098 1915........................ 375,977 1923....................... 2,913,908 1916........................ 449,394 1924....................... 3,583,000 1917........................ 644,074 1925....................... 4,221,000 1918........................ 897,402 1926....................... 4,470,000 1919....................... 1,081,919 1927....................... 4,735,000 1928....................... 5,340,000 The actual production represents only a small proportion of the potential production. The Masjid-i-Suleiman Field The present producing unit, the Masjid-I-Suleiman, is an anticline some 20 miles in length by 4 miles in width, including three domes as so far proved. The producing horizon is a very open, dolomitic limestone, which, while showing much variation in texture and porosity, is so connected that the lowering of pressure in one section of the field quickly affects the whole area. The topography of the field is broken, with differences of well elevations of as much as 1100 ft. The depths to the limestone vary from about 800 to about 5000 ft. The closed-in pressure of the oil wells varies from zero to 50 lb. per sq. in. and that of the gas wells is about 450 Ib. Although the static oil level may be several hundred feet below ground level, in oil wells of high elevation, all the wells flow continuously or intermittently after they are once started—the majority continuously. The gas is very high in hydrogen sulfide in gas-oil wells of low pressure.

Jan 1, 1929

Maps and Mining Data for the Library. I suggest that the Institute Library can increase its usefulness by systematically soliciting from the engineers and mining companies all published reports and maps, and parts of reports and records, not of a private or confidential nature. If this is done an engineer assigned to report on a well-known property or district may depend on finding at the library a large amount of preliminary data and general information from these official and engineering reports relating to his prospective investigation. In many mining reports a large part of the report is of a general technical and scientific character, and there could be no good objection on the part of either the engineer, or his client to making such information thus of record. I believe there would be a great deal of information furnished by the engineers if they were properly solicited under this plan. I have found on different occasions that much work could be saved by such records now in the Library. The plan would also have the advantage of keeping the Library in closer touch with the active engineers, a feature to be desired. I also have in mind that the Library should undertake to collect all the prospectuses and financial literature which is issued in connection with mining. Of course a great deal of this is unreliable, but even this is valuable for record in the matter of legal and other investigations. I shall be glad to co-operate with you, and do anything I can to develop more in detail the general ideas above noted. Yours very truly, KIRBY THOMAS NOTE.-The Library of the Institute makes a specialty of collecting and filing material of this character, and the letter printed above is brought to the attention of members in the hope that they will co-operate by sending in copies of maps, reports of properties, etc.

Jan 11, 1913

By J. J. Strohl, H. J. Schwellenbach

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.&apos; By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By Francis Pyne

THE following notes are presented in an endeavour to point out the disadvantages attending the use of chrome brick in reverberatory furnaces used in the treatment of materials that are too valuable to allow any great amount of metal absorption by the brick. Several years ago one of the large Eastern copper refineries decided to utilize basic material in place of siliceous material in the walls of their reverberatory furnaces for the treatment of very foul blister copper, as the latter rapidly corroded the siliceous linings. It was also expected that there would be less slag formation, with a consequent decrease in the cost of treatment and a reduction in the metal losses. Magnesite brick were first used, but while the corrosive action of the foul material was greatly reduced, and the amount of. slag formed was very much less, the magnesite proved to be unsatisfactory in certain parts of the furnace, because of its tendency to crack and span badly when subjected to the alternate heating and cooling that takes place in a reverberatory refining furnace. This made many repairs necessary, and consequent frequent interruption in the operation of the furnace, which, together with the high price of the magnesite brick, ran up the cost of maintenance to an unreasonable amount. It was, therefore, decided to substitute chronic brick for magnesite brick in the parts of the furnace affected. The results, as far as the reverberatory furnace was concerned, were very satisfactory. The corrosion due to the action of the foul blister was small and the amount of slag formed was no greater than when using magnesite, and the tendency to crack and spill shown by the magnesite was eliminated. Gradually the use of chrome brick was extended to all furnaces; those treating blister copper as well as those melting cathodes, and the results were so satin-, factory that the siliceous roofs were replaced by roofs of chrome brick except in certain places where experience showed a more satisfactory performance on the part of the silica brick.

Jan 12, 1917

By A. C. Dorenfeld

Ores of the Mammoth mining district, some 45 miles northeast of Tucson, Arizona, are treated by Mammoth-St. Anthony, Ltd. The ores now treated come from a complex system of veins and faults, comprising three mines, in close proximity: The Mohawk, Mammoth and Collins. The ore treated in the past was oxide ore, which contains economic amounts of gold, silver, lead, molybdenum, and vanadium (as wulfenite and vanadanite). The lead-zinc sulphide ore has never been treated before. The Ore The sulphide ore contains galena, sphalerite, marmatite, pyrite and some chal-copyrite, covellite, chalcocite and bornite, as well as small amounts of gold and silver. The value of the ore is principally in the lead and zinc content. The lead-zinc sulphides are substantially liberated at a minus 65-mesh grind. However, a large part of the chalcopyrite and other copper minerals is not freed, but remains as pin points on the sphalerite. Economically, the copper is unimportant. The gangue is a rhyolite granite, with very little limestone. There is no molybdenum or vanadikm in the sulphide ore. History of Sulphide Treatment Test work on samples of the sulphide ore was conducted by several equipment companies, the Arizona Bureau of Mines and the company&apos;s laboratory. Results of this test work indicated that for some reason, which at the time could not be explained, recoveries would be only about 65 per cent lead and 50 per cent zinc. Notwithstanding this, because of the nation&apos;s need for metals, the management built a 150-ton treatment plant. The mill was started in June 1942. During December 1942, using the standard method of treating lead-zinc ore, the monthly results were as shown in Table I. These results were poor. Too much lead was in the zinc concentrate, and too much zinc in the lead concentrate. It should be noted here. that prior to December attempts were made to depress the sphalerite with large and varied amounts of cyanide, zinc sulphate, and sodium sulphite with poor results.

Jan 1, 1947

By H. J. Schwellenbach, J. J. Strohl

This paper relates the results obtained by decreasing the size of feed to a rod mill screen circuit. As would be expected, added production was gained and a finer grind with no tonnage loss was made possible. THE MacIntyre mill of the National Lead Co. has four parallel sections in which magnetic and gravity separations are made. Ore from a 5000-ton bin is fed by Hardinge constant weight feeders to a 6x12-ft open-end Marcy rod mill in each section. Rod mill discharge is bucket elevated to two 4x10-ft Tyrock screens whose oversize is returned to the rod mill. Screen undersize is pumped to Crockett separators for magnetite removal, and the Crockett tailings are sized and tabled to make an ilmenite concentrate. Minus 200-mesh material, overflowed by the Crockett and Dorrco sizers is thickened, de-slimed, and concentrated by flotation.&apos; By 1948, the development of a successful slime flotation process and numerous improvements in the gravity circuit of the mill had increased the capacity of the available concentrating equipment. At the same time, improvements in the operating techniques of the dry crushing plant and improved feeding of the crusher by the mining department had appreciably increased the production rate per operating hour in that plant. Therefore, it was decided to take advantage of the added crusher capacity to produce a finer rod mill feed and in that way increase grinding capacity. At approximately the same time, finer screens were installed in the rod mill circuits to give finer grind and thus improve liberation of the magnetite and ilmenite from the gangue and from one another. Prior to the first part of 1948, a 3/4-in. limiting screen was used for sizing rod mill feed. The first reduction in crushing plant screen size was to 1/2 in. With this screen, the circulating load in the tertiary crushing circuit was too high and plant capacity dropped below a practical hourly rate. A 9/16-in.-sq opening screen, installed next, proved satisfac-

Jan 1, 1951

By C. M. Means

CONSIDERING the hazard involved in mining operations, statistics show, that a very small percentage of accidents is chargeable to electricity. These accidents do represent quite a large percentage of those that are preventable and they are the direct result of the introduction, for purely economical reasons, of a dangerous element. The introduction of electricity in mines should decrease the hazard, and in no case should electricity be applied to the mechanical operation of equipment if by so doing the dangers incidental to mining are increased. The greater number of accidents are the result of persons coming in contact with exposed conductors at potentials varying from 250 to 500 volts direct current. The use of alternating current at higher potentials than these is quite common in large mines, but it is very rare that an accident happens from this source. This is explained by the fact that where these high voltages are warranted, proper precautions are taken, and the installations are directed by those who fully appreciate the dangers incidental to the work. Direct current is in much more general use than alternating current because of its adaptability to haulage locomotives and the low cost of installation in small operations. We are, however, coming to a more extended application of alternating current for the operation of mechanical devices used in connection with mining. Our underground direct-current wiring system is an evolution from the surface trolley and feeder systems. This is probably due to the fact that our earlier successful applications were electric locomotives, which required the use of a trolley wire. The trolley current has been almost banished from industrial establishments on the surface because of the hazard involved, yet we use it indiscriminately underground, where the dangers from accidental contact or fire are infinitely greater. This does not mean that the underground trolley systems should be eliminated, but it does imply the proper safeguarding of such equipment as is necessarily a part of the trolley systems.

Jan 4, 1914

ALTHOUGH the committee appointed by the President of the American Iron and Steel Insti-tute, to consider the twelve-hour day work in the steel industry and report conclusions and recommenda-tions, has made a very careful and painstaking study of the facts and figures developed, it is not yet ready to make what it would designate a final report. Apparently the underlying reason for the agitation which resulted in the appointment of this committee was based on a sentiment (not created or endorsed by the workmen themselves) that &apos;the twelve-hour day was an unreasonable hardship upon the employees who were connected with it; that it was physically injurious to a large percentage of the employes; and that it interfered with family associations essential to the welfare of the children; that for these reasons it was, in a sense, opposed to the public interest. TWELVE-HOUR DAY NOT NECESSARILY INJURIOUS Whatever will be said against the twelve-hour day in the steel industry, investigation has convinced this committee that the same has not of itself been an injury to the employes, physically, mentally or morally. Whether or not, in the large majority of cases, twelve-hour men devote less time to their families than the employes working fewer hours is perhaps questionable. A part of the manufacturing of iron and steel is necessarily continuous. Therefore, practically, there must be two shifts of twelve hours or three shifts of eight hours. The workmen, as a rule, preferr the longer hours because it permits larger compensation per day. It is asserted with confidence that there is less phys-ical work, as a total per day, and less fatigue from the work of the twelve-hour day in the steel industry, than pertains to the large majority of the eight-hour men. This is because in the former case there are more rest periods during the twelve hours on duty.

Jan 6, 1923

By Albert Walcott

THE success of various operations in the manufacture of optical glass depends, in a large measure, on the ability to maintain proper heat control. A good pyrometer system is, therefore, a very necessary part of the equipment needed. Men of long experience in dealing with high temperatures, who have acquired through constant practice the ability of judging approximate temperatures by means of color, are undoubtedly valuable, but even the most competent are not as reliable as a good pyrometer system. They admit that their ability to judge accurately varies from day to day and that their judgment is affected considerably by the condition of the sky, whether it is clear or cloudy. Besides, where it is a matter of only a few weeks to train a man to operate a pyrometer system, it takes about as many years to train him to judge temperatures by the state of incandescence. The pyrometer outfit in use in the glass plant of the Bausch & Lomb Optical Co. consists of the following: Platinum elements, optical pyrometers; base-metal elements, indicators, potentiometers, and recording pyrometers. Platinum Elements and Optical Pyrometers.-Temperature control in - the melting operation is of such great importance that it is very desirable to know the exact temperatures, relative temperatures being of little value. The use of the platinum element is to determine the rate of heating or cooling of the furnace. The platinum couple is installed in either the crown or the wall of the furnace. It is placed first in a porcelain protecting tube and this is placed in a permanently fixed fireclay tube projecting from 1 1/2 to 2 in. into the fire-chamber. It has been found impractical to use the platinum couple for determining exact temperatures of the melting furnace, for the platinum element deteriorates rather rapidly at the high temperatures used, 2500° to 2700° F. (1371° to 1482° C.). When protected and installed as described, though,

Jan 9, 1919

By A. F. Banks

INTRODUCTION Selective flocculation-flotation as a method of desliming potash flotation feed has been practiced in the Cominco Ltd. potash plant since November. 1974 when it was installed to replace a less efficient mechanical desliming process. The plant was designed for a consistent feed containing 4.5% insolubles. In practice the quantity and quality of the insoluble material is quite variable. This variability led to serious problems in the deslime circuit and consequently had an adverse effect on sylvite flotation performance. A costly mechanical solution to the problem was being seriously considered when it was drawn to our attention that a selective flocculation-flotation process had been developed. The process was tried first in the laboratory on the Cominco Ltd. ore, then on a pilot basis on 25% of the mill feed and finally full scale. The time from being informed of the process to full scale operation was 18 months. It has since been incorporated into a tailings regrind circuit. The original mine design called for the mining of an 11 foot height up to the top of the three clay seams. These had to be taken as they were a zone of structural weakness. It also assumed that the beds would be uniform with only a few disturbances. Very soon after the initial start up it was found that permanent openings had to take in the layers right up to the upper clay seams and because of the mining process that this material would have to be milled when encountered. The number of disturbed areas in the mine is greater than the design anticipated. In operation the insoluble content of the mill feed varies from 4.5% to 8% plus with the peaks being very short- term to relatively long, from hours to day. In terms of quality, the insolubles in the mining horizon are relatively soft and easily liberated, whereas those above are hard and liberate slowly.

Jan 1, 1979