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By J. L. Bray

IT is well known that iron markedly improves the physical properties and casting characteristics of aluminum bronzes. The use of 1 to 4 per cent iron as the principal modifying addition has become common practice in many foundries melting aluminum bronzes. Although the beneficial effects of iron are unquestioned, there is some uncertainty regarding the structural changes caused by it in the basic copper-alumi-num system. It is agreed that a new constituent appears in abundance when the iron content reaches 2 to 4 per cent. This constituent, which frequently occurs in shapes resembling clover leaves, is sometimes assumed to be an aluminum-iron intermetallic compound. If this were correct we might expect a change in the structural relationships of the matrix in the direction of decreased aluminum content. On the other hand, it has been reported that iron has a negligible effect in altering the fundamental relationships of copper-aluminum alloys1. Others2,3 have reported an additive coefficient. Apparently it will be necessary to know the equilibrium relationships between the phases of the copper-aluminum-iron alloy system to establish any such theory as fact. When this has been done it may very well be that a further revision in the structural relationships may be necessary to describe the changes to be expected in commercial cast alloys. It was with this latter problem before them that the authors attempted to

Jan 1, 1936

By Charles Goodale

With notes on other mines of the Marysville district by WALTER McDERMOTT, London, England, and F. L. Sizer, Dos Cabezas, Ariz. (Salt Lake Meeting, August, 1914) THE purpose of this paper is to review the history of the Drumlummon, one of the famous old mines of the West. Mining engineers, when sent to examine new mines in old districts, or to decide whether an old property, under new conditions, will pay to re-open, have had difficulty in learning anything authentic regarding workings which are under water, and historical facts about the working character of the ore, the total production, and the value of the ore with increasing depth. Not many years elapse after the pumps are pulled, before maps, books and records are destroyed, and the value of the publications of the Institute would be increased if they contained the history of many old mines now abandoned. The unsuccessful efforts of the writer quite recently in trying to obtain reliable information about a gold mine which was reported to have produced several million dollars between 1880 and 1888, the workings of which are now partially caved and under water, impressed upon his mind the desirability of having a permanent record of old properties. In this particular case, the owners insisted that under present conditions of improved methods of treatment -particularly by the use of cyanide, which was not an applied process when the mine was operated, the property. could be made to pay a handsome profit on ore left in sight in the old workings, but no assay records could be found to confirm these assurances, and even the records of bullion shipments had been destroyed. If surveys and maps were made as the work progressed, no trace of them could be found, and therefore opening of the ground through a new shaft and drifts therefrom would be attended by some risk while exploring near the old workings. Fatal accidents by flooding under these conditions have occurred, and as an evidence that miners have fears of such dangers, it was urged upon the Mining Committee during a recent session of the Montana Legislature, by a delegation of

Jan 8, 1914

By C. H. Li, R. J. Stokes, T. L. Johnston

A portion of the Ni3Cr-Ni3Al phase dzagram has been investigated, including the precipitation of 1 (Ni3Al) as well as the existence of ordered Y (Ni matrix), Extensive metallographic studies by electron microscopy have been carried out and X-ray diffraction diagrams have been taken at both elevated and ambient temperatures, No long-range ordering of the phase was found above 893°C (1639°F) in the compositional range investigated. The high rate of formation of the ordered ?1 phase upon quenching has been confirmed, a phenomenon which up to the present Lime has retarded the understanding of this phase diagram, Specimens with controlled amounts of ?1 exhibited up to a six-fold improvement in yield and creep strength compared with single-phase Ni-Cv-A1 alloys. MOST of the important high nickel heat-resistant alloys are related to the nickel-chromium-aluminum system. One of the important functions of the aluminum, often accompanied by titanium, is to produce a precipitate of the very fine ?1 phase (Ni3A1, X) where X may be Ti, Cr, or other elements. A number of investigators1-" have demonstrate the presence of y' in these alloys but most of their materials contained large amounts of other elements resulting in appreciable amounts of con- fusing phases such as carbides. Further clouding of the precise role of y' has been the suspected existence of an ordering reaction in the y matrix itself, ro-3,12-17 It appeared from these investigations that aluminum in Ni-Cr alloys would raise the ordering temperature of the matrix from 540°C at NisCr to 1300°C at Ni,Al. In support of this possibility, the matrix phase in the parallel Ni-Fe-A1 system is ordered to very high temperatures.1 OBJECTIVES OF THIS RESEARCH From this survey of the literature it appeared desirable to conduct an investigation of the simple Ni-Cr-A1 system using the alloys of the highest available purity to avoid the complication of foreign phases and thereby isolate the structures in the Ni-Cr-A1 system. Procedures were developed to study ? - ?o and ? - ?1 equilibria. Preliminary mechanical tests at elevated temperatures were also conducted.

Jan 1, 1962

PRESIDENTS DAVID THOMAS 1871 Born November 3, 1794, Cadoxtan, Glamorganshire, South Wales. Died in 1882. Attended schools in Wales. Came to America in 1839. A.I.M.E.: Member, 1871; Honorary Member. Iron manufacturer; inventor. Built first anthracite fired hot-blast furnace in America. Known as the "Father of the Lehigh Valley." President, Catasauqua and Fogelsville Railroad; Director, Lehigh Valley Railroad; Founder, Thomas Iron Co. and St. Luke's Hospital at Bethlehem, Pa.; Trustee, Lafayette College.4 ROSSITER WORTHINGTON RAYMOND 1872-1874 Born April 27, 1840, Cincinnati, Ohio. Died in 1918. Brooklyn Polytechnic Institute, 1857; Lafayette College, Ph.D., 1868; Lehigh University, LL.D., 1906; University of Pittsburgh, Hon. LL.D., 1915., Gold Medal, Institution of Mining and Metallurgy (London). A.I.M.E.: Founder, Member of 1871; Vice President, 1871, 1876-1877; Secretary, 1884-1911; Secretary Emeritus, 1911-1918; Honorary Member. Consul¬ing mining engineer; practicing lawyer, sailor, soldier, writer, orator, editor, theologian, teacher, novelist, chess player. Recognized as the "Grand Old Man of the Institute." ALEXANDER LYMAN HOLI,EY 1875 Born July 20, 1832, Lakeville, Conn. Died in 1882. Brown University, 1853; Hon. LL.D., 1878. A.I.M.E.: Member, 1871; Vice President, 1874; Manager, 1877-1879. Bessemer Medal (Iron and Steel Institute, London), posthumously. Brought bessemer process to United States. A statue was erected in Washington Square in his memory and presented to the City of New York in 1889. The inscription notes: "foremost among those whose genius and energy established in America and improved throughout the world the manufacture of bessemer steel." ABRAM STEVENS HEWITT 1876 Born July 31, 1822, Haverstraw, N. Y. Died in 1903. Columbia University, 1842; Hon. LL.D., 1887. A.I.M.E.: Member, 1871; Manager, 1872-1875. Bessemer Medalist and Honorary Member, Iron and Steel Institute (London). Gold Medal, Chamber of Commerce of New York City. Philanthropist, lawyer, statesman, publicist, pioneer in the iron and steel industry. Mayor of New York City in 1887. Business partner of Peter Cooper, Executive Manager of Board of Trustees, Cooper Union, New York City.

Jan 1, 1947

By Charles W. Goodale

The purpose of this paper is to review the history of the Drumlummon, one of the famous old mines of the West. Mining engineers, when sent to examine new mines in old districts, or, to decide whether an old property, under new conditions, will pay to reopen, have had difficulty in learning anything authentic regarding workings which are under water, and historical facts about the working character of the ore, the total production, and the value of the ore with increasing depth. 'Not many years elapse after the pumps are pulled, before maps, books and records are destroyed, and the value of the publications of the Institute would be increased if they contained the history of many old mines now abandoned. The unsuccessful efforts of the writer quite recently in trying to obtain reliable information about a gold mine which was reported to have produced several million dollars between 1880 and 1888, the workings of which are now partially caved and under water, impressed upon his mind the desirability of having a permanent record of old properties. In this particular case, the owners insisted that under present conditions of improved methods of treatment —particularly by the use of cyanide, which was not an applied process when the mine was operated, the property could be made to pay a handsome profit on ore left in sight in the old workings, but no assay records could be found to confirm these assurances, and even the records of bullion shipments had been destroyed. If surveys and maps were made as the work progressed, no trace of them could be found, and therefore opening of the ground through a new shaft and drifts therefrom would be attended by some risk while exploring near the old workings. Fatal accidents by flooding under these conditions have occurred, and as an evidence that miners have fears of such dangers, it was urged upon the Mining Committee during a recent session of the Montana Legislature, by a delegation of

Jan 1, 1915

By Royal B. Jackman, Carle R. Hayward

Two comprehensive papers have appeared regarding the forms of copper that occur in smelter slags, one by Frank E. Lathe1 and the other by C. G. Maier and G. D. Van Arsdale.2 These authors comment on other published results and draw conclusions from their own investigations. The development of X-ray examination of minerals and improvement in high-power microscopic methods make it possible today to get additional evidence by means that were not available to earlier investigators. The present work was confined to reverberatory slags obtained from five different plants. Some definite conclusions were reached regarding the forms of copper present in these slags. In most cases these agree with some of the findings of Maier and Van Arsdale but new methods of attack were used in the investigation. Conclusions of Previous Investigators Various important conclusions reached by previous investigators are summarized in the following paragraphs, arranged chronologically. William A. Heywood:3 Slag losses increase with grade of matte but with the same grade of matte losses decrease as the acidity of the slag increases. Lewis T. Wright:' Slag losses are only partly due to prills of matte. It is suggested that part of the copper is dissolved in the slag. J. Parke Channing9 A considerable part of the copper enters the slag as silicate or oxide and not as sulfide. C. A. Grabill:6 Slag losses are both physical and chemical. Physical losses, due to poor settling, gas flotation and slight differences in specific gravity, are small under normal conditions. Chemical losses may- be due to formation of oxides or to solution of metals and sulfides. The last is considered the most important. John W. James:' The copper in converter slag is largely the oxide and as such enters the reverberatory slag when converter slag is added to the charge.

Jan 1, 1934

By E. R. Johnson

The copper concentrator of the Braden Copper Co. is at Sewell, Chile, on the westem flank of the main Cordillera of the Andes, at an air distance of approximately 50 miles southeast of Santiago, the capital of Chile. The region is mountainous and space is at a premium because of the rough terrain and the danger of snowslides. The ore, a porphyry copper, is mined by a block-caving method in the El Teniente mine. It is hauled to the crushing plant, a distance of 2.6 miles, through a tunnel because of the severe climatic conditions. The usual large, coarsely set, primary reduction machines found in many ore-treating plants are not required because of the mining method employed and the limiting 12-in. spacing of the grizzlies at the head of each ore pass in the mine. It is the purpose of this paper to describe the numerous improvements made in the crushing section of this large flotation concentrator during the past I5 years to increase the efficiency of the plant and to raise the daily output from around 20,000 to 30,000 tons. The reader should keep in mind that all changes were made within the existing building because of the limited space and the high cost of the heavy mill construction required to withstand the snow loads. The present plant, as well as the former, operates on a 24-hr, 7 days a week basis. The crushing plant commenced operations at its present site in 1921, the main equipment being No. 7½ gyratories, followed by 48-in. horizontal Symons disk crushers. Fines were removed by I-in. stationary grizzlies ahead of each unit. In 1924 a third stage of crushing was provided by the addition of 55½-in. rolls in open circuit. During 1925, the Symons disks were replaced by 73½-in. rolls. All the rolls were changed to closed circuit, with elevators and screens to give a final product of 9 pct plus 3-mesh. The flowsheet of a typical crushing section of this period is shown in Fig I. The ore was withdrawn from a common bill serving the five crushing sections in the plant by means of any number up to six 36-in. wide-apron feeders, which discharged onto a common conveyor belt. This conveyor was equipped with a magnetic head pulley and the crusher attendant was stationed near the discharge end to pick out pieces of mine timber that might be in the ore. The ore was discharged over a 13/4-in. spaced bar grizzly and the oversize of this grizzly was crushed by a No. 7½ McCully gyratory crusher set at 13/4-in. opening on the closed side. The grizzly undersize joined the gyratory discharge and was conveyed to a set of 73½ by 20-in. Garfield rolls set at I in. These rolls operated in closed circuit with a bucket elevator and two vibrating screens placed in parallel.

Jan 1, 1949

By W. F. Mattes

The manufacture of steel rails in the United States upon a large scale may be roughly dated from the years 1875-76, and the same years witnessed an active movement among the railroads toward the adoption of heavier and improved patterns. In this movement each road was a law unto itself, and the "personal equations " of some scores of engineers were so freely injected into the calculations that patterns were multiplied in the most absurd and mischievous manner, until at length the burden upon the mills became wellnigh unendurable, and called forth a vigorous protest in Mr. Holley's paper,* read before the Institute in February, 1881. There were then regularly manufactured in the various mills of the United States, no less than 119 different patterns of steel rails, subject to current order, of which 64 per cent. were embraced under five weights per yard. Commenting upon these and other figures there given, Mr. Holley stated as the primary object of his paper, " to show that,, while this multiplication of rail-patterns is an annoyance and an expense to rail-makers, it is a source of immense loss to railway companies." Since 1876 the increase in gross annual tonnage upon our leading lines has been very great, and has been accompanied by an increase of wheel-loads and running-speeds. As the average weight of the rails of 1876 was by no means excessive, the growing deman Is of traffic have slowly driven the roads toward stiffer sections, inaugurating a movement that now promises to become epidemic. In 1876 a 67-pound rail was a first-class equipment; to-day everything under 75 pounds is light, and many 80's are coming into use. European practice has long been heavier than ours; in part because the higher cost of sleepers has led to an average of 3-feet spacing, instead of two-feet as with us. The old 84-pound double-header, and Sandberg's 80- and 89-pound rails, respectively 51/4 inches and 591/2 inches high, are reported common on English roads; and he is strongly advocating the general adoption of the 100-pounder shown in Fig. 7, with what success I have not learned, except that the Belgian government has ordered a lot. Mr. Sandberg points out the undesirability of frequent changes of pattern, and the recent rapid growth in traffic-requirements; prognosticates a continuance of this

Jan 1, 1887

ASIDE from the John Fritz Medal, in which the Institute participates through its representation on the John Fritz Medal Board; the Institute itself has three awards to make annually, as follows: the Robert W.' Hunt Medal and Prize, the J. E. Johnson, Jr. Award, and the James Douglas Medal. ROBERT W. HUNT MEDAL AND PRIZE The partners of Robert W. Hunt established a fund which was presented to the Institute at a fitting ceremony on May 27, 1920, to establish an annual medal and a sum of money to be awarded under the following rules: 1. The award shall consist of two prizes; first of a. gold medal, second of a sum of money; a certificate to accompany each prize. The money prize shall not be awarded to a member over 40 years of age, but under unusual circumstances, both prizes may be allotted to one person provided that he is not over 40 years of age. In general it will be understood that the Committee shall award the money prize to the younger men, rather than the medal. 2. The awards shall be made not oftener than once a year to that person or persons contributing to the Institute the best original paper or Papers on iron and steel. The scope of the term "iron and steel" shall be determined by the sub-committee considering the awards. In general, papers dealing with the practical side of. the subject have preference over those dealing with the theoretical, side in recognition of the fact that Captain Hunt's main contributions to the industry have been in the improvement of production and quality of material. 3. A sub-committee of three to five, including the Chairman of the Iron and Steel Committee, shall be appointed by the Iron and Steel Committee annually to adjudge the. award subject to approval. 4. The awarding committee shall submit its report to the Iron and Steel Committee at its October Meeting, and the award shall be certified to the Secretary of the Institute in time to permit the presentation to be made at the Annual Meeting of the Institute. 5. The recipient of the award shall be designated "The Hunt Medallist." This prize is administered. by the iron and Steel Committee. J. E. JOHNSON, JR., AWARD This award is made from the income of a fund of $30000 donated by Mrs. Margaret Hilles. Johnson in memory of her husband, J. E. Johnson, Jr., who was a prominent engineer, author of two valuable volumes on iron blast-furnace construction and practise, vice-chairman of the Institute's Iron. and Steel Committee, and a frequent contributor of papers to the Institute's TRANSACTIONS. This prize is administered by the Iron and Steel Committee. The intent of the donor is to encourage young men in creative work in branches of the metallurgy or manufacture of pig iron with which the professional activities of Mr. Johnson were chiefly concerned. The control off the fund and distribution of the awards having been vested in the Board of Directors of the Institute, this Board has made the following regulations concerning it:

Jan 1, 1925

By E. R. Johnson

THE copper concentrator of the Braden Copper Co. is at Sewell, Chile, on the western flank of the main Cordillera of the Andes, at an air distance of approximately 50 miles southeast of Santiago, the capital of Chile. The region is mountainous and space is at a premium because of the rough terrain and the danger of snowslides. The ore, a porphyry copper, is mined by a block-caving method in the El Teniente mine. It is hauled to the crushing plant, a distance of 2.6 miles, through a tunnel because of the severe climatic conditions. The usual large, coarsely set, primary reduction machines found in many ore-treating plants are not required because of the mining method employed and the limiting 12-in. spacing of the grizzlies at the head of each ore pass in the mine. It is the purpose of this paper to describe the numerous improvements made in the crushing section of this large flotation concentrator during the past 15 years to increase the efficiency of the plant and to raise the daily output from around 20,000 to 30,000 tons. The reader should keep in mind that all changes were made within the existing building because of the limited space and the high cost of the heavy mill construction required to withstand the snow loads. The present plant, as well as the former, operates on a 24-hr, 7 days a week basis. FORMER PLANT The crushing plant commenced operations at its present site in 1921, the main equipment being No. 7 1/2 gyratories, followed by 48-in. horizontal Symons disk crushers. Fines were removed by 1-in. stationary grizzlies ahead of each unit. In 1924 a third stage of crushing was provided by the addition of 55 1/2-in. rolls in open circuit. During 1925, the Symons disks were replaced by 73 1/2-in. rolls. All the rolls were changed to closed circuit, with elevators and screens to give a final product of 9 pct plus 3-mesh. The flowsheet of a typical crushing section of this period is shown in Fig I. The ore was withdrawn from a common bin serving the five crushing sections in the plant by means of any number up to six 36-in. wide-apron feeders, which discharged onto a common conveyor belt. This conveyor was equipped with a magnetic head pulley and the crusher attendant was stationed near the discharge end to pick out pieces of mine timber that might be in the ore. The ore was discharged over a 1 3/4-in. spaced bar grizzly and the oversize of this grizzly was crushed by a No. 7 1/2 McCully gyratory crusher set at 1 3/4-in. opening on the closed side. The grizzly undersize joined the gyratory discharge and was conveyed to a set of 73 1/2 by 20-in. Garfield rolls set at I in. These rolls operated in closed circuit with a bucket elevator and two vibrating screens placed in parallel.

Jan 1, 1947

By O. W. Simmons, L. W. Eastwood, C. M. Craighead

Eighty-four quaternary titanium-base alloys from the following systems were investigated: 1. Titanium-chromium-carbon-nitrogen. 2. Titanium-chromium-carbon with copper, vanadium, molybdenum, manganese, iron, or nickel. 3. Titanium-chromium-nitrogen with vanadium, molybdenum, or nickel. The tensile properties, minimum bend radii, hardnesses, response to heat treatment and aging treatment, and phase relationships of these alloys are reported. HIS paper is the third in a series describing J- titanium-base alloys. The first paper describingdescribed titanium binary alloys, and the second contained data on a large number of titanium ternary alloys. Eighty-four heats, comprising three groups of quaternary titanium-base alloys, are described in this paper as follows: 1. Titanium-chromium-carbon-nitrogen alloys. 2. Titanium-chromium-carbon alloys containing copper, vanadium, molybdenum, manganese, iron, or nickel. 3. Titanium-chromium-nitrogen alloys containing vanadium, molybdenum, or nickel. The methods of preparing, fabricating, and testing the alloys are described in the paper Binary Titanium Alloys. Quaternary Titanium - Chromiurn - Carbon -Nitrogen Alloys The properties of titanium-chromium-nitrogen and titanium-chromium-carbon ternary alloys have been described in the paper Ternary Alloys of Titanium. Because small amounts of nitrogen and carbon both appear to be somewhat beneficial, these additions have been investigated fairly completely in quaternary alloys containing chromium at the following levels of composition: 2.5, 3.5, and 5.0 pct chromium 0.25 and 0.5 pct carbon 0.1 and 0.2 pct nitrogen Mechanical Properties: The mechanical properties of these alloys are listed in table I. Fig. 1 and 2 show graphically a comparison of the tensile properties and hardnesses of quaternary alloys with those of the titanium-chromium binary alloys. Fig. 1 shows the comparison at 0.25 pct carbon, whereas fig. 2 shows the comparison at 0.5 pct carbon. It is concluded from these data that 0.25 pct carbon and 0.1 pct nitrogen are desirable additions to the 5 pct chromium-titanium base. It is further concluded that most of this improvement is produced by the nitrogen and that carbon is either harmless or mildly beneficial, at least up to 0.25 pct. Structure and Transformation Range: The effects of carbon and nitrogen on the transformation range of the titanium-chromium alloys were determined as part of the routine investigation. Additions of 0.25 pct carbon and 0.1 pct nitrogen had no appreciable effect upon the transformation range. However, the addition of 0.25 pct carbon and 0.2 pct nitrogen definitely raised the beta solvus temperature, as shown by fig. 3.

Jan 1, 1951

By G. Chesneau

Two great discoveries of this century have diminished the dangers of fiery coal-mines,—the sifety-lamp, conceived in 1815 by Sir Humphrey Davy and successively improved by many engineers, such as Clanny, Marsaut, Mueseler, Fumat, etc., and safety-explosives, the appearance of which is comparatively recent, hut which, thanks to the persevering labors of several technical commissions (especially the French commission on explosives), and of the manufacturers and experimental users of these powders, appear likely to play, in the protection of the miner, a part as important as that of the safetylamp itself. Brit while it is true that by the use of a safety-lamp of approved form, and the exclusive employment, with all prescribed precautions, of these new explosives, complete security should be assured, even in highly fiery atmospheres, experience has shown that the imprudence of workmen, or the bad condition of a lamp, may nullify these protective means, so that it ought to be the constant care of the engineers of fiery mines to dilute the fire-damp by good ventilation so far as to maintain always and in all parts of a mine a proportion of this gas below the lower limit of explosiveness belonging to a mixture of air and methane. Good ventilation still remains the best preventive of explosions from fire-damp, and since, even with a rational and well-controlled mine-ventilation, the volumes of air traversing galleries and rooms are liable to vary within wide limits by reason of open doors, falls of rock, or other accidental causes, it follows that but a few thousandths of fire-damp should be left in the normal atmosphere of the mine-workings, so that a considerable

Jan 1, 1894

By F. G. McCutcheon

UNTIL about 1930, the universal practice of horizontal-retort zinc smelters in the United States, as far as the writers are aware, was to operate the retort furnaces on a 24-hr cycle; that is, the retorts were charged in the early morning of one day, and in the early morning of the following day the residues were cleaned out, the retorts recharged, and a new cycle started. During the business depression of the early thirties, zinc production fell off to the extent that zinc smelters had idle retort capacity. Rather than continue to run some of the retort furnaces on a 24-hr cycle and shut down the remainder, a few companies decided to try out a longer cycle. Although the output of zinc per retort per day is lower, the longer cycle permits a larger charge per retort and improved zinc recovery. Advantages of the longer cycle were soon evident, and the practice spread rapidly. Various lengths of cycle were used, typical examples being 28-hr, 32-hr, 36-hr, and 48-hr. The 48-hr cycle fits in best with the standard routine of smelter practice that was built up on the basis of the 24-hr cycle; that is, the retort furnace "maneuver" always comes at the same time of day, and can be performed in the cool hours of the early morning. Under some conditions, particularly in some of the older plants, where the retorts are smaller than in newer plants, a 48-hr cycle is longer than necessary, and some intermediate length of cycle may theoretically be most economical. Use of an intermediate cycle, however, means that the maneuver comes at different times on consecutive days, and that sometimes it comes during the hottest part of the day. The work on the retort furnaces is extremely hot and arduous at best, and it is much worse, of course, if it is done during the heat of a summer day. When labor was plentiful and tractable, many plants were able to operate on cycles of intermediate length, but the workmen did not like the working hours. Under present labor conditions, it is hardly practicable to operate with a cycle of intermediate length, so the choice has narrowed down to the alternatives of 24-hr or 48-hr cycles. In any discussion of economics, too comprehensive generalization is not warranted, because local conditions affecting economics vary greatly. The best economic practice at one plant at any given time may be uneconomical at another plant, or even at the same plant at another time. To determine the length of cycle that is the better practice at a given plant at a given time requires a weighing of the advantages and disadvantages of each cycle, taking into consideration all the local conditions. Some of the factors that must be considered in making the comparison are discussed briefly in the following paragraphs.

Jan 1, 1947

By T. L. Joseph

Blast furnaces are most efficient thermally when the CO2 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyeres, is converted later to CO2 depends upon the ratio of ore to coke in the charge and the degree to which CO2, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + CO2 = 2CO. About 300 to 400 lb. of carbon' per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower CO2 in the top gas. Rich2 has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher CO2 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen entering electrolytic iron, melted under various slags, was directly proportional to the concentration of free FeO in the slag and inversely proportional to the basicity of the slag. Herty and Gaines5 concluded that the silicate content of open-hearth steel before deoxidation is related to the concentration of silica in the pig-iron charge, indicating that most of the persistent inclusions are silica from the iron. They also found that

Jan 1, 1936

By T. L. Joseph

Blast furnaces are most efficient thermally when the CO2 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyeres, is converted later to CO2 depends upon the ratio of ore to coke in the charge and the degree to which CO2, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + CO2 = 2CO. About 300 to 400 lb. of carbon' per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower CO2 in the top gas. Rich2 has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher CO2 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen entering electrolytic iron, melted under various slags, was directly proportional to the concentration of free FeO in the slag and inversely proportional to the basicity of the slag. Herty and Gaines5 concluded that the silicate content of open-hearth steel before deoxidation is related to the concentration of silica in the pig-iron charge, indicating that most of the persistent inclusions are silica from the iron. They also found that

Jan 1, 1936

By J. W. Leonard, J. C. Anderson, C. T. Holland

PREPARATION CONTROL IN UNDERGROUND MINING Selective Mining According to district Depending upon the particular geographic location or seam characteristics modem raw coal preparation practices can be used to obtain improved product quality and uniformity. Nevertheless, certain coal markets may require only marginal quality while others may require very high quality and these objectives may be sought by commencing the preparation of coal underground at the "working face." Underground preparation involves choosing those portions of the bed which can be cleaned to satisfy predetermined market requirements involving (1 ) maintenance of uniformity in chemical quality, (2) elimination of removable impurities, (3) control of sizes and size ratios and (4) control of moisture. The methods employed to attain these objectives vary widely with the (1) seam under development, (2) mining equipment available and (3) changing trends in the consuming markets. When a mining property is opened and development progresses from one section to another the (1) petrography of the coal in the bed does not re- main uniform, (2) character of roof and floor are subject to change and (3) character and prevalence of mineral impurities are subject to considerable variation. The quality of coal shipped to market from any mine depends upon intelligent foresight, planning, execution and supervision. Mining becomes a highly technical art. The quality of coal as determined by chemical and petrographic analysis is without question one of the most important considerations in the development and life of any coal-mining property. Variations in such proper- ties as volatile matter, ash, sulfur, Btu content and petrography are of considerable interest and to a limited extent so are such properties as phosphorous content, free swelling index, plasticity, hardness, pressure formation during coking and ash composition. The maximum and minimum specifications of these characteristics as established by competitive market conditions place definite limits on the marketability of a coal. The prospecting and evaluation of any area containing one or more salable coal beds by means of channel samples taken from outcrop openings and/or from diamond-drill cores can yield valuable data on coal characteristics. By correlating the data developed from distinct petrographic sections of channel and/or drill core samples, the quality of the bed according to stratification will be evident. There are few coal beds that do not show wide variations of such characteristics as ash and sulfur

Jan 1, 1968

By E. A. Gulbransen, J. W. Hickman

INTRODUCTION DURING the past two decades considerable progress has been made in the art of manufacturing heater alloys. The conventional iron-chromium-nickel alloys have been improved by the addition of small quantities of other elements. Up to the present time, to our knowledge, there has not appeared a systematic study of these alloys where the newer tools of investigation have been brought to bear on the problem. It is not yet clear what roles the additional elements play in increasing life time in usage. As a first approximation one might guess that elements such as manganese, silicon, zirconium, calcium and aluminum would aid in the formation of a protective film on the alloy surface, thus retarding the oxidation rate and increasing lifetime. If these additional elements do form protective oxides in the outer surface of the oxide film, it may be possible to identify them by means of reflection electron diffraction. It is also possible that the additional elements may not diffuse to the surface but may remain at the interface between the oxide and the metal. Then, too, the major elements in the alloys undoubtedly form oxides which are stratified on the surface. If the minor elements concentrate in a layer in contact with the substrate where they undergo oxidation and if this oxide film offers resistance to the diffusion of iron, chromium and nickel ions, then the lifetime of the heater element might be expected to increase. It is also possible that preferential reduction in the diffusion rates of the several ions may occur to such a degree that oxides may form in the body of the film and on the outer surface which are different from those that would normally form if the additional elements were not present. The use of transmisson electron diffraction may aid us in interpreting these phenomena. Up to the present time all of the factors which influence the lifetime of a given heater element are not definitely known. Among these are: (I) The rate of oxidation of the alloy. This may be dependent upon the composition and the physical state of the alloy. (2) The stability of the structure of the coating which forms. Transitions from one structure to another or solid phase reactions resulting in a change of composition may occur. The tendency of the film to adhere to the substrate when heated and cooled successively will be important. (3) The evaporation of various constituents from the alloy such as the removal of carbon by oxygen in a decarburization reaction. With these possibilities in view, an electron diffraction study of the oxides occurring on six different nickel-chromium alloys was undertaken with the hope that the results of this study could be correlated with oxidation rate studies as determined by means of the vacuum microbalance method1 and with

Jan 1, 1948

By J. B. Warriner

Stripping is the name given to the process of removing clay, rock, or other cover from deposits of coal or ore. In this paper it is intended to cover the methods used in carrying on this operation in the anthracite coal regions of Pennsylvania. In doing this, attention will be chiefly devoted to those methods wherein practice in different sections of the region differs, or is susceptible to change or improvement, though it is intended to give as complete data as possible concerning all stages of the operation. The data from which this paper has been prepared have been collected and submitted by a committee† representing the larger users of stripping methods, of which committee the writer has acted as chairman. Data and assistance have also been obtained from various mining men, engineers, contractors and manufacturers of stripping equipment, to all of whom our appreciation is due. Early History and Statistics The earliest mining operation on a commercial scale in the anthracite region was a stripping operation. This was the famous Quarry mine at Summit Hill, where this method was used only because no other method of mining a deposit of the nature found there seemed feasible to the early operators. A description of this stripping from a report published in 1821 is as follows: "The coal mine at present worked by the company lies on the top of a mountain, and appears to extend over some hundred acres of land covered by about 12 ft,. of loose black dirt resembling moist gunpowder, which can be removed by cattle with scrapers, and thrown into the valley below so as never to impede the work. The thickness of the coal is not known, but a shaft has been sunk in it 35 ft. without penetrating through. More than an acre of mine has been uncovered and presents a huge rock of coal, which is easily quarried without blasting."

Jan 1, 1918

By B. M. Harra, F. G. McCutcheon, O&apos

Until about 1930, the universal practice of horizontal-retort zinc smelters in the United States, as far as the writers are aware, was to operate the retort furnaces on a 24-hr cycle; that is, the retorts were charged in the early morning of one day, and in the early morning of the following day the residues were cleaned out, the retorts recharged, and a new cycle started. During the business depression of the early thirties, zinc production fell off to the extent that zinc smelters had idle retort capacity. Rather than continue to run some of the retort furnaces on a 24-hr cycle and shut down the remainder, a few companies decided to try out a longer cycle. Although the output of zinc per retort per day is lower, the longer cycle permits a larger charge per retort and improved zinc recovery. Advantages of the longer cycle were soon evident, and the practice spread rapidly. Various lengths of cycle were used, typical examples being 28-hr, 32-hr, 36-hr, and 48-hr. The 48-hr cycle fits in best with the standard routine of smelter practice that was built up on the basis of the 24-hr cycle; that is, the retort furnace "maneuver" always comes at the same time of day, and can be performed in the cool hours of the early morning. Under some conditions, particularly in some of the older plants, where the retorts are smaller than in newer plants, a 48-hr cycle is longer than necessary, and some intermediate length of cycle may theoretically be most economical. Use of an intermediate cycle, however, means that the maneuver comes at different times on consecutive days, and that sometimes it comes during the hottest part of the day. The work on the retort furnaces is extremely hot and arduous at best, and it is much worse, of course, if it is done during the heat of a summer day. When labor was plentiful and tractable, many plants were able to operate on cycles of intermediate length, but the workmen did not like the working hours. Under present labor conditions, it is hardly practicable to operate with a cycle of intermediate length, so the choice has narrowed down to the alternatives of 24-hr or 48-hr cycles. In any discussion of economics, too comprehensive generalization is not warranted, because local conditions affecting economics vary greatly. The best economic practice at one plant at any given time may be uneconomical at another plant, or even at the same plant at another time. To determine the length of cycle that is the better practice at a given plant at a given time requires a weighing of the advantages and disadvantages of each cycle, taking into consideration all the local conditions. some of the factors that must be considered in making the comparison are discussed briefly in the following paragraphs.

Jan 1, 1949

By Robert B. Bossler, Parke A. Dickey

The presence of connate water in oil sands is of far greater practical significance in secondary oil-recovery operations than it is in primary operations. The percentage saturations of oil, water, and gas in a depleted field determine whether it will be economically possible to recover oil by secondary methods, and whether water-flooding or gas-driving, or either, will be successful. It is suggested that economic recovery of oil by secondary water-flooding can occur only when a rich bank of oil is formed by the encroachingm water. If the sand has a greater effective permeability to water than it has to oil before flooding, a bank will not be formed, and water-flooding will not be successful, although gas-driving may be. Published results of laboratory work indicate that the effective permeability of a sand to oil and water depends primarily on the saturation of the sand by these phases. They also indicate that a small difference in relative saturation produces a large difference in effective permeability to the phase in question. There thus exists a critical range of saturations, wherein a small change in oil and water content of the sands will make water-flooding possible or impossible. In gas-drive operations the water saturation is less critical but no less important. The presence of connate water in a sand increases the proportion of the residual oil that may be recovered by gas drive. Sands containing oil saturations too low and water saturations too high to make water-flooding possible can often be subjected to gas drive with economic recovery of oil. The results of laboratory work on test samples that do not contain water cannot be applied to normal field operations. More tests on cores containing water as well as oil are desirable, and the improvement of coring technique is necessary. However, sufficient data are already available to enable an operator to use core data intelligently in estimating the chances of a secondary oil-recovery program. Introduction In recent years laboratory work has led to a much better understanding of the principles of oil-reservoir performance. The general principles of the expulsion of oil from sandstone by gas or water were developed by geologists of the Second Geological Survey of Pennsylvania in the 1870lsl but only recently has the mechanism of gas or water drive been intensively investigated. Most published articles have large the facts established in the labora tory to the primary production of oil from flush fields- This Paper is not based on any new observations, but is an attempt to apply some of the conclusions obtained in the laboratory and described in the literature to the recovery of oil from depleted pools by secondary methods. One of the most important factors that determine whether or not a secondary-recovery operation is economically successful is the ratio of injected fluid to oil recovered. This ratio may be expressed as the over-all or cumulative gas-oil ratio if gas is the driving medium, or water-oil ratio if water is the driving medium. Recent theoretical and experimental work has shown that the relative amounts of

Jan 1, 1944