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By Wilber Stout

In Ohio, the annual output of coke made from native coals has averaged not more than 70,000 tons, or about enough to run a 200-ton blast furnace. Raw coal locally mined from the Sharon, or No. 1, bed was used for iron smelting in Mahoning furnace at Lowellville, Mahoning Co., as early as 1846. This practice was continued in the Youngstown district until the coal became scarce and the Connellsville coke more popular. At present, the Sharon bed contributes a part of the fuel for the production of ferrosilicon iron at Jackson, Jackson CO. About one-third native coal and two-thirds soft coal are used. In this state, coke, on a commercial basis, was first made in Jefferson County probably about 1861 or 1862. The coals used were the Steuben-ville Shaft, or Lower Freeport, and the Strip Vein or Middle Kittanning. The practice, however, was discontinued before 1880 because of the competition of superior foreign cokes. Beehive ovens, later built at Leetonia, Columbiana Co., were operated for only a few years; the coal used at this place is the Lower Kittanning or No. 5. In a desultory way, slack from the Pittsburgh coal has been coked at a few places in eastern Ohio, and attempts, either for experimental or for commercial purposes, have been made in other parts of the state. Small quantities of coke were made from the Upper Freeport coal at Salineville, Columbiana Co., and near Nelsonville, Hocking Co.; the results were poor. Tests were made of the Middle Kittanning coal near Zanesville, Muskingum Co., at Moxahala Furnace, Perry Co., and at Washington Furnace, Lawrence Co.; the cokes were all high in sulfur and ash. Attempts were made to coke the Clarion coal at Vinton Furnace, Vinton Co.; a small battery of molding ovens and a pile of high-sulfur coke yet remain to tell the story. The Pittsburgh coal was coked on a commercial scale at Utley and at Lathrop, Athens Co., but the operations were not successful and were soon abandoned. These failures and indifferent successes are not to be attributed to a lack of demand for a first-class coke, as Ohio has held a commanding place for many years in iron smelting; nor to a deficiency of coals, as the state has a vast reserve of such material. They are to be ascribed to the inferior quality of the native products offered to the trade. The cokes

Jan 1, 1920

By Augustus J. Bowie

THE discovery of the auriferous cement gravel deposits in and near Deadwood Gulch, Lawrence County, Dakota Territory, in 1876, created a mining excitement, and rush to the Black Hills. Numerous ten and twenty-stamp mills were built on Deadwood Gulch ; Central City and the town of Deadwood sprang into existence, and became the centres of the mining industry of the Hills. Gold quartz deposits were subsequently discovered and located on a so-called mineral belt, extending from Whitewood Creek on the south, to Deadwood Gulch on the north. A northerly group of these quartz locations was purchased in December, 1877, by parties in San Francisco,, who formed a company, incorporated under the name of the Father de Smet Consolidated Gold Mining Company. This company immediately commenced the work of exploration. The deposits were proved to be extensive, and easily mined, but it was found that the ore, on an average, would not. yield more than $10.00 per ton* of 2000 pounds. At this time the cost of supplies and materials in the Hills was high. Miners' wages were $3.50 per diem, and custom mills charged $6:00. per ton. The successful working of mines on "the belt" necessitated cheap milling. The ores being of so low grade,† this could be accomplished only with large and economically running mills. To insure the greatest possible economy, the Father,& Smet mill * Up to December 31st, 1879; the de Smet mine had produced bullion of the value of $718,211.68 from 76,155 tons of ore, being an average value of $9.41 per ton. During the year ending December 31st, 1879, the mine produced 46,344 tons, which yielded $425,360.42, or $9.18 per ton. The total gross cost of working for the corresponding period, including deadwork, exchange, legal expenses, fete., wits $219,914.82. The yield of the mine for the year 1880 (Engineering and Mining Journal, February 5th, 1881), was $600,011, with total expenses of mining and milling, of $313,108. † The first 4083 tons extracted from the de Smet mine, were selected, yielding $62,783.97. The next 25,728 tons produced $230,067.24. The total 29,811 tons were weighed, and worked at custom mills. The yield of the first 409 tons from tine Golden Terra Mine was $11.44 per-ton.

Jan 1, 1882

By S. B. Patterson, A.R. Amos

Log washers have been used for many years in the washing of clay iron ores, phosphate rock and manganese ores, but not until the past 15 years have they been employed to any extent in the preparation of aggregate materials and fluxing stone. For a machine of such importance where tenacious clay or soft rock must be eliminated, there are few data generally available. The purpose of this paper is to set forth the results obtained from several operations, together with a few suggestions for planning a specific installation. Where the material to be washed out consists of loose surface soil, silt, or small amounts of quickly soluble clay, there is little competition with other forms of washers, because greater power and maintenance are required for the operation of log washers, but where the foreign matter is a stiff clay, such as occurs with high-calcium limestones and in marginal gravel deposits, the ability of other kinds of washers to eliminate completely the hard clay balls is at least uncertain. Then the log washer comes into its own field, for with the breaking and churning action of the logs it will clean out the toughest clay. Even with material from the hydraulic dredge, if it contains tough clay, the log washer is desirable unless the gravel is subjected to a costly soaking and rehandling process. Recently there has been a considerable development in other types of machines to accomplish these results but it is not within the scope of this paper to make comparisons between competitive types, but to present the best obtainable data on log washers only. From the primitive design that gave it its name, a long log hewn from a tree, with paddles forged to shape from strap iron and attached by lag screws, the log washer has gradually been developed to the present all-steel type. As now manufactured, the log washer consists essentially of two built-up logs on which are mounted a series of paddles, carried by shafts fitted at the ends of the logs on bearings fastened outside the enclosing steel box, or trough, which is set at an angle of about 5" to 10" and into which the material is fed. Stuffing boxes at the places where the shafts pass through the box at the lower end are equipped with water-

Jan 1, 1938

By A. R. Amos, S. B. Patterson

Log washers have been used for many years in the washing of clay iron ores, phosphate rock and manganese ores, but not until the past 15 years have they been employed to any extent in the preparation of aggregate materials and fluxing stone. For a machine of such importance where tenacious clay or soft rock must be eliminated, there are few data generally available. The purpose of this paper is to set forth the results obtained from several operations, together with a few suggestions for planning a specific installation. Where the material to be washed out consists of loose surface soil, silt, or small amounts of quickly soluble clay, there is little competition with other forms of washers, because greater power and maintenance are required for the operation of log washers, but where the foreign matter is a stiff clay, such as occurs with high-calcium limestones and in marginal gravel deposits, the ability of other kinds of washers to eliminate completely the hard clay balls is at least uncertain. Then the log washer comes into its own field, for with the breaking and churning action of the logs it will clean out the toughest clay. Even with material from the hydraulic dredge, if it contains tough clay, the log washer is desirable unless the gravel is subjected to a costly soaking and rehandling process. Recently there has been a considerable development in other types of machines to accomplish these results but it is not within the scope of this paper to make comparisons between competitive types, but to present the best obtainable data on log washers only. From the primitive design that gave it its name, a long log hewn from a tree, with paddles forged to shape from strap iron and attached by lag screws, the log washer has gradually been developed to the present all-steel type. As now manufactured, the log washer consists essentially of two built-up logs on which are mounted a series of paddles, carried by shafts fitted at the ends of the logs on bearings fastened outside the enclosing steel box, or trough, which is set at an angle of about 5" to 10" and into which the material is fed. Stuffing boxes at the places where the shafts pass through the box at the lower end are equipped with water-

Jan 1, 1938

By W. Edeburn, H. R. Fife

An extended study of dustproofing Freeport-seam coal has emphasized several fundamental features in the successful treating of bituminous coals. The extent of the exposed surface area and its absorption propcrties for the dust-retaining film were found to be the most critical factors affecting the efficiency of fly-dust prevention. From recently published screen data and specific surface-area measurements on United Kingdom coals, an approximation of the surface exposed by mine-run coal and by 1-in. slack was made. The total surface area exposed by an average tori of this mine-run coal was estimated at 3768 sq. yd., while a ton of 1-in. slack represents a total surface area of 8779 sq. yd. with no additions for porosity. It is important in coal treating to consider that in mine-run coal the fraction above 1/4 in. in size represents 79.5 per cent of the total weight, but only 4 per cent of the total surface, while the fraction below 150 mesh contains 47 per cent of the total surface and only 1.2 per cent of the total weight. Also, one should appreciate that, in a 1-in. slack, the domestic stoker fraction (1 in. to 8 mesh) contains 76.3 per cent of the weight and 8.2 per cent of the surface, while the fraction below 150 mesh represents 3.3 per cent by weight and 54.8 per cent by surface area. Wax extraction data on 1-in. to 0 slack from mine-run treated Freeport-searn coal showed that the coating was distributed uniformly as a function of exposed surface; hence, nonuniformly on a weight basis. This distribution of the coating on mine-run coal therefore limits the extent of treating that is desirable on bituminous coal at this point in the preparation. The treatment of Freeport mine-run coal before it reached the tipple gave a 70 to 80 per cent reduction of fly-dust in the tipple, a satisfactory treated commercial slack coal, and an indicated change in the trend of tipple degradation. The premium-treated domestic sizes required a supplemental treatment for complete dustproofness. The actual cost of the pretipple and supplemental treatments was very low per ton of coal treated, while the net or true cost was sufficiently

Jan 1, 1938

By W. Edeburn, H. R. Fife

An extended study of dustproofing Freeport-seam coal has emphasized several fundamental features in the successful treating of bituminous coals. The extent of the exposed surface area and its absorption propcrties for the dust-retaining film were found to be the most critical factors affecting the efficiency of fly-dust prevention. From recently published screen data and specific surface-area measurements on United Kingdom coals, an approximation of the surface exposed by mine-run coal and by 1-in. slack was made. The total surface area exposed by an average tori of this mine-run coal was estimated at 3768 sq. yd., while a ton of 1-in. slack represents a total surface area of 8779 sq. yd. with no additions for porosity. It is important in coal treating to consider that in mine-run coal the fraction above 1/4 in. in size represents 79.5 per cent of the total weight, but only 4 per cent of the total surface, while the fraction below 150 mesh contains 47 per cent of the total surface and only 1.2 per cent of the total weight. Also, one should appreciate that, in a 1-in. slack, the domestic stoker fraction (1 in. to 8 mesh) contains 76.3 per cent of the weight and 8.2 per cent of the surface, while the fraction below 150 mesh represents 3.3 per cent by weight and 54.8 per cent by surface area. Wax extraction data on 1-in. to 0 slack from mine-run treated Freeport-searn coal showed that the coating was distributed uniformly as a function of exposed surface; hence, nonuniformly on a weight basis. This distribution of the coating on mine-run coal therefore limits the extent of treating that is desirable on bituminous coal at this point in the preparation. The treatment of Freeport mine-run coal before it reached the tipple gave a 70 to 80 per cent reduction of fly-dust in the tipple, a satisfactory treated commercial slack coal, and an indicated change in the trend of tipple degradation. The premium-treated domestic sizes required a supplemental treatment for complete dustproofness. The actual cost of the pretipple and supplemental treatments was very low per ton of coal treated, while the net or true cost was sufficiently

Jan 1, 1938

By Raymond J. Kujawa

For general purposes, insulating materials may be any of those mineral substances that provide a barrier between a desired human environmental feature and an unwanted condition. In this sense, radiation shields, electrical insulations, cryogenic insulation, and acoustical barriers would qualify. For this chapter only the thermal and acoustic insulations used for the median range of normal comfort control will be covered. Because of their magnitude and complexity the other fields are left to other chapters. In another sense, this section deals only with the ultra-lightweight aggregates and insulations, rather than the multitude of light-weight structural materials. Definitions and Uses The insulating materials covered in this chapter are the three groups of mineral products that provide thermal or acoustic barriers through the median temperatures encountered in home and industry. The three mineral categories are perlite, vermiculite, and the silicate wools whether they be categorized as glass, mineral, or rock. Perlites are those amorphous aluminum silicates with minor impurities that take the form of a felsitic, glassy lava possessing a pearl-like luster and concentric cracking. In a more commercial sense the term is used to describe either the natural or processed acidic, volcanic glass, containing from 2 to 5% combined water, that may be converted to a frothy mass by heating rapidly to the softening point. Vermiculites are micaceous, hydrated magnesium iron aluminum sheet silicates of varying composition. Commercial vermiculites are any of these sheet silicates that are capable of sufficient expansion on rapid heating to make an acceptable, highly porous product. Mineral wools, rock wools, silicate cottons, glass fiber are generic names for manufactured fibers in which the fiber forming substance may be slag, certain rocks, or glass. The fibers may be made by a blowing or spinning process. The resultant fibers may be felted or matted or gathered in the manner best suited to the use. The three groups may be used competitively to insulate ceilings and walls. The fibers have strong advantages in this area as they may be manufactured in batts that are stable in residential wall construction. The perlites and vermiculites are more suited to insulate cores and cavities of masonry walls and as plaster and concrete aggregates for thermal insulations. Vermiculite is widely used for insulating attic floors. Because of manufacturing techniques and controls, the glass fibers have made great strides as composites in the plastics industry but only a few of these uses classify in the insulation category.

Jan 1, 1975

By A. Bar-Or, G. Hirsch, P. Wynblatt

The rates of nucleation and growth at various temperatures in U-0.45, 1.1, and 1.35 at. pet Cr alloys were determined, providing a qliantitative proo-f that the transformation under study is controlled by nucleation and diffusional growth. The linear relationship, log NV US 1/T2, suggests that the nucleation of the alloys is probably controlled by the classical law of nuclei formation. The established linear relationship, log G us 1/T, in the three alloys suggests that growth in each alloy is controlled by bidimensional nuclei formation. It was established that both the rate of growth and the rale of nucleation are retarded by chromium. Using classical formulae, sApN and sApG (the ß/a apparent interphase energies) were calculated from the nucleation and growth results, respectively. It was found that their values are of the same order of magnitude. However, why sApN is unaffected by the addition of chromium, while sApG increases with it, needs further clarification. In the last few years, an appreciable amount of research has been conducted into the mechanism of the ß —a transformation in uranium and uranium alloys. Recently, Burke1 reviewed the published work concerned with this transformation. He concluded that in dilute uranium alloys it is firmly established that, at high temperatures in the a range, ß decomposes by a diffusional process. However, he based his conclusion either on experiments dealing with the over-all transformation,2 or on qualitative results.3 It was felt that a better insight into the mechanism of transformation could be gained from the quantitative measurements of growth and nucleation rates, which we aimed to carry out in the present work. In a previous work,4 we have shown that a rim of a grains is formed during isothermal transformation at the high a range. The thickening of this rim with time was used for measuring the rate of growth in U-0.45 at. pet Cr.5 This method was also used in the present work for two more alloys. For meas- uring the rate of nucleation, a method of quenching and macroetching was developed to allow differentiation between a grains formed during isothermal transformation at the upper a range and those formed after quenching, Fig. 1. EXPERIMENTAL WORK AND RESULTS The Rate of Nucleation. The measurement of the rate of nucleation was effected by counting the num-

Jan 1, 1965

By L. Zwell, H. A. Wriedt

In describing their procedure for measuring the lattice parameters of solid solutions, Abrahamson and Lopata report (p. 77) the final heat treatment of their alloy specimens as "700°C for 100 hr, followed by a furnace cool". The isothermal part of this treatment may be expected to precipitate solute in excess of the stable a solid-solution limit at 700°C; the subsequent slow cooling is likely to reduce the dissolved solute still further. Fig. 1 (upper) presents the data for titanium alloys so he;rt-treated and shows the lattice parameters of solid solutions containing up to 3.5 at. pct Ti; Fig. 17 (lower) shows 2 at. pct Ti as the solubility limit in a iron at 700°C. Figs. 1 and 17 are thus incompatible. The paper contains comparable incon~patibilities for zirconium alloys, for niobium alloys, and for tantalum alloys. The final heat treatment at 700°C may also be responsible for the anomalous "knee" in the lattice-parameter data for nickel alloys, Fig. 3. The "knee" composition is located near the 700°C limit of the stable a solid-solution field;37 alloys of the experimental composition range with more nickel than the "knee" composition are two-phase at 700 C and prolonged heat treatment of them at 700°C would tend to produce inhomogeneity rather than homogeneity. In Figs. 4 and 7, the composition ranges of the zirconium, hafnium, and niobium alloys examined terminate at alloy concentrations of 0.1 to 0.2 at. pct. These metals form very stable oxides and their partial removal from solid solution in iron by impurity oxygen may be expected. We believe that the authors should have reported the procedure by which they distinguished the amounts of alloying elements in solution and in oxide. The diagrams pertaining to copper in Figs. 3 and 18 (lower) are almost certainly incorrect. Three recent investigations of this system38-40 all substantially agree on the solubility of copper in a iron, finding it to be much larger than that reported by Abrahamson and Lopata. The change of the lattice parameter with composition reported by Abrahamson and Lopata is too large by a factor of three to five, as seen by comparison with the findings of other investigators.41738 Moreover, the authors violate, in Fig. 18, the accepted Fe-Cu constitution diagram37 by ignoring the eutectoid at about 850°C.

Jan 1, 1967

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

By W. A. Binsacca

The concentrator treats sulfide ore containing lead, copper, zinc, gold and silver, at the rate of 850 tons per 24 hr., and produces lead, copper, zinc and iron concentrates by selective flotation methods. The ore occurs below the oxidized zone along a main fracture, the wall rock of which is either greywacke or calcareous shale. The predominating sulfide minerals are galena, sphalerite, pyrite and chalcopyrite; the others are quartz and calcite. In minor amounts, arsenopyrite, pyrrhotite and hematite occur, and such secondary minerals as smithsonite, chalcocite, covellite, bornite, cerussite and anglesite are detected in the ore, the last two named occurring principally as surface alterations of the galena. The ore is hauled from the mine bins in 10-ton cars which pass over the track scales before entering the car dumper ahead of the crusher plant.' A conveyor belt delivers the crushed ore to the sulfide bin, which has a capacity of 2800 tons. The primary grinding unit consists of two 6 by 12-ft. Marcy rod mills operating in parallel and in open circuit, each taking 425 tons per day. One rod mill discharges into an 8 by 25-ft. Dorr classifier with a 12-ft. bowl and the other into a 6 by 25-ft. Dorr duplex classifier. These classifiers are arranged so that each unit rakes the sands into a 6 by 14-ft. Traylor ball mill, which discharges to the bowl classifier only. The duplex classifier therefore operates in open circuit and the bowl classifier in closed circuit, the overflow of both classifiers furnishing the feed for the flotation section. Other groupings have been tried, but this has been found the most satisfactory. The Marcy mills are driven by 150-hp. slip-ring induction motors with speed reducers to give 17 r.p.m. The Traylor mill operates at a speed of 24 r.p.m. and is driven by a 200-hp. slip-ring induction motor, direct connected to pinion and gear. The bowl classifier operates at a speed of 22 r.p.m., with the bowl rake making 4.5 r.p.m. and the duplex classifier at 31 r.p.m. High-carbon steel rods of Mexican manufacture, 12 ft. long by 3-in. diameter, are used in the Marcy mills, with a resulting consumption of 1.52 Ib. per ton ore. The Traylor mill is fed 3-in. cast-iron balls cast in

Jan 1, 1935

By Eugene Jablonski

Statistical data on the Polish producing business are available only for the first half of 1939. Because of the invasion of the country during September by Germany and Russia all sources of information have stopped, and it is impossible to get even an approximate picture of the status of the industry after that date. The long declining trend of Polish production, which was finally checked during 1938, kept a further upward tendency during the first half of 1939, and there is good reason to believe that the upward trend would have been maintained. It is true, there was little hope of reviving the slowly dying Boryslaw field, although there still remained a slight possibility for an extension in its southwestern direction. Repressuring plans were still discussed, but. could not be put into effect because of extremely complicated and split-up lease ownership. The Gorlice-Lipinki ficld continued to be the main factor in the increase of the country's production. During the first half of 1939 this field received a further important extension in a southern direction. The Schodnica field continued to use repressuring with success. At the end of June, 21 exploratory wells were drilling, 14 of which were within thc Carpathians and 7 in the Miocene geosyncline of the Carpathian Foreland. All of the wildcat wells drilled within the Carpathians were on the struct,ural cxtcnsion of old established fields, but part of the wildcats of the Foreland were on new geological and geophysical prospects. The Opary gas field of the Foreland discovered last year has been further developed. The Government-owned Polmin completed its third gasser. This well penetrated two sands; the upper one, 118 ft. thick, was found at 1312 ft., with an initial pressure of 570 lb. and capable of produeiring 10 million cu. ft. per day. The second sand, 91 ft. thick, was found at 1690 ft., yielding 20 million cu. ft., with a pressure of 760 lb. The well Mieczyslaw, drilled about 3 miles further north on the flank of the structurc, found the first sand watered. The Pioneer No. 5, at Chodonice near Przemysl, which reported somc gas last year, was 5696 ft. deep at the end of June, still drilling in Miocene. Anothe well in the same vicinity drilled by the Gasolina was 1708ft. deep.

Jan 1, 1940

By Eugene Jablonski

Statistical data on the Polish producing business are available only for the first half of 1939. Because of the invasion of the country during September by Germany and Russia all sources of information have stopped, and it is impossible to get even an approximate picture of the status of the industry after that date. The long declining trend of Polish production, which was finally checked during 1938, kept a further upward tendency during the first half of 1939, and there is good reason to believe that the upward trend would have been maintained. It is true, there was little hope of reviving the slowly dying Boryslaw field, although there still remained a slight possibility for an extension in its southwestern direction. Repressuring plans were still discussed, but. could not be put into effect because of extremely complicated and split-up lease ownership. The Gorlice-Lipinki ficld continued to be the main factor in the increase of the country's production. During the first half of 1939 this field received a further important extension in a southern direction. The Schodnica field continued to use repressuring with success. At the end of June, 21 exploratory wells were drilling, 14 of which were within thc Carpathians and 7 in the Miocene geosyncline of the Carpathian Foreland. All of the wildcat wells drilled within the Carpathians were on the struct,ural cxtcnsion of old established fields, but part of the wildcats of the Foreland were on new geological and geophysical prospects. The Opary gas field of the Foreland discovered last year has been further developed. The Government-owned Polmin completed its third gasser. This well penetrated two sands; the upper one, 118 ft. thick, was found at 1312 ft., with an initial pressure of 570 lb. and capable of produeiring 10 million cu. ft. per day. The second sand, 91 ft. thick, was found at 1690 ft., yielding 20 million cu. ft., with a pressure of 760 lb. The well Mieczyslaw, drilled about 3 miles further north on the flank of the structurc, found the first sand watered. The Pioneer No. 5, at Chodonice near Przemysl, which reported somc gas last year, was 5696 ft. deep at the end of June, still drilling in Miocene. Anothe well in the same vicinity drilled by the Gasolina was 1708ft. deep.

Jan 1, 1940

NAME AND OBJECT. SEC. 1. This Institute is incorporated under the Membership Corporation Law of the State of New York ; its corporate name is AMERICAN INSTITUTE OF MINING ENGINEERS; and its objects are such as are stated in its Certificate of Incorporation. ARTICLE II. MEMBERS. SEC. 1. The membership of the Institute shall comprise four classes, namely (1) Members ; (2) Honorary Members , (3) Associates , and (4) Honorary Associates Only Members and Associates residing within the United States of America, Republic of Mexico and Dominion of Canada shall be entitled to vote at the meetings of the Institute. SEC. 2 All Members, Honorary Members, Associates and Honorary Associates of the American Institute of Mining Engineers as the same existed on the day of the incorporation of this Institute, are Members, Honorary Members, Associates and Honorary Associates, respectively, of this Corporation SEC. 3. The following classes of persons shall be eligible for membership in the Institute, namely, as Members and Honorary Members, all professional mining engineers, geologists, metallurgists or chemists, and all persons practically engaged in mining, metallurgy or metallurgical engineering ; as Associates and Honorary Associates, all persons desirous of being connected with the Institute who, in the opinion of the Council, are suitable. SEC 4. Every candidate for election as a Member or Associate of the Institute must be proposed for election by at least three Members or Associates; must be approved by the Committee on Membership, as prescribed in the By-Laws ; and .must be elected by the Council. Not less than three-fourths of the votes cast shall be necessary to an election Every person so elected shall become a Member or Associate, as the case may be, upon payment of his first dues as herein-after prescribed Each candidate for Honorary Member or Honorary Associate, must be recommended by at least ten Members or Associates ; must be approved by the Council ; and must be elected by ballot at a meeting of the Board of Directors by the unanimous vote of all the Directors present ; provided, however, that the number of Honorary Members and Honorary Associates shall not at any time exceed twenty SEC. 5. If any person elected a Member or Associate does not, within sixty days after notice of his election, accept the same and pay his initiation fee and dues for the current year, his election may be cancelled at the discretion of the Council.

Jan 1, 1910

By Lucien Eaton

The use of filling in mines is less common in the United States than it is in Europe, where in some places it is required by law. In most cases the filling is placed by hand, and the material used for this purpose is the waste from development and that sorted from the ore in the stopes. When these sources provide an insufficient supply, additional rock must be quarried on surface or mined underground, or some other material must be used. In different parts of the world sand, clay, gravel, culm, granulated slag, mill tailings, and even ashes, have been used as filling material in place of or in conjunction with rock. Of all the materials used for filling the best is probably clean sand. Gravel, finely crushed rock, granulated slag and coarse mill tailings are almost as good, and the methods used for placing and confining sand will apply equally well to any of these. The advantages of sand are its comparatively slight compressibility and the ease, economy and convenience with which it may be placed. On account of the voids, which necessarily occur in fills made of coarse waste, such fills will compress 10 to 25 per cent before they develop their full strength, whereas clean sand seldom compresses more than 5 per cent. Sand is especially desirable when practically complete filling is required, and sand derived from mill tailings, especially when not thoroughly cleaned, has been remarkably effective in sealing off and extinguishing mine fires. When natural sand is available on surface, or when it can be obtained from the mill tailings or tailings dump, the cost of using this material for fill will probably be considerably less than that of rock. In this paper it is not proposed to discuss all methods of filling, but to describe some of the methods used for placing sand fill and to make a brief comparison of their merits. Methods of Placing Sand Sand may be transported and placed underground by several methods, which may be classified as follows: 1. By Hand.—The sand, which must be comparatively dry, is taken down the shaft in cars on cages and is distributed to its destination in these cars, or it is dumped down raises and is drawn off and distributed by means of cars, wheelbarrows or scrapers.

Jan 1, 1932

By Raymond M. Howe

Some time ago, a prominent engineer asked a representative of the firebrick industry to prepare a comprehensive paper on blast-furnace refractories. It was to have been the purpose of this paper to gather practical experiences from widely different sources, in the hope of determining what kind or kinds of firebrick were best adapted for certain uses. This engineer sent questionnaires to the large consumers of blast-furrace linings and stovebrick and received very detailed replies as to their experiences. These replies are conflicting, but, having been turned over to the writer, form the basis of this attempt to meet the request. Because of the impossibility of making final recommendations at this time, a few of the more interesting and commonly known methods of studying refractories are also discussed, as well as the more recent developments in this field. The first reply to the questionnaire stated that good service was being secured from the linings. This satisfactory condition was attributed to the use of a very good distributor. The writer believed that manufacturers try to and do produce a good product and that the proper handling of equipment is of the utmost importance, if good results are to be secured. The type of bricks used was not mentioned. The second reply stated that good life was secured from the furnace linings, but that the writer had had soft-fired stovebrick crumble and, for that reason, preferred a hard-fired product for use in this position. The type of bricks used was not mentioned. The third reply stated that firebrick had crumbled from the mantle up; the cause was not known. The fourth reply stated that the stove linings were satisfactory. Hand-made furnace linings failed after.4 years' service so a steam-pressed lining- was installed which was still good after 4 years of service. Consequently, a second steam-pressed lining was installed but this disintegrated at the top after the furnace had been in blast 7 months. The firebrick manufacturer attributed this failure to the expansion of iron blocks that had been installed at the stock line to prevent erosion. The fifth reply stated that the present "hard-burned" hand-made furnace lining had worn away very uniformly to the depth of only 5 in.

Jan 1, 1920

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 N. S. Stoloff, R. G. Davis

It is concluded from an X-ray stztdy that the formation of the hcp Mg3Cd superlattice is a nuclea-tion and growth reaction. A two-phase, ordered-plus-disordered, region is observed between 153" nnd 141"C. A maximum in the flow stress is observed both at temperature and in quenched specimens within the two-phase region. The peak arises fro112 the extra energy required to create antiphase boundary upon cutting the ordered regions. Hardening during isothermal annealing in Mg3Cd has been shown to be similar to the domain hardening in Cu3Au-type alloys in that up to the maximum the structure is ordered domains growing into a short range ordered matrix. The hardening is due to a combination of short range order strengthening and the strengthening produced when antiphase boundary has to be created. In a recent study1 of the effect of long-range order upon the mechanical properties and deformation modes of hcp Mg3Cd a comparison was made be- tween quenched (disordered) and slow-cooled (fully ordered) specimens. The most striking observations were that the disordered alloy yields at a much higher stress than the ordered alloy, and that there is a change in slip mode from primarily (_0001) ba_sal slip in the disordered state to (0001), {1010}, {1122}, and {l011} slip in the ordered state. The main purpose of the present work is to extend these observations to determine whether MgsCd exhibits a peak in flow stress at an intermediate degree of order, or at a critical domain size, as do other orderedstems. However, disagreement exists about the Mg3Cd order transformation with respect to: 1) the location of the critical temperature for order, 165°C or 153"~,' and 2) how the degree of order varies with temperature. Welber et a1.5 concluded from specific heat measurements that the formation of Mg3Cd is a second-order reaction with the degree of order a continuous function of temperature, while Moore and Raynor concluded that it is neither a first-order reaction, with the degree of order a discontinuous function of temperature, nor a second-order reaction. Consequently it was considered necessary to undertake an X-ray investigation into the change in degree of order and structure in the

Jan 1, 1964

By H. D. Colson, Robert A. Rapp

When a binary alloy is oxidized to form simultaneously an internal oxidation zone and an external scale which grows at a constant rate, the oxidation should reach a condition of steady shale, in which the internal oxidation zone is maintained at a time-independent thickness. Assuming an ideal model for the steady-state oxidation, the diffusion equation is solved for the steudy-state concentration profiles for oxygen and alloying- element in the metal. A theoretical prediction of' the steady-state thickness of the internal oxidation zone is made in terms of the experimental parameters. FOR the oxidation of binary alloys, Rhines, Johnson, and Anderson,1 as well as Mask: have derived equations describing the kinetics for the simultaneous formation of an internal oxidation zone and an external scale which was assumed to mow according to a parabolic rate law. Experimental determinations of internal oxidation zone thicknesses beneath a Cu2O scale for Cu-Be and Cu-A1 alloys3,4 have shown some deviation from Maak's theoretical expressions, because the ideal model assumed for the calculation is not met experimentally. Some pure metals, such as columbium and tantalum oxidize at elevated temperatures and high oxygen pressures to form nonprotec-tive external scales which thicken at a constant rate. If such a base metal is alloyed with a solute element which may form a more stable oxide than the solvent oxide, then internal oxidation of solute may occur beneath the growing external scale. A review of the internal oxidation of alloys has been published recently.5 The purpose of this paper is to analyze theoretically for a binary alloy the kinetics for the simultaneous formation of an internal oxidation zone and an external scale which grows at a constant rate. Consider an alloy A-B in which A is the more-noble solvent and B is the less-noble solute. The solute oxide is assumed to precipitate in internal oxidation as stoichiometric particles of BO, which are assumed to have a negligible solubility product in the A matrix. It is further assumed that the diffusion of oxygen through the zone of internal oxidation is not impeded by previously precipitated BO, particles. The specimen is assumed to be sufficiently thick so that the solute content in the center of the specimen does not change. For consistency: the following notation of Maak2 has been used:

Jan 1, 1967