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LOCAL SECTIONS (There are 60 Local Sections in the United States, 4 Foreign ) ADIRONDACK Established June 18, 1952 Paul W Allen, Chairman William Webb, Vice-Chairman Edwin Briggs, Secretary-Treasurer National Lead Company Mac Intyre Development Tahawus, New York Paul W Allen, Section Delegate R G Fleck, Alternate Section Delegate

Jan 1, 1954

By W. R. Crane

Mining of the red iron ores of the Birmingham district has been carried on energetically during the past 60 years, and their development has created a large iron and steel manufacturing center, the only important one in the South. The district produces approximately 10 per cent. of all the iron ore of the United States (80 per cent. of the Alabama ore mined is red ore); also, 40 of the 400 blast furnaces of the country are in the district tributary to Birmingham. The rapid growth of the district has been made possible through investigations that resulted in radical changes in furnace practice, and a still greater impetus will come from the study of the low-grade, high-silica ores, as a result on which they will be made amenable to treatment by concentration. Mining practice in this district has been comparatively simple because of the occurrence of the ore; but with the rapid extension of the workings and the disturbed condition of the ore bed at some distance from the outcrop, more difficult conditions are encountered and the tendency is toward worse rather than better conditions. Should the high-silica ores of the lower bench of the Big Seam become available through beneficiation, mining practice will have to be modified to meet the new conditions, which will be rendered more difficult by the increased weight of cover that will exist at considerable distances from the outcrop. Support of workings will require greater attention and the efficient and economical operation of the mines will depend largely on the successful solution of the problems of working and handling the ore. History and EaRly Development The first explorers of the coal and mineral lands of Alabama were blacksmiths and mechanics mustered out of the army after the war of 1812. These men recognized the red rock of the Birmingham district as iron ore, and utilized it in making cooking utensils and farm implements. The first blast furnace was built and operated at Russellville in 1818, where also was a foundry and rolling mill. Soft ore was used at that time, but, in 1864, the first red hematite ore of Red Mountain was smelted, near Irondale. Birmingham was founded in 1871, and the successful use of coke in making pig iron in 1876 was the beginning of

Jan 1, 1925

By R. S. McClellan

In order to meet the ever-increasing demand by consumers for uniformity of ground talc, a new method of blending its finished product has been developed by Gouverneur Talc Company, Inc., at its plant near Gouverneur, N. Y. The method has been in use for a sufficient length of time to be now judged a success. The Gouverneur district is the leading talc producing area of the world, inasmuch as more tonnage is produced there than from any other state in the U.S. or, so far as is known, from any other locality in the world. Almost all of the output goes into industrial uses, with ceramics and paint being the principal consuming industries. The physical and chemical specifications imposed by these industries are stringent and the trend, as in most industries, is toward ever increasing severity. The blending method described herein has aided Gouverneur Talc to keep up with this trend.

Jan 3, 1961

By Ira B. Joralemon

Neighboring mines, like men who live in the same environment, often assume similar characteristics. In the Bridge River area of the Lillooet district in central British Columbia there is a remarkable exception to this rule. The Pioneer mine has a simple fissure over 3000 ft. long, with no faults of more than a few feet and with only minor variations in width; the Bralorne mine, 2½ miles away at the other end of the same intrusion, has developed one of the most intricate vein and fault systems in the West. Hinged faults displace the main vein for hundreds of feet, and sometimes become valuable veins themselves. The width of ore varies in a few yards from 2 to 60 ft. Changes in grade of ore are just as abrupt. The rocks that follow the veins or are cut by them often grade into one another or into the ore with no definite contacts. Although the complexity of the structure has left many problems unsolved, the development has indicated a fascinating geological pattern. The Bralorne is in the eastern portion of the Coast Range, 55 miles by spectacular mountain road from the Pacific Great Eastern Railway station of Bridge River. The adit tunnel runs easterly from Cadwallader Creek, a tributary of Bridge River that flows into the Fraser River. Steep mountains rise from 3300 ft. elevation above sea level at the creek to peaks and ridges at 8000 to 9000 ft. altitude. The lower slopes are heavily wooded while above 5000 ft. bare rocks, glaciers and snow-fields extend far above timber line. See Fig. 1. Story of the Bralorne The Lorne claim, staked in 1897, was one of the earliest lode locations in the Bridge River district. At that time it was 70 miles by steep and treacherous trail from the railway. For the next 15 years Arthur Noel, Will Haylmore and other local prospectors mined the small, irregular, oxidized quartz veins to a maximum depth of 200 ft., and treated the ore in arrastres and later a five-stamp mill. The total production to 1912 was only $55,000. Noel and associates later formed the Lorne Amalgamated Mines Co., which they worked in a desultory way until 1925. Increasing sulfide and arsenide content in the veins

Jan 1, 1935

By Ira B. Joralemon

Neighboring mines, like men who live in the same environment, often assume similar characteristics. In the Bridge River area of the Lillooet district in central British Columbia there is a remarkable exception to this rule. The Pioneer mine has a simple fissure over 3000 ft. long, with no faults of more than a few feet and with only minor variations in width; the Bralorne mine, 2½ miles away at the other end of the same intrusion, has developed one of the most intricate vein and fault systems in the West. Hinged faults displace the main vein for hundreds of feet, and sometimes become valuable veins themselves. The width of ore varies in a few yards from 2 to 60 ft. Changes in grade of ore are just as abrupt. The rocks that follow the veins or are cut by them often grade into one another or into the ore with no definite contacts. Although the complexity of the structure has left many problems unsolved, the development has indicated a fascinating geological pattern. The Bralorne is in the eastern portion of the Coast Range, 55 miles by spectacular mountain road from the Pacific Great Eastern Railway station of Bridge River. The adit tunnel runs easterly from Cadwallader Creek, a tributary of Bridge River that flows into the Fraser River. Steep mountains rise from 3300 ft. elevation above sea level at the creek to peaks and ridges at 8000 to 9000 ft. altitude. The lower slopes are heavily wooded while above 5000 ft. bare rocks, glaciers and snow-fields extend far above timber line. See Fig. 1. Story of the Bralorne The Lorne claim, staked in 1897, was one of the earliest lode locations in the Bridge River district. At that time it was 70 miles by steep and treacherous trail from the railway. For the next 15 years Arthur Noel, Will Haylmore and other local prospectors mined the small, irregular, oxidized quartz veins to a maximum depth of 200 ft., and treated the ore in arrastres and later a five-stamp mill. The total production to 1912 was only $55,000. Noel and associates later formed the Lorne Amalgamated Mines Co., which they worked in a desultory way until 1925. Increasing sulfide and arsenide content in the veins

Jan 1, 1935

Neighboring mines, like men who live in the same environment, often assume similar characteristics. In the Bridge River area of the Lillooet district in central British Columbia there is a remarkable exception to this rule. The Pioneer mine has a simple fissure over 3000 ft. long, with no faults of more than a few feet and with only minor variations in width; the Bralorne mine, 2½ miles away at the other end of the same intrusion, has developed one of the most intricate vein and fault systems in the West. Hinged faults displace the main vein for hundreds of feet, and sometimes become valuable veins themselves. The width of ore varies in a few yards from 2 to 60 ft. Changes in grade of ore are just as abrupt. The rocks that follow the veins or are cut by them often grade into one another or into the ore with no definite contacts. Although the complexity of the structure has left many problems unsolved, the development has indicated a fascinating geological pattern. The Bralorne is in the eastern portion of the Coast Range, 55 miles by spectacular mountain road from the Pacific Great Eastern Railway station of Bridge River. The adit tunnel runs easterly from Cadwallader Creek, a tributary of Bridge River that flows into the Fraser River. Steep mountains rise from 3300 ft. elevation above sea level at the creek to peaks and ridges at 8000 to 9000 ft. altitude. The lower slopes are heavily wooded while above 5000 ft. bare rocks, glaciers and snow-fields extend far above timber line. See Fig. 1. Story of the Bralorne The Lorne claim, staked in 1897, was one of the earliest lode locations in the Bridge River district. At that time it was 70 miles by steep and treacherous trail from the railway. For the next 15 years Arthur Noel, Will Haylmore and other local prospectors mined the small, irregular, oxidized quartz veins to a maximum depth of 200 ft., and treated the ore in arrastres and later a five-stamp mill. The total production to 1912 was only $55,000. Noel and associates later formed the Lorne Amalgamated Mines Co., which they worked in a desultory way until 1925. Increasing sulfide and arsenide content in the veins

Jan 1, 1935

By Ira B. Joralemon

Neighboring mines, like men who live in the same environment, often assume similar characteristics. In the Bridge River area of the Lillooet district in central British Columbia there is a remarkable exception to this rule. The Pioneer mine has a simple fissure over 3000 ft. long, with no faults of more than a few feet and with only minor variations in width; the Bralorne mine, 2½ miles away at the other end of the same intrusion, has developed one of the most intricate vein and fault systems in the West. Hinged faults displace the main vein for hundreds of feet, and sometimes become valuable veins themselves. The width of ore varies in a few yards from 2 to 60 ft. Changes in grade of ore are just as abrupt. The rocks that follow the veins or are cut by them often grade into one another or into the ore with no definite contacts. Although the complexity of the structure has left many problems unsolved, the development has indicated a fascinating geological pattern. The Bralorne is in the eastern portion of the Coast Range, 55 miles by spectacular mountain road from the Pacific Great Eastern Railway station of Bridge River. The adit tunnel runs easterly from Cadwallader Creek, a tributary of Bridge River that flows into the Fraser River. Steep mountains rise from 3300 ft. elevation above sea level at the creek to peaks and ridges at 8000 to 9000 ft. altitude. The lower slopes are heavily wooded while above 5000 ft. bare rocks, glaciers and snow-fields extend far above timber line. See Fig. 1. Story of the Bralorne The Lorne claim, staked in 1897, was one of the earliest lode locations in the Bridge River district. At that time it was 70 miles by steep and treacherous trail from the railway. For the next 15 years Arthur Noel, Will Haylmore and other local prospectors mined the small, irregular, oxidized quartz veins to a maximum depth of 200 ft., and treated the ore in arrastres and later a five-stamp mill. The total production to 1912 was only $55,000. Noel and associates later formed the Lorne Amalgamated Mines Co., which they worked in a desultory way until 1925. Increasing sulfide and arsenide content in the veins

Jan 1, 1935

By L. S. Darken, H. A. Wriedt

THE constitution diagram of the iron-copper system derived by Daniloff' exhibits, at the iron-rich end, phase fields similar to those of the iron-carbon diagram. At 1484° 1094°, and 850°C there are, respectively, a 6-y-L peritectic, a y-E-Lperitectic, and a a-y-E eutectoid. The phase designated E is copper-rich, containing 4 pct Fe or less according to the temperature. The composition of copper-saturated ferrite (a-solid solution) has been determined previously only twice as a function of temperature. By electrical conductivity measurements, Buchholtz and Köster2 found 3.4 pct Cu in ferrite at a eutectoid temperature of 810°C. They represented the composition of saturated ferrite by the equation, log (%Cu) = 4.32-4125 (1/T) Gregg and Daniloff3 reported that Köster later modified his opinion, raising the indicated eutectoid temperature to 833 °C and the copper content of eutectoid ferrite to 3.7 pct. The literature contains several indications4-7 that the work of Köster and his co-workers is not reliable. Norton,8 using an X-ray method, found that 1.4 pct Cu dissolved in ferrite at the approximate eutectoid temperature of 850°C and that the solubility decreased to about 0.3 pct at 700°C. Below 700°C his experimental data show almost no change in solubility. These findings, although inconsistent with the work of Ruer and Goerens4 and of Smith and Palmer,7 have been used generally since their publication. In investigations involving the observation of arrests in temperature during continuous heating or cooling the precise location of the eutectoid temperature has been obscured by a large hysteresis. The more reliable work 579 where the results are not effected by this hysteresis indicates a eutectoid temperature between 840° and 850°C. The foregoing discrepancies, along with a few observations in this Laboratory relating to the precipitation of copper from ferrite, prompted us to rein-vestigate the lower temperature portion of the Fe-Cu system. In this investigation the solubility of copper in ferrite was obtained by the chemical analysis of high-purity iron equilibrated with copper-silver alloys. This method eliminated the necessity for retaining by a quench the physical state of the specimen existing at the equilibration temperature. EXPERIMENTAL The reaction of the iron with the copper alloy was

Jan 1, 1961

By Paul A. Beck

THE occurrence of the (100) [lOO] or "cube" texture upon annealing of cold-rolled copper has been much investigated.' The conditions favorable for its formation were found to be a high final annealing temperaturez or long annealing time," a high reduction of area in cold rolling prior to the final anneal,' and a small penultimate grain size." The effects of penultimate grain size and of rolling reduction were found by Cook and Richards4 to be interrelated in such a way that any combination of them giving lower than a certain value of the final average thickness of the grains in the rolled material leads to a fairly complete cube texture with a given final annealing time and temperature. Also, according to the same authors, at a higher final annealing temperature a larger average rolled grain thickness, i.e., a lower final rolling reduction, is sufficient than at a lower temperature. These somewhat involved conditions can be understood readily on the basis of recent results obtained at this laboratory. Hsun Hu was able to show recently by means of quantitative pole figure determinations that the rolling texture of tough pitch copper, which is almost identical with that of 2s aluminum: may be described roughly as a scatter around four symmetrical "ideal" orientations not very far from (123) [112]. In the case of aluminum, annealing leads to retain-ment of the rolling texture with some decrease of the scatter around the four "ideal" orientations, and to the appearance of a new texture component, namely the cube texture." A microscopic technique, revealing grain orientations by means of oxide film and polarized light, showed that the retainment of the rolling texture is achieved through two different mechanisms operating simultaneously, namely "re-crystallization in situ," and the formation of strain-free grains in orientations different from their local surroundings, but identical with that of another component of the rolling texture. Thus, a local area in the rolled material, having approximately the orientation of one of the four "ideal" components of the texture, partly retains its orientation during annealing, while recovering from its cold-worked condition, and it is partially absorbed at the same time by invading strain-free grains of an orientation approximately corresponding to that of another "ideal" texture component. The reorientation here, as well as in the formation of the strain-free grains of "cube" orientation, may be described as a [Ill] rotation of about 40°, see Fig. 1 of ref. 6. The preferential growth of grains in such orientations is a result of the high mobility of grain boundaries corresponding to this relative orientation.' " It appears very likely that in copper the mechanism of the structural changes during annealing is similar to that observed in aluminum (except for the much greater frequency of formation of annealing twins in copper). In both metals the new grains of cube orientation have a great advantage over the new grains with orientations close to one of the four components of the rolling texture. This advantage stems from their symmetrical orientation with respect to all four retained rolling texture components of the matrix; they are oriented favorably for growth at the expense of all of these four orientations. As a result, the growth of the "cube grains" is favored over the growth of the others, as soon as the new grains have grown large enough to be in contact with portions of the matrix containing elements of more than one, and preferably of all four component textures. It is clear that this critical size is smaller and, therefore, attained earlier in the annealing process if the structural units, such as grains and kink bands, representing the four matrix orientations are smaller, i. e., if the average thickness of the rolled grains is smaller. Hence, for a given annealing time and temperature, a smaller penultimate grain size and a higher rolling reduction both tend to increase that fraction of the annealing period during which the above condition is satisfied. Consequently, the percentage volume of material assuming the cube orientation increases. The same is true also for increasing time and temperature of annealing when the penultimate grain size and the final rolling reduction are constant, since the average size attained by the new grains during annealing increases with the annealing time and temperature. For the same reason, at higher annealing temperatures a given volume percentage of cube texture can be obtained with larger rolled grain thickness (larger penultimate grain size, or smaller rolling reduction) than at lower annealing temperatures. The well-known conspicuous sharpness of the cube texture may be interpreted as a result of the fact that selective growth of only those grains is favored that have an orientation closely symmetrical with respect to all four components of the deformation texture and exhibit, therefore, a high boundary mobility in contact with each. The effect of alloying elements in suppressing the cube texture, as described by Dahl and Pawlek,' appears to be associated with a change in the rolling texture. For face-centered cubic metals, such as copper, which do exhibit the cube texture upon annealing, the rolling texture is always of the type described above, i. e., scattered around four "ideal orientations" of approximately (123) [112]. The addition of certain alloying elements, such as about 5 pct Zn or 0.05 pct P in copper, has the as yet unexplained effect of changing the rolling texture into the (110) 11121 type. This texture consists of two fairly sharply developed, twin related components. In such cases, as in 70-30 brass and in silver, the annealing texture again is related to the rolling texture by a [lll] rotation of about 30°, however, because of the different rolling texture to start from, it has no cube texture component. At higher temperatures, both in brassm and in silver," grain growth leads to a further change in texture: A [lll] rotation of the same amount, but in reversed direction, back to the original rolling texture.

Jan 1, 1952

By J. F. McLaughlin

The installation of the Chance flotation system for the preparation of anthracite demonstrated the need for some means by which the specific gravity of the fluid mass in the separating cone could be determined with a fair degree of accuracy. The means used (with the original cones) was a set of wooden balls attached to ropes or chains, each with a different specific gravity. In the Northern Anthracite field, where the cones are generally operated with the specific gravity at 1.65, these gravity balls are provided with a range of 1.6, 1.65, 1.7 and 1.75. The operator of the cone from time to time would place one or more of these balls in the cone and estimate its specific gravity by the action of the balls. The margin of error with this system was very great, because there was the possibility that the test ball might be held in suspension and carried around by material floating beneath the surface of the cone in what is known as "island " formation—a formation resulting from a low sand line in the separating cone. The inaccuracy of this system of determination of specific gravity was indicated also by periodic tests of the final refuse bank, which showed excessive losses of coal to the refuse. Further evidence of the unsatisfactory operation of the cone was the uneven appearance of the market product, which caused spasmodic condemnation. Maintenance of Even Preparation With the critical market condition of today, it is essential that the preparation units operate consistently to maintain an even preparation. If a plant wishes to maintain a 1.65 specific gravity preparation, which means that all material of a specific gravity greater than 1.65 is classed as "reject," if the specific gravity of the fluid mass runs to 1.70, the market product will not be satisfactory; or if the specific gravity of the fluid mass is 1.60, the market product will be entirely satisfactory but the losses at the slate ends will be excessive. The desired specific gravity of the fluid mass in the Chance cone is maintained by agitating the sand within the separating cone by means

Jan 1, 1931

By C. R. Tinsley

A strictly limited number of specialist project finance banks exists, with the necessary specialist personnel able to address project cost-competitiveness, cost and commodity forecasting, joint venture negotiations, and completion and sales undertakings. Different techniques of internal bank appraisal are applied, with the usual equity investment project valuation criteria being seldom useful for loan assessment purposes. Company attention paid to the preparation of a sensitivity structured bankable presentation can prove most effective for securing development project finance. Substantial scope is possible for savings of time and money.

Jan 1, 1985

By Thomas Wright

No symposium on die casting could be complete without consideration of the methods of formula and impurity control. No consideration of control would be complete without discussion of that new tool of the analytical chemist and inspector, the quartz spectrograph. Just as there is an increased appreciation of the possible applications of the die-casting process and of die castings, so too is there, slowly but surely, of the spectrograph. At present the role of the instrument might well be compared with that of brass or magnesium die castings. Neither is as widely understood or known as they will be. Product specifications are increasing and are getting more explicit and often more rigid. The designer, metallurgist and production engineer or superintendent will be forced to depend more and more on the chemical laboratory. Sampling, as well as analyses, testing and inspection problems will grow. The chemical laboratory must meet its increased responsibility and use of the spectrograph will become more common. Advantages of increased precision, greater speed, lessened labor and often lower costs cannot be ignored. It will be the purpose of this paper in part to outline the respective fields of chemical and spectrographic analyses and to indicate where one may overlap the other and in part to indicate certain possible sources of contamination of zinc-base die-cast alloys in particular.

Jan 1, 1935

By Luis Toron, Salvador Cortes-Obregon

John D. Price (Colorado Fuel and Iron Corp., Pueblo, Colo)—The paper on the coal fields of the Oaxaca district as prepared by engineers Toron and Cortes-Obregon of the staff of the Bank of Mexico bears witness to the thorough and careful way in which the men associated with this organization perform their work. There is little to be added to their paper in way of discussion other than to confirm and amplify some of their statements. Since the only extensive and well-developed field of coking coal lies in the northeastern section of the country adjacent to Sabinas in the state of Coahuila, it follows that blast furnace plants would be located in that same region. Two such plants are now operating at Monterrey and Monclova, using coke produced at the Sabinas district mines. But the nearer of these two plants is 600 miles from Mexico City and even farther from the center of population. Transportation of products from these mills to the market area is therefore expensive, both because of the distance and the difficulty of the terrain over which it must be carried. The development of an integrated steel industry closer to the center of population has therefore long been a goal toward which the Mexican technicians have been striving. While the presence of coal of some grade has been reported in many of the states, and many ideas have been advanced regarding its possible uses in iron and steel production, deposits of anthracite in Sonora and the various coals of the Oaxaca district as reported on in this paper are the only ones that have been explored in a serious manner. The coking coal from the Mix-tepec zone appears to offer promise of producing a coke which could be used in a standard blast furnace. Several problems are indicated, however: 1—The ash in the coal is high as mined, but indications are that it can be washed to an ash content of 15 pct with a recovery of 70 pct of washed coal. 2—Such washing would increase the volatile content from 17.4 pct to about 20 pct, and in a byproduct oven this should give a coke yield of close to 80 pct with an ash content of coke under 20 pct. 3—A free swelling index of 5 appears low for a good coking coal, and below that of the coals from the Sabinas district, which show between 6 and 9. But washing of the coal should result in an improvement in this regard; in the United States coals from Utah with an index even lower than 5 have made a usable coke. 4—A coal with volatile as low as 17.4 in raw coal and 20 in washed coal would come close to being classed as a low-volatile rather than medium-volatile coal, and low-volatile coals are notorious for their high expansion properties. Several plants in the United States are making coke from straight medium-volatile coal of 26 to 28 volatile content, and one at Rosita,, Mexico, from coal of 25 volatile. But no plants to my knowledge are using coal as low as 20 volatile. Since the Rosita coal appears to be a borderline coal from the angle of its expansion properties the coking of one of the straight lower volatile must be approached with caution. 5—There are few coals possessing any degree of coking properties which cannot be used in coke production by careful attention to its preparation and blending. The fact that coals of other types are available in this same region make improvement through blending very possible. 6—There are other workable methods of reducing iron ore other than the conventional coke-blast-furnace method. These will not be discussed here but it is known that their use has been considered. The technicians of not only Mexico but also of the other Latin American countries are keenly aware of their natural resources and their national needs. This paper emphasizes the fact that the Mexican technicians are working on their problem and attempting to speed the day of self-sufficiency for their country. Salvatore Cortes-Obregon (author's reply)—I wish to thank Mr. Price for his kind remarks. The Mixtepec coal as shown in Table II has 30 pct ash and a free swelling index of 5, but when the same coal is washed to 15 pct ash it has a free swelling index of 8 to 9 and the volatiles increased from 17.4 to 20.7 pct. A satisfactory coke has been produced from blends made in the Mexican laboratory using at least 40 pct of the Mixtepec coking coals with the other Oaxaca non-coking coals. Koppers in Germany report good coke obtained from the Oaxaca coal with a blend of 80 pct Mixtepec coal. Consideration is being given the possibility of using methods other than the conventional blast furnace for the reduction of iron ore near the Oaxaca area; electric furnaces appear promising. The non-coking coals could be used to produce cheap electric energy and the coking coals to make metallurgical coke.

Jan 1, 1955

By C. T. Holland

This discussion is concerned with those com-J- paratively infrequent bumps that eject material from the failed mass with enough energy to wreck heavy machinery and seriously injure or kill people. In such cases there is a loud report, and the pressure waves set up in the mine atmosphere wreck ventilation controls several hundred feet from the site of the failure. Vibrations are induced in the earth's crust which may be detected several miles from the point of origin, and explosive gases may be released in large quantities." Usually a dense suspension of dust is produced from the failed mass, together with dust placed in suspension by the resulting air blast. These rock bursts or bumps have also been called mountain shots, bounces, pillar bursts, and crumps. Such occurrences, when extensive, are serious catastrophes, and may force abandonment of the mine.' CAUSES OF ROCK BURSTS Rock bursts in coal mines have occurred in sev-eral places under diverse geological conditions, mining methods, and/or practices. A study of many such occurrences indicates that they depend on the following conditions: 1) cover thickness of 500 ft or more, 2) cover composed of strong members and having a strong member lying close to or on the coal bed, and 3) a floor that does not readily heave. The minimum thickness of cover under which coal mine rock bursts occur is based on observation. Under unusual circumstances involving very strong roof members and/or mining systems, and/or practices especially likely to induce bursts, a burst conceivably could occur under less roof. Under special circumstances rock bursts have occurred in marble quarries almost at the surface.' Furthermore, under very heavy cover it is likely that controlling roof members need not be as strong and as close to the coal bed to cause bumps. Just what constitutes a strong roof or roof member is open to question. Obviously the strength of a roof member is determined by its thickness, the distance between bedding planes across which little or no bond exists, the spacing and development of joints, and the inherent strength of the rock material. Strata sections showing roofs under which rock bursts in coal mines have developed are illustrated in Fig. 2. In several instances the strength of rocks composing the roof of mines developing bumps has been tested. The tests (Table I) show that these rocks develop considerable strength in compression, bending, and shear and that in some cases the strengths are very high. No quantitative data are available concerning joints in roof members over coal beds where rock bursts have occurred, but the author has examined many mines subject to bursts and has received the impression that joints in the shales immediately over the coal beds are always present and well developed. In the main sandstone or conglomerate members themselves, however, where observed, he has yet to see a fracture system indicating well developed joints. Undoubtedly joints are present, but the evidence indicates that they are few, poorly developed, and widely spaced. In a mine subject to bumps, moreover, actual measurements of roof spans up to 225 ft between supports under a cover of about 800 ft have been made before roof failure occurred. Such spans signify strong roof. In coal beds subject to rock bursts, the bottom rock is usually a sandy shale. A few tests (Table I) have been made on the floor rock of coal beds. Observations in mines indicate that they are usually weaker than roof rocks and that under load they show more plastic effects. Observation in ground subject to bursts invariably indicates a relatively strong floor highly resistant to heaving, although some heaving may occur. Other factors listed as favorable to rock bursts are: 1) mountainous surface; 2) steeply dipping beds; 3) proximity of faults, folds, and other similar geologic structures. Many bursts have occurred in mines under mountainous country, although many bursts in mines under a surface with little relief indicate that thickness of cover, not relief, is the essential factor. In steeply dipping beds, many bursts occur in mines that are essentially flat, and rock bursts have been recorded in beds dipping as much as 35o Most rock bursts in the U. S. occur in beds that dip less than 10 pct. It is also true that most U. S. coal is produced from beds dipping less than 10 pct. Geologically, faults are regarded as providing stress relief, but in the vicinity of faults, as well as sharp folds, it would seem that residual stresses could be present. The fact that rock bursts occur in areas remote from faults and sharp folds suggests that their presence is not necessary for this phenomenon to take place. There have been many bursts close to faults and folds, and it may be assumed that occasionally such structures accentuate their occurrence3 The conditions and circumstances under which they occur, coupled with the effects produced, leave In the Crowsnest Pass coalfield, British Columbia, a rock burst is estimated to have released 1.8 million cu ft of methane.

Jan 1, 1959

By R. W. Raymond

In a paper on the Divining-Rod (Transactions, xi., 411), presented at the Boston meting, in February, 1883,I suggested, among other signs of which the skilled prospector might consciously or unconsciously avail himself in the search for springs of water or mineral deposits, the presence of particular grasses or shrubs, and the general appearance of the flora. There are not wanting, in the literature of our art, indications that greater attention was paid to such signs in ancient days than at present. It is the tendency of a scientific age to ignore, or regard as superstitious, what cannot be classified or explained. But the acute though unscientific observation of earlier times often, doubtless, accumulated facts which it would be worth our while to re-examine. It is highly probable that the mining prospectors of the sixteenth and seventeenth centuries, to go no further back, were close students of nature, and knew, though they could not explain, and, for reasons of private gain, would not communicate, many curious and significant facts. We can generalize better than they could; but for that very reason we are, perhaps, their inferiors in observation. They let nothing pass, because all things were equally significant to them. We take wider views, and miss more. And, no doubt, there still exists among practical millers, prospectors,, hunters, Indians, etc., an unwritten.

Jan 1, 1887

By Arthur L. Hall

The purpose of this symposium is to challenge industry to better air classification, and as a first step toward fulfillment, to spotlight three fairly recent, aerodynamically designed, highly efficient air classifiers. This discussion does not intend to imply that present practice is bad. Actually, when present air separators are used for the jobs for which they were originally designed, they give high production and uniform operation, with a minimum of maintenance. The customers of nonmetallic mineral pigment producers, however, are making increasing demands for finer products, such as 99.9 pct -20µ, which appears to be beyond the economic capabilities of standard air separators industrially used over the last 10 to 30 years. Here is an actual example. A 10-ft diam pebble mill is in closed circuit with a conventional 12-ft diam centrifugal air separator, grinding 20-mesh wollastonite to a series of fine grades up to 99.97 pct - U.S. No. 325 mesh (44µ). Fig. 1 shows variations in fineness of mill discharge (or air separator feed), recycle for regrinding (or air separator tails), and product as measured by a wet rotating sieve test. For example, when the mill is turning out a product 80 pct -325, the mill discharge or separator feed is 30 pct -325, and the coarse tailings or recycle is only 5 pct -325. Hence 25 pct of the 30 pct available is accepted as product. This is efficient separation, and the ratio of recycle to feed is only about 2: 1. How- ever, to make a product testing 99.97 pct -325 the separator feed is 62 pct -325 and the coarse tailing 60 pct -325. In this case the recycle ratio is about 13: 1 and the economic limit to produce fineness has been substantially reached with this given equipment and process arrangement.

Jan 10, 1957

By J. Chipman, G. G. Hatch

One of the important functions of the iron blast furnace is the desulphur-ization of pig iron before it enters the steelmaking furnaces. However, the increasing concentrations of sulphur in the metallurgical coke, source of approximately 90 pct of the sulphur present in the blast furnace charge, and demands for higher rates of production within recent years have increased the need for greater desulphurization within the iron blast furnace. Furnace operators are beginning to look for desulphurizing agents other than blast furnace slag to accomplish the desired degree of desulphurization. A considerable amount of work has been done on desulphurization outside the furnace with soda ash, calcium carbide and various synthetic slags. Whether the desulphurization of pig iron is accomplished wholly inside the furnace or partly inside and the remainder outside, will be determined by the economics involved. Regardless of which is the case, it is believed that it is necessary to have a better understanding of the physical chemistry of desulphurization by blast furnace slags. To this end, it is the object of the present investigation to attempt what is believed to be the first equilibrium study of the distribution of sulphur between liquid pig iron and a wide range of blast furnace slag compositions. Review of Literature There is a considerable amount of information in the literature concerning the desulphurizing power of iron blast furnace slags, the solubility of various sulphides in the slags, and the effect on desulphurization of temperature, of elements dissolved in the liquid iron, and of viscosity. However, there is nothing to indicate that the equilibrium distribution of sulphur between liquid iron saturated with carbon and iron blast furnace slags has been studied experimentally. Wentrupl has made probably the most detailed study of the desulphurization of pig iron to date. He considered that there are three distinct aspects involved, namely: 1. Desulphurization within the blast furnace (by lime and manganese). 2. Subsequent desulphurization by manganese. 3. The effect of subsidiary reactions on the desulphurization by manganese. The experimental work carried out by Wentrup was devoted mainly to obtaining a better understanding of how desulphurization by manganese was accomplished in the mixer and the ladle. Particular attention was given to the part played by carbon, silicon, and phosphorus associated with manganese in the iron, and the effect of temperature on desulphurization. The experimental results indicated that desulphurization by manganese is purely a process of crystallization of manganese sulphide. The addition of silicon to iron melts containing 3.5 pct carbon and less than 0.5 pct manganese had no noticeable effect on desulphurization, but with 1-2 pct manganese the silicon additions improved the desulphurization. Additions of phosphorus also resulted in improved desulphurizati011 by manganese, but the effect was not as marked as in the case of silicon. It was also found that desulphurization by manganese was further improved by lowering the temperature. In order to explain desulphurization inside the blast furnace, Wentrup considered the system iron, sulphur, calcium, oxygen, manganese. (silicon). The distribution of sulphur between the metal and slag was represented by the following equation: (SS) _ (S)Fe + (S)Ca + (S)Mn .... [S] = [s] [1] The parentheses and the brackets represent the equilibrium concentrations in weight per cent of the slag and metal constituents, respectively. Since FeS D (FeS) _ (FeS) (S)Fe LfeS - [FeS] [S] [2] (CaO) + S e (FeO) + (S)Ca _ (FeO)(S)Ca. (S)Ca _ (CaO) Kl = (CaO)[S] [S] ~K1(FeO) [3] Mn + S D (S)Mn (S)Mn (S)Mn K' = [MnpT "1ST = *lIMnJ !4) Substitution of Eq 2, 3, and 4 into Eq 1 resulted in if = L- + *> (Sol + K^ (S) [51 Eq 5 was used to calcu1;lte -f^j and [S] at 1480°C for slags containing 30-50 pct lime, 0.1-2.5 pct iron oxide, 0-26 pct silica, 2 pct sulphur and iron analyzing 1.5 pct manganese. The value for LFaB at 1480°C was found to be equal to 4.5, based on the experimental work of Bardenheuer and Geller.2 The results of the calculations are shown in Table 1. Although the slags are hypothetical and do not represent the range of compositions found in ordinary blast furnace practice, the calculations indicate that lime is effective in controlling desulphurization only if the iron oxide and silica contents of the slag are kept low. Schenck3 did not claim K1 to be a true equilibrium constant, but an empirical value which varied with the silica content of the slag.

Jan 1, 1950

By George M. Schwartz, Ian L. Reid

THE Soudan mine in the Vermilion district of northeastern Minnesota is the oldest iron mine in the state. It has shipped ore every year since 1884 and still contributes a yearly quota of high grade lump ore. No comprehensive report on the Vermilion iron-bearing district has appeared since Clements' monograph,' but Gruner2 discussed the possible origin of the ores in 1926, 1930, and 1932, and recently Reid and Hustad have added data on mining and geology .3, 4 For many years geologists of the Oliver Iron Mining Div., U. S. Steel Corp., have kept up to date a series of plans and vertical sections of the Soudan mine. In connection with mine operation considerable diamond drilling has been done, and this, together with the mine openings, has permitted a reasonably accurate picture of the structure of the orebodies and wall rocks. It has long been evident to geologists familiar with the mine that the ores were not a result of weathering, a point emphasized by Gruner in 1926 and 1930. As the deeper orebodies were developed it also became clear that replacement had played an important part in their development. In recent years it has been recognized that other iron ores were formed by replacement, as Roberts and Bartly5 have argued strongly for the deposits at Steep Rock Lake. On the basis of these facts G. M. Schwartz suggested to members of the Oliver staff that it would be desirable to study the evidence of replacement, particularly the possible alteration of the wall rock which would be expected if the replacement was a result of hypogene solutions. Rock Formations: The formations directly involved in the iron orebodies of the Soudan mine are few though far from simple. The country rock is largely the Ely greenstone of Keewatin age consisting of a mass of metamorphosed lava flows, tuffs, and intrusives which have been more or less altered by hydrothermal solutions. The predominant rock is chlorite schist. Interbedded with the original flows and tuffs are a series of beds and lenses of jasper to which the name Soudan formation has been applied. In the Vermilion district the term jaspilite has been used for interbanded jasper and hematite. According to modern usage these jasper or jaspilite beds do not comprise a formation separate from the Ely greenstone, inasmuch as the beds of jasper are interbedded with the flows and tuffs of the upper part of the greenstone. It would more nearly accord with modern usage to consider the Soudan beds a member of the upper part of the Ely formation. Because of incomplete rock exposure and exploration the number of interbedded jaspilite beds is unknown. In the mine, however, as many as nine major beds of jasper are known on a cross-section of one limb of the syncline, with an equal number on the other limb. In addition diamond drill cores show beds of greenstone down to half an inch in thickness. The thin beds are probably always tuffs. Structure: Rock structure in the Soudan area is complex, and because there are no recognizable horizons within the greenstone it is extremely difficult to work out the details. Generally speaking, the major regional structure is an anticlinorium, the axis trending east-west, with a westerly pitch. The Soudan mine is related to a synclinal structure on the north limb of the anticline about a mile from the west nose of the folded iron formation. The general structure at the mine is that of a closely folded minor syncline on the major regional anticline. A cross fault has dropped the east side so that the bottom of the syncline has not been reached, whereas to the west it is well shown by the mine openings and diamond drill exploration. Throughout the mine the beds of jasper, and ore-bodies that have replaced the jasper, normally dip northward at angles of 80" or steeper. In detail the jasper beds are extremely folded, probably as a result of deformation while they were still relatively unconsolidated. Orebodies: Ore in the Soudan mine is mainly a hard, dense, bluish hematite. Locally ore has been brecciated and cemented by quartz. The vugs commonly occurring near the borders of orebodies are lined with quartz crystals. They seem to have formed as part of the ore-forming process and are evidence that no folding or compression of the ore has taken place. The orebodies are numerous, varying greatly in size. Many lenses of high grade hematite are too small to be mined. Some of the larger orebodies have been followed vertically for as much as 2500 ft and horizontally up to 1500 ft. The large ore-bodies are extremely irregular in outline in the plane of the beds of jaspilite. In width they are more regular, as they are strictly governed by the width of the jaspilite beds and the greenstone wall rock, which seems to have resisted replacement by hematite. At many places the orebodies replace the jaspilite completely and have a footwall and hanging wall of greenstone. At other places either one or both walls may be jaspilite. Geologists who have studied the orebodies in recent years agree that evidence for the replacement origin of the hematite bodies seems conclusive. AS noted above, many of the orebodies replace jaspilite beds from wall to wall with no evidence whatever of compaction. The replacement origin is also supported by details of the banding which is characteristic of the

Jan 1, 1956

By O. U. Bradley

The methods of handling the oil output of the Mid-Continent fields are not unlike those practised in other oil fields of the United States, and it is not expected that this paper will present any entirely novel ideas. Some features of the practice of handling this oil may be of interest, however. Wooden Gaging Tanks Practically all of the oil sold to pipe line? and other purchasing agencies throughout the Mid-Continent field is gaged from wooden tanks. Steel gaging tanks are slowly coming into use, and are preferable in many ways—leakage and evaporation are reduced, and the strapping and recalculation of capacities are unnecessary. The wooden tanks are slightly coned to facilitate tightening the hoops. The sizes in common use are 250, 500, 800 and 1600 bbl. The number and the capacity of the gaging tanks upon any producing oil property depend upon its production, and the facilities for disposing of the output. The 1600-bbl. size was the most common in the Cushing field, particularly during the period of flush production. The tanks are distributed over the leased premises in groups of two, three, and four tanks each, and these are filled from flow tanks, or through direct pipe connections with the heads of the wells. Flow tanks are merely temporary reservoirs to facilitate the transfer of oil from the wells to the gaging tanks. They diminish fire risk, and are considered indispensable on small producing properties where water is pumped with the oil, as they permit the refuse and water to 'settle out before the production is run to the gaging tanks. strapping and Calculations After a gaging tank has been strapped (measured) and its contents computed, a deduction of 3 per cent. is made from its computed volume. The strapping of a tank requires four steps: (1) The outer circumferences

Jan 1, 1920

By Donald B. Evans, Robert D. Pehlke

The solubility of nitrogen in liquid Fe-A1 alloys has been measured up to the solubility limit for formation of aluminum nitride using the Sieverts method. The activity coefficient of nitrogen decreases slightly with increasing aluminum content in the range of 0 to 4 wt pct Al. Based on a nitride composition, AlN, the standard free energy of formation of aluminum nitride from fhe elements dissolved in liquid iron has been determined to be: ?F" = -59,250 + 25.55 T in the range from 1600º to 1750ºC. The solubility of nitrogen in liquid iron alloys and the interaction of nitrogen with dissolved alloying elements in liquid iron have been the subject of a number of research investigations.' Most of this work, however, has been reported for concentrations well below those necessary for the formation of the alloy nitride phase. Data in the concentration region near the solubility limit of the alloy nitride, particularly for systems exhibiting stable nitrides, are important in evaluating the denitrifying power of various alloying elements. They are also useful in determining the stability of a given nitride if it is to be used as a refractory to contain liquid iron alloys. In view of the importance of aluminum as a deoxidizing agent in commercial steelmaking and the fact that its nitride, AIN, is a highly stable compound and has merited some consideration as an industrial refractory, the following investigation was undertaken. The use of the Sieverts technique provided a measurement of the equilibrium nitrogen solubility in liquid Fe-A1 alloys as a function of nitrogen gas pressure up to 3.85 wt pct A1 in the temperature range of 1600º to 1750°C. The values obtained by the Sieverts method were checked by means of a quenching method in which liquid iron was equilibrated with an A1N crucible under a known partial pressure of nitrogen gas, and the solubility of A1N in liquid iron determined by chemical analysis. EXPERIMENTAL PROCEDURE The theoretical considerations involved in determining the solubility product of a solid alloy nitride phase in liquid iron by measuring the point of departure of the nitrogen gas solubility from Sieverts law have been discussed by Rao and par lee.' The principal problem is to determine the variation of nitrogen solubility in an alloy as a function of the pressure of nitrogen gas over it with sufficient precision to establish the break point in the curve at the solubility limit of the alloy nitride phase. A fairly large number of data points are required to do this. A second problem is the determination of the composition of the precipitated solid nitride phase. This is necessary in order to define completely the thermodynamic relationships. The Sieverts apparatus used to make the nitrogen solubility measurements in this investigation is of essentially the same design as that described by Pehlke and E1liott.l The charge materials were Ferrovac-E high purity iron supplied by Crucible Steel Co. and 99.99+ pct pure aluminum. Recrystal-lized alumina crucibles were used, and were not attacked by the liquid alloys. The hot volume of the system which was measured for each melt ranged from 46 to 50 standard cu cm and was found to decrease linearly with decreasing pressure and with increasing temperature. The temperature coefficient of the hot volume at 1 atm pressure of argon gas was essentially constant for all experiments at a value of -6 X 10-3 cu cm per "C. The melt temperature was measured with a Leeds and Northrup disappearing filament type optical pyrometer sighted vertically downward on the center of the melt surface. The temperature scale was calibrated against the observed melting point of pure iron taken as 1536°C. The emissivity of all melts was assumed to be that of pure iron, taken as 0.43. The charge weights ranged from 110 to 140 g and the range of aluminum contents covered was from 0 to 3.85 wt pct. Aluminum additions were made as 12 to 15 wt pct A1-Fe master alloys previously prepared in the system under purified argon. The compositions of the master alloys were checked by chemical analysis and found to be in agreement with the charge analyses. Vertical cross sections of the master-alloy ingots were used as charge material for the equilibrations in order to minimize the effect of any segregation which might have occurred during solidification of the master alloys. Determinations of the solubility product of

Jan 1, 1964