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ANYONE who wishes to make the patterns for statues in order to cast them later in bronze, even though he himself is the artificer who made them, must take many things into consideration: first, whether they are large or small, and whether they are wholly round or only in part; then of what material they are made, for the way that can be followed if they are of wax cannot be used if they are of clay, wood, marble, or even of plaster; likewise he must consider whether or not it has irons inside to support it. There is also a great difference in facility if the support is placed more in. one way than in another. Great consideration should be given to whether it is to be made hollow or solid. If you have not been the workman yourself, you must at least have been an active helper in this and in every other part in order to be able to observe everything without a fault. For if they are large it is necessary to make their moulds in pieces, with marks and measurements so that you can replace them-at will exactly in the same order and in the same places after they have been taken out. But if you are the artificer [as well as sculptor] the process is made easier. Taking the ordinary way and, over an iron, make out of well beaten moulding clay, compounded with cloth clippings a core exactly as the statue is to be or however large you wish the hollow to be. Dry and bake it. On top of this work the wax, thickening it all over as much as you wish the bronze to be and finish it completely, remembering to arrange that the base may come out solid with bronze so as to support the whole work. Then take some rods of hammered bronze one dito thick and one palmo long. Pass these throughs the thickness of the wax and fasten them in several places in the core of clay that you made. [81] Do it so that they, project out of the wax a good three or four dita. These are, put in so that when the wax is removed the core is held in the middle of the mould that forms the outer part, and this outer part is held by the inner part, namely the core, so that the mould may not move in any part from its place even when it is handled. This done, take some of that clay made with well-prepared cloth clippings as I taught you, and mix with it if you wish some wash ashes or

Jan 1, 1942

By Clarence E. Jackson, George G. Luther

THE ever increasing array of information concerning the measurement of the effect of the welding process on the properties of a steel is an indication of the tremendous effort that is being expended in this field. The methods that have been considered may be divided into two groups: first, those that measure directly the effect of welding on the ductility of a steel, and, second, those that observe the change in some property, such as the hardness of the heat-affected zone, and assume a direct correlation between this change and the resultant ductility. The nick-bend test is a direct method designed for studying the effect of welding on the ductility of a steel. The test specimen (Fig. I) consists of a bar cut from a plate with a bead weld on the surface. The specimen is prepared with a saw-notch placed so as to restrict deformation to the heat-affected zone under the bead weld when tested as a static bend. Failure is caused by rupture of the low-ductility metal in the heat-affected zone. The angle at maximum load is used as an index of the effect of the welding process on the material. In the development of the nick-bend test considerable scatter was noted in some of the results, although the agreement of duplicate specimens welded at the same time and tested consecutively was excellent. Upon further investigation it was found that the change in ductility was dependent upon the interval of time between the welding operation and the testing of the specimen. For the six steels of Fig. a, the angle at maximum load in four cases was doubled and in one case was tripled. All steels that have been studied show an improvement in the angle of bend for longer intervals between welding and testing. This same behavior has been found in weldability tests using the Navy T-bend

Jan 1, 1945

By G. E. Steudel

THE study of the possibility of effecting a lower cost in the manu-facture of pig iron reveals the importance of the ever present question of slag chemistry and volume. Factors that determine slag chemistry and volume are the impurities in the ore available, the ash-sulphur content of the coke, and a flux material necessary to promote removal of sulphur. Of these, the ores present by far the major source of impurities and what may prove to be the lowest practical slag volume with desired chemistry in one locality may be considered low or high in some other center of iron manufacture. The Chicago district is not only blessed with ores of high quality, but with excellent coke and flux materials as well. It is the intent of this paper to discuss slag problems at the South Works only, where burdens consist of Mesabi ores, Michigan stone, and coke ranging from 7 to 9 per cent ash and containing an average of 0.57 per cent sulphur. Early in the development of modern blast-furnace practice, the con-sideration of slags was secondary to the many adjustments of mechanical difficulties, such as furnace lines, stock distribution, stove design for obtaining higher blast temperatures, improved chemical segregation and size of raw materials. Furnace operators, after making these improve-ments, turned their attention to the study of other refinements in practice, of which slag chemistry appeared to be the most important. It is only within the last two decades that operators have become familiar with the various aspects of slag compositions. The industry, while rather slow in its recognition, is indebted to such investigators as Rankin,1 Wright,1 Feild,2 Royster,2 McCaffery,3 Kinney,4 Holbrook,5 and Joseph5, whose researches concerning the elimination of sulphur within the blast furnace and the effect of slag composition upon melting point, viscosity, and desulphurizing power have established criteria for the selection of slags of the requisite quality.

Jan 1, 1939

By R. I. Jaffee, J. W. Roberts, D. N. Williams

DISPERSION hardening as an alloying process has aroused increasing interest in the past few years. This alloying procedure, in which an insoluble phase is dispersed randomly through a metal or alloy, is believed to be especially suited for the attainment of creep strength at elevated temperatures since the dispersed phase is metallurgically stable. Another advantage is that the physical properties of the matrix phase, such as electrical conductivity, remain relatively unaffected by alloying. These two characteristics resulted in interest in the applicability of dispersion hardening to copper and copper alloys. A study of copper alloys prepared by powder metallurgy methods has shown some strengthening from oxide dispersions.' However, powder metallurgy processing would not be economical for large scale production of dispersion hardened copper-base alloys. Therefore, the research described in this article was initiated to examine methods of forming dispersion alloys which would not require powder metallurgy processing. If powder metallurgy processing is to be avoided, a method must be found by which the dispersion material can be introduced into and dispersed through the melt before casting. When the dispersed phase is an oxide or similar nonmetallic, it is extremely difficult to form a satisfactory liquid: solid dispersion since successful mixing requires that the melt wet the dispersed phase. It is almost impossible to force a nonwetted particle beneath the melt surface. This difficulty can be avoided, however, if the oxide particles are produced in molten copper by reaction within the melt. The problem would then reduce to maintaining randomness until solidification. Four procedures were considered for producing an oxide particle within molten copper by reaction. In all cases, the intent was to react oxygen (0) with a reactive metal (X) to form an oxide (XO) within the copper melt. These reactions were as follows: 1) (Cu-X), + (O)g (Cu)1 + (XO)S [1] 2) (Cu-X)1 + (Cu2O)s (Cu)x + (XO)S [2] 3) (Cu-OK)1 + (Cu-X)1 (Cu)1+(XO)9 [3] 4) (Cu-OK)1 +(X)s (Cu)1 +(XO)s [4] The reactions were carried out in an 8-lb melt of 99.98 pct Cu maintained in the vicinity of 2200oF by induction heating. Where possible the amount of the reactants was controlled to give 1 to 5 vol pct oxide. Mechanical stirring during the reaction period was used to minimize agglomeration of the reaction product. The presence of a reaction product was determined metallographically by study of an ingot cast from the melt at the end of the reaction period. Reaction [I] was studied by dissolving aluminum in the melt followed by forcing oxygen through the melt. The melts became quite sluggish during the reaction period suggesting that A12O3 was formed. However, the reaction product rose out of the melt rapidly, possibly due to the flushing action of the oxygen stream, and little dispersed phase was formed. Reaction [2] was examined by dissolving aluminum, silicon, or nickel i; the melt and then adding Cu20 (melting point 2250F). This procedure was unsuccessful. Although Cu2O was easily stirred into cop-

Jan 1, 1961

By Carl Zapffe

The object of the Bradley process is to free manganese oxide from its associated gangue and separate the contained iron oxide by dissolving the manganese and precipitating it from the solution. This results in producing two oxides as concentrates, the one to be converted into a manganese sinter and the other into an iron sinter. The manganese sinter may be made to contain as high as 70 per cent. inanganese and as low as 1 per cent. silica, with phos'phorus not to exceed 0.02 per cent, and from that figure downward to about 0.005 per cent., and with sulfur at 0.007 per cent. The iron sinter contains about 2.50 per cent. manganese, making the total recovery of the manganese 98 per cent. The method involves selective roasting, which converts the manganese into a form in which it is soluble in an aqueous solution of ammonium sulfate, then precipitating the manganese as a hydrate by using ammonia gas, then stabilizing and washing the hydrate, followed by sintering; concurrently most of the iron oxide is converted by roasting into an insoluble form and becomes a residue with magnetic qualities, which makes possible the recovery of much of the iron in the ore. Leaching manganese by using sulfur dioxide gas was shown possiblc Inany years ago. Mr. Bradley tried it for a year, working in conjunction with the original experimenters of that method, and found that while relatively pure inanganese oxide could be produced, the likelihood of developing a commercial process along this line was not very encouraging. He also attempted leaching with nitric and sulfur acids, but the cost of recovery mounted and the quality of the product was not as good Mr. Bradley tried nearly every leaching method proposed, and the expen-ments were made on a large scale in the spacious laboratories of John C. Wiarda & Co., in Brooklyn, N. Y. Insufficient encouragement was discernible in the results obtained. Upon the suggestion of N. Arthur Laury, of John C. Wiarda & Co., Bradley undertook using arnrnoniurn sulfate. Because a high rate of recovery appeared feasible, Bradley planned the erection of a 4-ton pilot plant at the State of Minnesota Mines Experiment Station, at Minneapolis. The results of that work are described in this treatise.

Jan 1, 1929

By W. A. Thomas

Steam is, and doubtless always will be, the basic power in the anthracite industry, either directly applied through engines and pumps or electrically. The rapidity with which electric power is being applied to the mining and preparation of anthracite leaves little doubt as to its utility and economy. There is practically no operation, except possibly the drilling of hard rock, that has not been successfully electrified and the matter of drilling is handled economically with electrically driven compressors and the standard compressed-air drills. Conservation and the seemingly ever-increasing cost of production, justify careful consideration of the possible results of the more extensive use of electric power in the industry. It is not planned to discuss, herein, the relative economies of individual applications but to consider the relative economy and flexibility of steam and electric power. The reports available for the year 1920 indicate a production of slightly over 80,000,000 tons of anthracite, of which 8,843,500 tons were used for power and heating purposes; in other words, over 11 per cent. of the production was used at the mines. Some of this was used to generate electric power at the mines, about 160,000,000 kw-hr., while 150,000,000 kw.-hr. was purchased from the public utility companies. One large producer, who is buying some electric power and generating some, used over 13 per cent. of the production for power and heating at the mines. Another producer, where considerable electric power is generated on the properties, is using less than 9 per cent. of the output for power and heating. One of the largest producers has a fuel consumption of over 15 per cent. of the output, besides which considerable power is purchased. A careful study of the field indicates that steam-operated plants use upwards of 13 per cent. of the production for power and heating, while electrically operated plants use about 1 per cent. for heating; so that, for steam-power purposes at least 12 per cent., or 268.8 Ib., are used for power out of each ton produced. In the complete electrification of anthracite mining operations, no set rule can be given for the power requirement, in kilowatt-hours per ton, as conditions vary widely in the matter of pumping, ventilation, and transportation, but the following cases are given as a basis of determining a range and a possible average power requirements:

Jan 1, 1922

By C. E. Berry

Natural stone or manufactured porcelain pebbles are used as the grinding elements in pebble mills and the mills are lined with stone or porcelain blocks. Steel balls usually form the grinding medium in ball mills, and the mills may be lined with steel. Pebble mills are used to avoid contamination by metal of the material being ground, to avoid rapid wear due to corrosion, and for other specialized reasons. Belgian silex (silica) blocks formerly were used widely for lining pebble mills, because of their good resistance to wear, light color, and low cost, but the importation of these blocks was suspended early in the war. The stocks of Belgian silex diminished so rapidly that it was found desirable to compare the properties of a number of domestic substitutes for use in critical operations. Accordingly, a study was made to determine the best substitutes for fine grinding of white pigment and for certain heavy-chemical process work such as the disintegration of phosphate rock in weak phosphoric acid solution. A comprehensive survey of substitute lining materials was not contemplated. It was stipulated at the beginning of the investigation that only materials on the market, available in large quantities, would be tested and that these should be reasonable substitutes from a wear-resistance viewpoint as determined from known physical characteristics. Marble and limestone, for example, could be eliminated immediately. Quartzite, granite and porcelain were selected as likely substitutes for Belgian silex and were subjected to quality tests. Materials Tested 1. Belgian Si1ex.—Sample blocks of a standard grade were taken from crates in which they had been imported. The stone was light gray in color and contained inclusions of small clusters of quartz crystals as well as light colored portions that might be calcite. This was a siliceous rock which appeared to be of igneous origin. 2. Quartzite.—Samples of quartzite were obtained from a quarry in Minnesota. The blocks exhibited a crystalline fracture. The stone was pink in color and composed of fine quartz crystals bound together with a siliceous cement. 3. Granite A was a fine-grained granite quarried in South Carolina. 4. Granite B was coarse-grained granite quarried in South Carolina. 5. Granite C was a coarse-grained granite quarried in North Carolina. 6. Porcelain A.—Sample blocks of the pebble-mill lining material were obtained direct from the manufacturer. It was white in color and of lower specific gravity than the natural materials. It did not have a crystalline structure. 7. Porcelain B.—Sample pebble-mill lining blocks of the porcelain were received direct from the manufacturer. It was white

Jan 1, 1947

By Herbert H. Kellogg

Deposits of high-silica kaolinite clays occur at many places in central Pennsylvania. These white clays were formed apparently by weathering of argillaceous quartzite and limestone. Their geology, distribution and use have been reported by Leighton.1 The larger deposits have been mined from time to time, and the clay, after washing to remove coarse sand, has found use as rubber filler, foundry clay, and in white cements. In general, the clay from these deposits contains too much grit (quartz) to be useful as paper filler, and the silica content is too high to make it useful as china clay for ceramic ware. Gravity methods for the separation of quartz from these clays have failed because of the extremely fine size of the quartz grains, but by the methods of froth flotation described in this paper these sandy kaolins can be made to yield a purified kaolinite that has a chemical composition close to that of good-grade Georgia kaolins, and is free enough from grit to be used as a paper filler. As by-products of the separation, a fine ( — 325-mesh) sand assaying 95.5 per cent SiO2, and a very fine (—3-micron) clay are produced. The possible ceramic uses of these products are being investigated at The Pennsylvania State College. Experimental Procedure The Sample.—The clay sample used in these tests was obtained from a small mine near Tyrone, Pa. It had already been washed to remove coarse sand, and represented the finished product of this small mine and washery. Microscopic examination showed that the clay was composed almost entirely of the minerals quartz and kaolinite, and that the quartz was present mostly in particles larger than one micron, while the kaolinite particles generally were less than 20 microns in size. The kaolinite and quartz particles were physically separable by blunging the clay in water with a suitable dispersant. Preparation of Flotalion Feed.—By the addition of NaOH in an amount equal to 3 lb. per ton of solids, the clay was completely dispersed in water. Flotation feed was prepared by allowing a dispersed suspension (10 per cent solids) of the clay to stand for the period of time required for a 2.5-micron particle of quartz to settle from the top of the container to the bottom. After this settling period the suspension was decanted from the settled solids. The settled solids were redispersed in water and the separation was repeated once. The two fine fractions were then combined. Thus the clay was divided into a fine and a coarse fraction by size separation at approximately 2.5 microns.* The fine frac- . tion was found to have a considerably lower sand content than the raw clay (Table I), and was not treated further. The coarse fraction formed the flotation feed for all the tests described in this paper. A size separation of this type can be

Jan 1, 1948

By F. W. McQuiston, L. J. Bechaud

3.2-1. Introduction. Much has appeared in the technical literature on theory of sampling (18, 37), weight of sample required in relation to particle size (28), probability of error in sampling (3), and statistical analysis of sample data (20). Each industry, whether it be copper, iron ore, or coal, has its own standards and methods. All of the rather extensive theoretical treatments appearing in published literature on sampling apply to material already on hand such as coal, iron ore, or copper concentrates in rail cars, or in moving streams such as on conveyor belts. The attempt usually is to apply statistical techniques in order to improve reliability and precision in sampling, and to reduce differences between the analyses of seller and buyer. In contrast to this type of sampling, very little has appeared in the literature on sampling a virgin deposit from the metallurgical point of view. Perhaps it would he more descriptive to designate this point of view as "Gm-metallurgical," since geology is inextricably interwoven with metallurgy in gaining an understanding of the complexities of a deposit, eventually leading to a definition of mineable reserves, with the development of a flowsheet and engineering criteria for the planning of a successful and profitable operation. The degree of cooperation between these two departments often determines whether the processing plant is a success at the inception of operations, or whether expensive changes and modifications must he made, with the plant finally reaching rated capacity at some later date. The first concern of the metallurgical department should be to establish proper accreditation of the samples which will be used in any ore-testing campaign. It should he made mandatory, as part of operating management policy, for the responsible metallurgist to he aware of the significance of the samples and to become intimately knowledgeable in all aspects of metallurgical behavior and response to treatment of all varieties of ore in the deposit under consideration.

Jan 1, 1968

By N. S. Greensfelder

THE firing of explosive charges by electricity dates back to 1745 when a Doctor Watson is said to have used an electric spark for igniting gunpowder. His method failed in practical application because, to ignite blasting powder, heat must be applied for a period longer than the duration of an electrostatic spark; Wheatstone found that this does not exceed 14000 sec. The usual result, therefore, of an electrical discharge in gunpowder is to scatter the grains, without igniting them. Benjamin Franklin partly overcame this difficulty by confining the powder and crushing grains that lay in the line of discharge. In 1822, after the primary cell had been discovered by Volta, Robert Hare, professor of chemistry at the University of Philadelphia, noticed the heating of a short piece of wire placed in the circuit of a voltaic battery. Later work by Oersted, Faraday, Wheatstone, and Henley resulted in the development of the magneto blasting machines. Fulminates of gold and silver are thought to have been discovered by Cornelius Drebbel, about 1600. Pepys, in his diary for November 11, 1663, recounts a conversation with Doctor Allen, who told him about aurium fulminans "of which a grain ... . put in a silver spoon and fired, will give a blow like a musquett and strike a hole through the silver spoon downward, without the least force upward.1," This fallacious belief that explosives act downwards only, is still held by many who should know better. Fulminate of mercury was discovered by Howard in 1799. J. Shaw of Philadelphia is said to have invented a fulminate-of-mercury cap in 1814; his idea was later adopted by gunmakers in England and France. The principal advantages of

Jan 9, 1922

By E. H. Rose

THE present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: I. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. ANALOGY WITH GRINDING In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1944

By John J. Croston

IN the closing months of 1936 the copper industry gave every evidence that it was at last on the threshold of an improved era. At the beginning of the year prices stood at 9 ¼ ¢, which in itself was a great improvement over prices prevailing for the three previous years. Industry generally was showing signs of moving forward in this country, and business was definitely good in certain foreign quarters. The threat of war and preparations for rearmament in Europe and Asia acted as a further stimulus to purchasing. By midsummer the situation had so improved and stocks had been so reduced that the cartel of foreign producers (including American producers with foreign mines) decided to increase the production quota to 75 per cent on Aug. 1. Large-scale buying gave continued strength to the price and by Oct. 1 the quota was increased to 80 per cent, and again on Oct. 15, to 85 per cent. Notwithstanding this the price of copper continued its advance and on Nov. 1, with stocks declining continuously, the cartel increased operations to 95 per cent, then to 105 per cent on Nov. 5. In the meantime the domestic producers, because of the great increase in home consumption, had increased their production by nearly one-half. At the close of the year copper sold at 12¢ per pound, and since the turn of the present year has reached 17¢. This rapid increase has not been wholly justified. Much of the increase was due to the legitimate demands of recovering industry, but rearmament orders plus accumulation of reserve stocks in case of war, purchases in anticipation of labor strikes and purely speculative activities on the London Metal Exchange accounted for the balance. Certainly in view of the state of industry and the general world price level, it appears unnecessarily high. Talk of a copper shortage is sheer nonsense, the recent condition in the market being created by overbuying, leaving but a small volume for speculative supplies. It takes some months for an increase in mine output to reach the market, but production is again overtaking apparent consumption, stocks have begun to increase, and this, together with the reopening of many small

Jan 1, 1937

By Carl Zapffe

A Rise in less than ten years from obscurity into great prominence economically, tersely summarizes the history and status of the Cuyuna manganiferous iron ores. The Cuyuna district produces and ships mostly high-phosphorus iron ore, but the annual production of manganiferous iron ore is increasing rapidly and productions from individual mines are becoming more regular and more dependable as to grades and tonnages. This ore bids fair to become, shortly, the principal product of the district and is responsible for its present economic prominence. The iron ores of the district have been described, by the author and others, in many publications and will not be discussed here. Lately the manganiferous iron ores have been given some attention, but the range and variations of chemical composition of the ores, the principal physical characteristics of the ores and their deposits, the relative importance of the different ores and their possibilities in future developments, and the life of the respective reserves have not been discussed. It is impossible to give all of this information with the positiveness necessary for it to constitute a reliable guide for investment or even for future mine development. Each property and each party presents a special case and deserves critical consideration. We are dealing with special ores and always with a set of new conditions. Sufficient information of a general nature can be given, however, to indicate the reason for and the trend of the presently augmented developments in the district. In a great thickness of a big variety of sedimentary rocks, probably of Upper Huronion age, prevailingly of green, schistose phases of slaty and volcanic material intricately intermixed, lies a thin bed containing lenses of iron-bear ing formation. These lenses occur singly or in groups, thus causing the variation in the thickness of the formation; or they are separated at their ends and give rise to discontinuity or they overlap and display continuity with lateral offsets. Intense folding has added to the complications.

Jan 1, 1925

By Donald W. Scott

GENERALLY speaking, there is nothing very difficult about taking good micrographs of photogenic thin sections or grains with a petrographic micro-scopecamera setup. However, sometimes it is desired to show some detail that cannot be illustrated satisfactorily by the usual micrographs taken with parallel polarized light, with crossed nicols, or even with two micrographs. A simple procedure giving good results has been used at Battelle Memorial Institute for subjects falling into this classification, and it is believed to be of interest, especially to research people who must sell a point to executives through the medium of well-illustrated reports. The following problem illustrates a typical application: It is desired to photograph a thin section of an iron ore to illustrate the character and association of the iron oxides and the individual quartz crystals. A magnification of 75 dia. is acceptable. A micrograph of the field in question (Fig. I[a]) was taken with parallel polarized light. Due precaution was taken to provide critical illumination, correct exposure, a sharp focus, a good darkroom technique, and other essential conditions. It illustrates correctly the iron oxides, but it fails to show the individual quartz crystals. A micrograph of the identical field (Fig. I[b]) was taken under the same precise conditions, but with crossed nicols. It gives a better idea of the character of the individual quartz crystals but is likely to be misinterpreted unless accompanied by a lengthy descriptive note. This micrograph suggests that the centrally located, large black area is composed entirely of iron oxides. That this is an incorrect interpretation is evident by reexamining the micrograph of Fig. I[a]. The correct interpretation is that the dark area contains quartz crystals oriented with their optical directions parallel to, or nearly parallel to, the vibration directions of the polarizer and analyzer of the microscope. Thus, the executive, or anyone who had not done the work personally, would get no more than a fair interpretation, even if two micrographs were included; and in the interests of both brevity and clarity, one micrograph usually is more desirable. Exactly how many quartz crystals occur in this region cannot be told from the two micrographs, and a printing of the two negatives in register would likewise not show the detail desired. Thus, it is evident that some subjects cannot be satisfactorily illustrated by the usual procedures. In photographing Fig. is the light was subdued and, during the entire exposure of 75 sec., the analyzer nicol was rotated uniformly in its mounting from a crossed position to 45° uncrossed position to introduce the desired amount of contrast in the various individual quartz crystals. The true reproduction of the character and association of the iron oxides, and the additional structure and detail in the quartz areas, is apparent. It might be

Jan 1, 1946

By John P. Bryant

Plant nutrients are obtained by plants from both the air and the soil. Carbon dioxide, a gaseous form of carbon and oxygen, supplies the carbon which usually makes up 50% or more of plant structure. Plants are capable of absorbing carbon dioxide directly from the air. Most of this nutrient comes from the air, although some carbon dioxide is absorbed by soil water and taken into the plant as liquid. Growing plants use more water than any other substance. Plant nutrients in the soil dissolve in soil water; soil water entering the plant carries nutrients into the plant structure. Water is then given off into the air as the plant breathes or transpires. The inorganic or mineral portion of the nutrients which plants require comes from the soil. Inorganic or mineral composition of plants is usually less than 5% of their weight, and this mineral requirement, usually lacking in the soil, must be provided by fertilizers. Proper fertilization of a soil to improve its mineral level and productivity is the basis for the fertilizer industry (Thomas, 1960). There are 13 essential plant nutrients which often must be added to soils or plants and are generally considered as fertilizer minerals. These nutrients are divided into the categories: primary minerals (nitrogen, phosphorus, potassium); secondary minerals (calcium, magnesium, sulfur); and micro or trace minerals (boron, iron, manganese, copper, zinc, molybdenum, chlorine). Further, cobalt is often added to fertilizers. Although it is not considered to be an essential plant nutrient, it is essential to animal health and is often applied to forages as a fertilizer constituent. One point of major significance to the fertilizer industry is that not all nutrient carriers supply that nutrient in a chemical form which will be available to the plant. Thus, a product which may be considered a fertilizer mineral may require substantial processing or treatment before it becomes useful as a nutrient carrier. Fertilizer Industry Scope Until about 1900 demands for higher crop yields were met simply by bringing new agricultural lands into cultivation. Since that time great progress has been made in increasing yields through fertilization and improved varieties and cultural practices. The importance of the fertilizer industry to the economy and welfare of the U.S. is widely recognized. [Fig. 1] illustrates the relationship of lime and fertilizers to crop yields. In terms of tons of fertilizer products, the growth of the industry has been phenomenal, particularly between 1950 and 1970. Table 1 shows the increase in fertilizer usage. Primary Plant Nutrients in Fertilizers Primary minerals (nitrogen, phosphorus, potassium) are designated as such because of their significance to plant growth. Practically all soils are deficient in both nitrogen and phosphorus, and most of them are also deficient in potassium. Consequently, these nutrients

Jan 1, 1975

By John A. Emery, Frank G. Snyder

MINING geology has a threefold objective: to guide prospecting for new ore, to evaluate known orebodies as development risks, and to supply the detailed knowledge of ore structures necessary for more economical extraction. The role of geology in exploration is well established. This article will deal solely with orebody evaluation and the application of geology to development and mining problems as practiced by St. Joseph Lead Co. in the southeast Missouri district, which for many years has supplied approximately 30 pct of U. S. lead production. The primary concern of the present discussion is geology in the district, not geology of the district. The southeast Missouri district, locally called the Lead Belt, includes parts of St. Francois, Madison, and Washington counties. The first recorded mining was about 1720 at Mine La Motte in Madison County on surface showings of lead ore. Later discoveries near Bonne Terre, twenty-five miles to the north, led to extensive prospecting and underground mining. To date the district has yielded well over 8 million tons of pig lead, valued at more than $1 billion. By far the greater part of the lead production has come from an area of 50 sq miles, centering on the city of Flat River, as shown in Fig. 1. Role of Geology in Metal Mining The history of the use of geology in the older metal mining districts follows a simple consistent pattern that may be considered in terms of four distinct phases. Each geologic phase is a response to the economic status of the mining district; each has definite characteristics setting it apart from other phases; and each passes into the succeeding phase as the mining operation gradually exhausts the large, rich, easily discovered deposits and lives upon increasingly smaller, lower grade orebodies that are less easily found. The geologic phases and their major characteristics are as follows: Period of No Organized Geologic Program: During the early history of a great metal district ore is easily found by random prospecting and wildcat drilling. Many small operators are active, as profitable mining can be done with a relatively small capital outlay and a minimum of technical staff. Mining practice consists largely of sinking a shaft on an ore showing and following the ore in sight. When the ore is exhausted or becomes unprofitable to mine, the process is repeated on a new ore showing. The miners become familiar with certain ore habits and use this practical geologic knowledge to guide mining. The method is successful as long as large tracts of virgin mineralized area are available. This phase ends when small-scale mining methods are no longer profitable and the mines become consolidated into the hands of a few operators.

Jan 1, 1957

By Leslie G. McMillion

A field project to develop and demonstrate a method for alleviating problems of highly mineralized ground water where they occur as isolated zones or pockets in fresh-water aquifers is being conducted in the Ogallala aquifer of eastern New Mexico by the Robert S. Kerr Water Research Center. Removal of contaminated water is by a large-capacity well and the pumped poor-quality water is used locally for secondary recovery of oil. Benefits of the operation are threefold in that the aquifer is being reclaimed for future fresh-water production, contaminated water is being used beneficially, and fresh water which otherwise would have been used for waterflooding is available for other purposes. Ground-water quality in the site has improved markedly after two years of operation. Use of water in the arid and semiarid Southwest has grown so rapidly that demands in many localities far exceed available supplies. In the High Plains region of New Mexico, Oklahoma, and Texas, the only large sources of water are underground reservoirs, being depleted at an alarming rate. In some parts of this region, the aquifers contain pockets of highly mineralized waters; and the mineralized waters in some of them have started spreading in response to the drawdown effects of pumping wells. The aquifers of this region are primarily composed of sand and gravel beds; any cleanup by natural processes will require many tens or even hundreds of years. Most of these pockets of highly mineralized water were caused by oil-field brines seeping downward from earthen disposal pits. Thousands of these pits were in use in the region until only a few years ago when they were prohibited by the oil regulatory agencies of the respective states. The resulting areas of poor quality water are serious enough to warrant much corrective attention. Correction will involve one or more of three approaches: 1) Restrict the pumping of fresh water in the zones where the drawdown of the pumping well will significantly influence the movement of the contaminant. 2) Remove the highly mineralized water entirely from the affected area. 3) Develop and pump the poor-quality water in such a manner that its movement is held in check, even though only a part of the contaminant is actually removed. The principal hurdle to overcome in applying methods 2 or 3 is financing of the operations, since there are very

Jan 1, 1972

By John Birkinbine

IN the spring of 1878 the Pine Grove Furnace, located in Cumberland County, Pa., was blown in after lying idle for several years. The furnace was constructed in 1770, and for over a century it has been in almost continual operation. The plant consisted in 1877, when the writer was first called there, of a stone stack 32 feet in height, inclosing a shaft and boshes, the latter being 8 feet in diameter. The blast was supplied by two wooden blowing-tubs discharging into a third, having a floating piston sustaining a box weighted to give the desired pressure, the power being furnished by a waterwheel. A small 18-pipe hot-blast stove heated the blast. The furnace was remodelled during the winter of '77-78, and a Weimer blowing-engine with a blowing-tub 5 feet in diameter and 2 feet stroke, and the necessary boilers, were substituted for the wooden tubs and water-wheel. The stack was raised and enlarged, and provided with a bell and hopper having a central drop. Water dam and tymp were added, and general repairs were made. The hot oven, however, was riot increased. The furnace had always been operated with charcoal as fuel, and its reconstruction was made with a view to continue the use of charcoal, but provision was made for ample blowing capacity should other fuels at any time be employed. Although liberal arrangements had been made in cutting wood, the stock of charcoal supplied in 1878 was insufficient for the increased requirements of the remodelled plant, and notwithstanding the purchase of some 80,000 bushels from an adjacent idle furnace, there was riot enough to keep the furnace in constant blast until a new supply could be obtained, particularly as the season was unusually backward. Instead of following the established precedent of many charcoal furnace managers, i. e., blow out every February, it was determined to continue in blast, using coke as fuel. Accordingly, when the charcoal stock was exhausted, on March 22d, 1879, Connellsville coke was substituted, the change of fuel being made at once; that is, the coke charges followed immediately upon the last charcoal charges. After working a few days,-the strike in the Connellsville coke region cut off the supply, and anthracite coal was obtained, a mixture of the two fuels being employed. The strike continuing,

Jan 1, 1880

By J. Visman

A simplified method of assessing the characteristics of a cleaning unit, including washability curves, yield figures, ash error, separating gravities, and error curve. FOR more than 25 years evaluations of coal cleaning units have appeared in many publications, and during this time considerable progress has been made in developing the theory and techniques of calculating washery results. A variety of characteristics have been defined over the years, many receiving general recognition during the first post-war Coal Preparation Conference in Paris in 1950. Coal preparation has become a science, and the approach to preparation problems is being revolutionized. Formerly, for example, if a coal operator wanted to know the performance of a given cleaning unit for a particular grade of his run-of-mine, he either shipped a carload to the manufacturer for a test run or relied on a report of a test carried out on a coal similar to the one he intended to clean. In modern practice the expensive, time-consuming test work generally can be avoided if the washability curves of the coal in question and the error curve of the cleaner are known. Normally a guarantee can be given on the strength of this information. General acceptance of modern standards has thus reduced the need for pilot-scale models of existing types of coal cleaners. Needless to say, the method requires exhaustive and accurate information on the cleaning unit and on all the additional circumstances affecting its results, information that ultimately can be secured only from actual tests made in the laboratory or in the washery. It is not the purpose of the present discussion to compare the merits of various characteristics, but to describe the general method of calculation, including a method of obtaining data with a minimum of laboratory work. All data presented here were derived from tests made at the plant with portable field equipment, using small quantities of varsol, carbon tetrachloride, and bromoform. It will be explained later that the use of standard fractional ash contents designated normalized ash contents aids considerably in reducing the routine testing of coal cleaners. Like any other statistical procedure, this method has limitations, and it is not recommended for any work in connection with guarantees. Even a varying commodity such as coal, however, will follow the law of large numbers, which is the basic law of statistics, thus allowing simplification of test work without appreciable loss of accuracy. The characteristics needed for evaluating cleaner performance are normally derived from a float-sink analysis of samples of the clean coal and the refuse, followed by ash determinations of the various fractions. Usually a sample of feed is also collected to find its ash content and to make a single specific gravity cut, both in connection with determination

Jan 1, 1955

By J. B. Austin, R. H. H. Pierce

IN developing a consistent picture of the thermal expansion of the commercial chromium steels, which cover a range of, carbon contents, it is desirable to know the expansion in the limiting case of the binary, carbon-free, iron-chromium alloys. We have, therefore, determined the thermal expansion of a series of five special alloys, ranging in chromium content from 3 to 10.5 per cent, with not more than 0.03 per cent carbon, over a temperature range extending beyond the transformation region; this low carbon content enables us to draw reasonably reliable conclusions as to the behavior of the carbon-free system. Moreover, the measurements serve to determine the beginning and the end of the alpha-gamma transformation; and they emphasize once more the fact that a transformation temperature determined by any method in which the specimen is .cooled at an appreciable rate is apt to be seriously in error, particularly if the metal contains a transformation-retarding element such as chromium or managanese. The lag, or undercooling, observed at the ?-a transformation is explained in terms of the influence of temperature on transformation velocity; the rather unusual shape of the ? loop is attributed to a change with temperature and with composition of the distribution of chromium between the alpha-iron and gamma-iron solid solutions. The exact composition of the alloys which were obtained through the courtesy of Dr. W. Rohn, of the Heraeus Vakuumschmelze A-G., is given in Table 1. Determinations of the linear thermal expansion and of transformation temperature were carried out in the vacuum dilatometer previously described,1 using a rate of heating and of cooling of 3.0° to 3.5° per minute unless otherwise stated. Magnetic susceptibility, which was measured in studying the rate of transformation, was determined with a magnetic balance, which also has. been described.2

Jan 1, 1935