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By C. E. Beecher

The application of gas or air under pressure to obtain more oil from a sand which has been practically exhausted by ordinary production methods has been practiced to a limited extent for many years. Until recently such methods were confined to the Eastern fields, principally those of Ohio and West Virginia, where in many cases the production obtained after applying pressure for a few years was nearly equal to the total production prior to that time which had been obtained by the usual production methods. As early as 1917, J. 0. LewisL called attention to this method of increasing the recovery of oil and cited the phenomenal results obtained by H. E. Smith, I. L. Dunn and 0. C. Dunn, who pioneered the application of air or gas pressure recovery methods. With this information available it would appear that other oil-producing sections of the country have been extremely slow to realize the advantages of repressuring. However, after a few successful projects in the Mid-Continent area, the oil producers appreciated t' e value of this method for obtaining further and profitable production from their practically exhausted leases and are now making extensive use of it. In general, it has been the practice to apply pressure to the sands when production has declined to a small yield per well. The purpose of this paper is to discuss some of the advantages which may be expected from returning gas under pressure to the oil sands during the early stages of development, before the rock pressure has been dissipated and the gas exhausted, and to stimulate discussion of this subject on which comparatively little actual field data are available. The Function of Gas Gas associated with or dissolved in the oil is the source of energy causing movement of the oil to the well. When this energy is dissipated production ceases. The exception to this general statement is found in fields where hydrostatic water pressure may supply energy after the gas

Jan 1, 1929

By C. Chu

A theoretical investigation has been made of the forward combustion process using a one-dimensional linear mathematical model, taking into consideration the effect of the vaporization-condensation which occurs on the leading edge of the heat wave. This work involves the solution of five coupled partial differential equations. Besides the vaporization-condensation phenomenon, these equations account for conduction, convection, combustion, heat loss, diffusion and bulk fluid flow. For the one-dimensional linear model studied, the vaporization-condensation phenomenon does not induce appreciable change in the temperature at the combustion front; and its primary effect is to create a steam plateau and to increase the length of the heated zone ahead of the combustion front. This effect becomes more pronounced at lower pressures, higher porosities or reduced gas saturations. The peak temperature and the temperature profile on the leading edge of the heat wave stabilize after a certain period. The length of the steam bank remains practically constant, although the length of the water bank increases as the heat wave advances. INTRODUCTION The existence of the vaporization-condensation phenomenon in the heat-wave process and the important role played by the phenomenon have been recognized by several investigators. Kuhn and Kochl stated that steam plateaus were frequently observed on the temperature records of their experiments. The steam plateaus were attributed primarily to the vaporization and subsequent condensation of the interstitial water existing in the oil sand. Szasz2 suggested that both lighter hydrocarbons and water are vaporized on the leading edge of the heat wave, carried forward in the gas stream, and then condensed to create banks of oil and water. Martin et al suggested that the vaporization-condensation phenomenon is one of the main mechanisms of the heat-wave process, along with thermal expansion and viscosity reduction. Wilson et al. reported the existence of a steam plateau several inches in length in their small-scale tube-run experiments. However, this important phenomenon has never been taken into consideration in the numerous theoretical analyses by various authors. 5 The purpose of this work was to study the thermal aspects of a linear heat wave, taking into consideration the vaporization-condensation on the leading edge of the wave, to determine the effect of this phenomenon on the temperature profile of the reservoir, and to investigate how this effect varies when other process variables are changed. THEORY We consider a reservoir of porous medium of cross-sectional area A, extending from x = 0 to x = L. This reservoir contains, aside from the solid matrix itself, a gas phase and a "combined liquid phase" which is a combination of two immiscible liquid phases — namely, an oil phase and a water phase. The oil present in the reservoir is assumed to consist of three fractions, a noncondensable gas, a nondistillable residuum, and a vaporizable oil fraction which may be present in both the gas phase and liquid phase. Before the heat-wave process begins, preheating has taken place and has imparted an initial temperature distribution T,(x) to the reservoir. At the start of the process, a stream of oxygen-bearing gas is introduced through the face at x = 0. This gas supports the combustion of the residual fuel and supplies the heat throughout the process. In order to develop a tractable system of equations, the model properties are postulated as follows: (1) the system is isotropic and homogeneous; (2) there is a uniform fuel content throughout the reservoir; (3) combustion takes place in a moving vertical planar region of infinitesimal thickness; (4) the temperatures of the fluids and the solid matrix are identical; (5) radiation is negligible; (6) the gas and liquid saturations are uniform throughout the reservoir and remain unchanged

Jan 1, 1965

By Hoyt Stoddard Gale

Recent work on borate deposits near Kramer in the extreme southeast corner of Kern County, California, is of special interest because of the information it seems to give concerning the mode of origin of the borate minerals colemanite and ulexite, especially when considered in conjunction with the valuable clue to the mode of origin of colemanite in its present known deposits given in a paper by W. F. Foshag,1 of the United States National Museum. Colemanite is found in or near original deposits of ulexite, from which it is supposed to have been derived by a change in the character of the ground water or rate of its movement or both. Indications that an alteration of this sort had produced colemanite were first recognized by the writer2 in specimens collected in the Kramer district in 1920, and sent to W. T. Schaller, at U. S. Geological Survey, for identification. The identity of the two borate minerals in these specimens was suspected when collected but not at once determined, but the relations of the minerals as seen in the specimens seemed to afford a manifest clue to the origin of the deposits. Similar specimens were obtained by the writer from deposits in Death Valley soon after he had examined the deposits near Kramer. After studying the specimens submitted to him Dr. Schaller wrote as follows: " I have finally positively identified the boron minerals you sent me with your letter of Sept. 27. The 'white fibrous mineral like ulexite, is ulexite; the 'harder nodules' in the clay are colemanite, apparently formed from the ulexite as a result of leaching, with recrystallization in place—a theory that Foshag is working on." The present paper is a report on the facts observed at the depo its near Kramer, and some general observations on the geological relations noted, which, unfortunately, still form only a very incomplete record. The intimate association of deposits of colemanite and ulexite in California and Nevada with certain volcanic rocks affords a strong pre-

Jan 1, 1926

WE the undersigned, being ell persons of full age and citizens of the United States, and a majority resident of the State of New York, desiring to form a corporation pursuant to the provisions of the membership Corporations Law for the purpose of incorporating, as provided in section 5 of Article I. of said law, the existing unincorporated association known as American Institute of Mining Engineers, do hereby make, acknowledge and file this Certificate for that purpose, and do Certify as follows: I. That the American Institute of Mining Engineers is an unincorporated association organized and existing with the object of promoting the arts and sciences connected with the economic production of the useful minerals and metals and the welfare of those employed in these industries by means of meetings for social intercourse and 6he reading and discussion of professional papers, and to circulate by means of publications among its members and associates the information thus obtained. II. That the persons duly appointed or designated to manage the affairs of said association are designated by the rules thereof Members of its. Council; that the undersigned are all members of said Council as the same was constituted on the 296h day of December, 1904. III. That on said last-mentioned date a regularly called meeting of said association was held at its office in the Borough of Manhattan, City of New York; that thirty days before such meeting notice of the intention to incorporate said association was given by mail to each member thereof whose residence or post-office address is known; that at said meeting the following resolutions were offered, seconded and duly adopted by the unanimous vote of all its members then present, to wit: "Resolved, That it is the sense of the members and associates of the American Institute of Mining Engineers in general meeting assembled that it is desirable and necessary for the well being of said association and its members and for the furtherance of the objects for which the same has been formed, that said association incorporate under the Membership Corporations Law of the State of New York; "And Further Resolved, That the Members of the Council of this Association, or a majority thereof, be and they hereby are authorized, in accordance with the provisions of Section 5 of Article I. of the Membership Corporations Law, to incorporate this association for the same purposes for which it has been organized and conducted, in the manner provided in Article 11. of said law; "And Further Resolved, That the name of said corporation as hereby adopted by this meeting shall be American Institute of Mining Engineers; "And Further Resolved, That the said incorporators shall be named in the Certificate of Incorporation as directors of such corporation until its first annual meeting, and that such directors and their successors in office shall be and they hereby are authorized to enact and adopt a Constitution and By-Laws for the government of said corporation." IV. That the name of the proposed corporation is American Institute of Mining Engineers. V. That the purposes for which this corporation is to be formed are: To promote the arts and sciences connected with the economic production of the useful minerals and metals, and the welfare of those employed in these industries by all lawful means; to hold meetings

Jan 1, 1928

THE Institute assembled in the rooms of the Geographical Society at the Cooper Union, Tuesday evening at 8 P.M. President Raymond called the meeting to order, and introduced Mr. Abram S. Hewitt, who welcomed the Institute to New York. President Raymond, in behalf of the Institute, replied. The President then read the following report of the Council for the past year. The Council respectfully report that the American Institute of Mining Engineers has held four meetings-at Wilkes-Barre, Bethlehem, Troy, and Philadelphia, respectively-all of which have been well attended, and socially and professionally most successful. The proceedings of these meetings having been published in the Engineering and Mining Journal, and sent to all members and associates, are sufficiently well known. Forty-Bur papers, by twenty-six authors, have been read. These papers embrace a wide range of subjects, and prove the practical character of the proceedings of the Institute. The finances are in a satisfactory condition. By the report of the Treasurer it appears that $1290.10 have been received, $519.90 expended, and that $770.20 now remains in the treasury exclusive of fees for 1872. This fund, it is hoped, will be so increased by voluntary subscriptions as to enable us before long to issue a volume of our most important transactions in a permanent form. The Council has made a favorable arrangement with the Scientific Publishing Company, publishing the Engineering and Mining Journal, by which that journal became the organ of the Institute for the first publication of the proceedings, papers, and notices to members. This arrangement it is proposed to continue through the present year unless the Institute shall otherwise direct. I t is not intended to supersede the publication of an annual or occasional volume, but to maintain an interest in the Institute, and to serve the convenience of members while our funds do not permit a more ambitious publication.

By H. L. Talbot, R. S. Young, A. Golledge

The differentiation of oxidized forms of lead from lead sulphide in complex products by chemical analysis is of considerable importance to certain mining and metallurgical companies. A method for the separation of "oxide" lead, (non-sulphide), from sulphide lead, based on the solubility of the former and virtual insolubility of the latter in concentrated ammonium acetate solution, has been employed for many years in the industry.1,2 Unfortunately this solvent, while satisfactory for lead sulphate and carbonate, is without action on lead phosphates or vanadates, with the result that at .Rhodesia Broken Hill Development Co. where these latter minerals are present, they have counted as sulphides when using this procedure. For the control of operations in the new concentrator it became necessary to have a closer check on sulphide lead than could be obtained by the only current method of differentiating this constituent from oxidized lead. For some time this problem has engaged our attention, and after a large number of solvents had been tried, the following procedure was evolved. Like all methods depending on selective solution in a com- plex material, it will not give 100 pct separation into oxide and sulphide portions. Because of the ease of oxidation of galena this is almost an impossibility. Our procedure will, however, give a recovery of 99-100 pct of the oxidized lead, contaminated with very small quantities of lead sulphide, and a corresponding recovery of 97-98 pct of the sulphide lead. The procedure is simple and reproducible, and gives results quite .satisfactory for routine control purposes. Experimental For this work standard samples of the following pure lead minerals were obtained from Broken Hill, Northern Rhodesia. These were hand-picked specimens, and were carefully ground to —200 mesh. Since it was realized that a fair-sized sample would be necessary for this work to thoroughly test the large number of procedures we had in mind, a compromise had to be effected between a few grams of very pure minerals and a large sample of lower grade material. For this work the presence of gangue was not deleterious, so long as the oxide lead minerals did not contain lead sulphide, and vice versa. Table I shows the lead minerals used, and their total lead content, both theoretical and actual. Apart from the pyro-morphite the minerals are seen by analysis to be quite pure. Many solutions were tried in an effort to dissolve the oxidized lead minerals and

Jan 1, 1949

By C. E. Sims, F. W. Boulger, H. A. Saller

Most of the data on equilibrium constant and the deoxidations potentialities of those elements, considered to be stronger deoxidizers for steel than is silicon, have been calculated from thermodynamic data. The reason for this is, primarily, the obvious difficulty of obtaining direct experimental evidence of equivalent accuracy. This is an excellent use of the principles of thermodynamics and has given valuable data not otherwise available. Such results, of course, can be no more accurate than the physical constants used in the calculations, and one can never be sure that the basic data are either complete or accurate. In fact, as in the case with silicon,1 there are not only discrepancies among the calculated theoretical values of the equilibrium constant for deoxidation of steel but also between the theoretical and experimental values. It is highly desirable, therefore, to obtain experimental values for checks on calculated results whenever possible. If they disagree, both cannot be right, but if there is good agreement, their value is enhanced. The present work was done in an effort to obtain experimental evidence in regard to some of the common alloying additions but more particularly the so-called "strong" deoxidizers for steel. The method used was to determine the minimum concentration of the deoxidizer that would effect a certain definite degree of deoxidation in steel. The criterion of deoxidation was the change from the large globular Type I sulphide to the eutectic Type II as described by Sims and Dahle.2 This change is sharp and definite, and inasmuch as it can be produced with equal facility by aluminum, zirconium, and titanium, it is considered a manifestation of a certain degree of deoxidation and not an alloying effect. Ostensibly such a procedure could give only a comparison of deoxidizing powers and no absolute values. Nevertheless, repeated observations have shown that, when increasing increments of aluminum are added to a steel, the residual aluminum content begins to increase simultaneously with the appearance of Type II inclusions. Thus it seems warranted to postulate that the Type II inclusions appear coincident with the virtual elimination of FeO as an active constituent of the steel. Experimental Procedure The data obtained were primarily from the microexamination of polished and unetched specimens and from chemical analysis. Experimental heats weighing 200 to 250 lb were made in a basic-lined high-frequency induction furnace. The base composition was nominally that of a medium-carbon casting steel to which the appropriate additions were made. Specimens were poured into sand-cast ingots 3 in. in diam as shown in Fig 1. Sand-cast ingots were used to prevent chilling and to allow sufficient time in freezing for normal inclusions to form of a size large enough to be studied readily. In the first few heats, the tapered wall ingot was used, but in the majority, the extra large riser was used to prevent piping in heavily deoxidized steels. Specimens for microexamination were taken from the location shown in Fig 1, and drillings for chemical analysis were taken from a similar location. The procedure was to melt the base composition and deoxidize with the usual manganese and silicon additions and then to pour an ingot. The furnace was then tilted back, and the first increment of strong deoxidizer or special alloy was added and allowed to disseminate through the melt, with enough power on to hold the temperature constant, for 45 sec. Then a second ingot was poured. After this, another increment was added, and after the same holding time another ingot was poured. In this way from 9 to 12 ingots were poured from each heat, each successive ingot having progressively larger total additions of alloy. Eighteen heats were made altogether, and the range of alloys used and additions made are outlined in Table 1. The three principal types of sulphide inclusions found are illustrated in Fig 2. The globular Type I sulphides are characteristic of silicon-killed steels, the eutectic Type II are characteristic of steels deoxidized with a small amount of aluminum, while the larger, angular Type III are usually found in steels with a residual aluminum content above about 0.02 pct. In all specimens studied, the transition from Type I to II either did not occur at all or was very abrupt and clear cut. There never was any doubt as to just which increment produced the change, although the individual additions were small, in the order of 0.01 pct. The change from Type II to Type III was considerably less sharp, and, in some cases, both types were found together. Inasmuch as the formation of Type III sulphides is apparently not a deoxidation phenomenon, they will not be discussed here.

Jan 1, 1950

By Charles C. Hawes

DETERMINATION OF SILICA OXIDE iron ores usually are found in intimate association with silica and phosphorus compounds. Their quantitative separation and purification sometimes present difficult problems to the chemist. Silicic acids may represent various degrees of dehydration. They are exceedingly weak, but combine with metallic acids and form some of the most stable oxides in nature, the silicates. When iron ores are treated with hydrochloric acid, they yield practically all their iron. However, apart from rare exceptions, varying amounts of insoluble residues of silica, alumina, lime, magnesia and compounds of phosphorus, sulphur, titanium, chromium and others will be left. If the oxides of silicon, aluminum and titanium are present in small quantities, the insoluble residue is generally very small, but the presence of these oxides more usual in larger quantities. Therefore, only a special treatment can break the bond between these elements. Ore Residues Acids remove the soluble constituent of the ore, leaving the insoluble residue, which chemically is rock or gangue and must be treated accordingly. Alkali carbonates are used frequently to break up the residue, but other fluxes have also been recommended, such as a number of lead and boron salts. The introduction of these salts, which offers certain advantages, requires, however, their removal before the other constituents can be analyzed, and this elimination is sometimes difficult. Usually in quantitative analysis silicates are dissociated by fusing them in a platinum crucible with sodium carbonate, thereby forming sodium silicate, Na2SiO3, sodium aluminate, Na2A1204, sodium titanate, Na2TiOs, and carbonates of other elements. All the carbonates are soluble in hydrochloric acid, but also some silicic acid is formed that must be rendered insoluble. To this end the solution is evaporated several times, as the first evaporation still leaves as much as 5 per cent of the silica soluble. Even after a number of evaporations, minute quantities of silica may be found with the iron and alumina present. Perchloric Acid as an Agent for Dehydrating Silicic Acid Perchloric acid, HCI04, (60 per cent), has been used for years as a solvent, but few chemists have used it to dehydrate silica. It is derived from anhydrous perchloric C1207, which is most unstable and dangerous. Perchloric acid is a strong oxidizing agent, which boils at 203°C. It is more effective in rendering silica insoluble than any other known agent, and most rapid in action. After perchloric acid is added to a mixture containing silicic acid, the resultant solution is

Jan 1, 1945

By E. T. Peters, L. E. Tanner

In the course of a study concerned with the rnechanical behavior of the intermetallic compound CO3V, both structural observations and anomalous electrical-resistivity behavior indicated the presence of a new high-temperature phase in an alloy with the stoichiometric composition, Co-22.4 wt pet V. The evidence, which is given below, suggests that the new phase is situated between the low-temperature ordered hexagonal phase, k,1, 2and the high-temperature disordered fee solid solution, a,3 within a small range of temperature in the vicinity of 1070°C. Structural studies included X-ray diffraction analysis* and electron-transmission microscopy examination* of stoichiometric composition speci- mens homogenized for 1 hr at 1200oC and either air-cooled or water-quenched. The air-cooled specimens yielded X-ray diffraction evidence of the k phase, while electron microscopy revealed a microstructure of both heavily and moderately faulted regions as shown in Fig. 1. The high density of stacking faults (or fine twins) in region A is reminiscent of hexagonal cobalt following its allotropic transformation.4 Selected-area diffraction from this region revealed distinct superlattice reflections of the ordered hexagonal phase, as well as continuous rel-streaks which were normal to the fault planes in the micrograph. The latter is indicative of extremely small spacings between the faults.' Regions such as ß were found to be only partially transformed to the k phase. Attempts to study the antiphase domain structure by either bright- or dark-field techniques were unsuccessful in these initial studies. This was due to the low intensity of the superlattice reflections arising from the closeness of the cobalt and vanadium scattering factors. Quenching Co3V from above Tc appears to suppress the formation of the k phase. Both X-ray and electron-diffraction results indicated that the disordered phase, a, was partially ordered during severe cooling and that the structure of the ordered phase was of the L12 (Cu3Au) type,6 designated here as a'. As in the case of the hexagonal phase, examination of the domain structure was unsuccessful; however, significant information concerning the morphology of this condition was revealed by the appearance of dislocation arrays. In particular, it

Jan 1, 1965

Jan 1, 1970

By Walter R. Hibbard

Whether we have occasion to think of it or not, an adequate and economic supply of energy fuels is vital to each of us. It is interesting to look ahead at the way in which fuels from supplementary sources will begin to enter into the energy mix to help meet the long-term national energy needs. From 1947 to 1965, the average annual rate of growth of energy consumption was 2.7%, with total energy demand in 1965 being equivalent to 54 quadrillion Btu's. During this short period of 18 years, the cumulative output of coal in the U.S. was 9.6 billion tons; natural gas, 193.5 trillion cu ft; and crude oil, 45.9 billion bbl. During the next decade and a half the nation's energy requirements are expected to grow at an average rate of 3.3% yearly, with consumption of energy expected to reach an annual rate of about 88 quadrillion Btu's by 1980, or 63% over the 1965 level.

Jan 12, 1967

By Benjamin C. Koskiniemi

Introduction When evaluating any ore body, one of the first questions concerns the ore reserves. In the case of an open pit mine, this is not possible to answer reliably until the ultimate (final) pit limits have been established. Techniques used in designing an ultimate pit are classed as (1) manual, (2) computer, and (3) combined manual-computer. This chapter will describe how manual techniques can be utilized in designing an ulti¬mate pit Certain economic and design criteria must be established before the actual design begins In order to begin designing an ultimate pit, it will be assumed that the engineer already has the following data available (1) vertical sections, (2) horizontal sections for each level, (3) stripping curve, (4) bench height, (5) bank slope angle between levels, (6) level berm width, (7) roadway width, (8) pit slope angles at ultimate pit limits (estimated average including roads and ramps and between roads and ramps), and (9) minimum width of pit bottom

Jan 1, 1979

Jan 1, 1963

By R. Kossowsky

USING etch-pit techniques, Johnston and Gilman' have shown that the stress dependence of dislocation velocity can be represented by the expression where r, is the dislocation velocity, a is the applied stress, and oa, is the stress to produce a unit velocity. Noble and Hul12 showed that the exponent m is strain-dependent and increases rapidly in the work-hardening region. This was explained by arguing that in the work-hardened material the applied stress required for any given velocity will be greater than in the annealed material by an amount equal to a back stress due to work hardening. schadler3 has stressed the importance of substructure as a possible source of error in dislocation velocity measurements, and took care to make velocity measurements in which dislocations moved such short distances that interactions with boundaries were avoided. schadler3 also pointed out that, if standard testing equipment is used, measurements of m from flow stress data could be seriously in error in materials with a high initial work-hardening rate. Noble and Hu112 have further argued that, due to the strain dependence of m , the extrapolation to zero strain4 to obtain the true value of m can be justified only in the preyield region. Microstrain experiments in high-purity iron5,6 have led to the conclusion that aiA (Ref. 5) (defined as the stress to first produce a permanent set of 10-6 in. per in.) and oflow are not controlled by the same deformation mechanism. It has been proposed5,6 that aA is governed by the lattice resistance to the motion of screw dislocations in a given slip plane and that oflow is controlled by the lattice resistance to a full-scale cross slip. The purpose of this note is to extend Noble and Hull's argument further and to show that measurements of m from flow stress data can be in error due to the rapid increase in stress in the microstrain region6 when stress is increased beyond aA. Double z-one-refined single crystals of molybdenum (with the (101)[ 111] slip system operating and a Schmidt factor of 0.475) were used in an experimen-

Jan 1, 1968

By T. L. Joseph

Sulfur balances calculated from operating data covering thirty-five blast furnaces indicate that approximately 92 per cent. of the total sulfur charged enters with the coke, 7 per cent. with the ore, and 1 per cent. with the stone.. Most of the sulfur is carried from the furnace by means of the slag, the total amount eliminated being the product of the slag volume and the percentage of sulfur it contains. Additional sulfur in the charge must be eliminated either by increasing the percentage of sulfur in the slag or the quantity of slag. From relations found between slag volume and coke consumption and sulfur in the coke and slag volume, curves have been prepared showing the relative value of blast-furnace coke as its sulfur varies from 1 per cent. CHARCOAL was the predominant blast-furnace fuel until 1838, when it was found, by the operation of a 2-ton experimental furnace, that anthracite could also be used. This information was a stimulus to the iron industry, particularly in eastern Pennsylvania where anthracite was available at low cost. The production of anthracite pig iron increased steadily and, in 1855, its tonnage exceeded that of charcoal iron. Of the 680 furnaces in operation in 1874, the 226 using anthracite produced nearly one-half the total tonnage. In 1869, the production of coke furnaces exceeded that of charcoal furnaces and, in 1875, coke surpassed anthracite as a metallurgical fuel. In 1907, the output of anthracite pig dropped well below a million tons and its present significance is largely historical. In recent years, the production of coke has gone hand in hand with the production of pig iron and it has gradually attained the position it now occupies as a universal blast-furnace fuel. Charcoal, however, has great merit as a blast-furnace fuel chiefly because of its freedom from ash and sulfur. The harmful effect of the latter and the necessity of keeping the percentage of this element in pig iron at a minimum are well recognized. Other contaminating elements can be reduced to specified limits in the steelmaking process, but the control of sulfur belongs essentially to the blast furnace. The purpose of this paper is to point out the distribution of this impurity in blast-furnace

Jan 9, 1924

By Chester A. Fulton

APATITE, the most abundant crystalline phosphate mineral, is found in igneous rocks and probably is the primary origin of all other phosphates, whether mineral or organic. Its chemical formula may be expressed as Ca5(Cl,F) (PO4) 3. The chlorine may be replaced almost entirely by the fluorine. It is found in large or small hexagonal prisms, usually of a vitreous luster and a green or red color, but also violet, white, or yellow. It is found largely as an accessory mineral in granitoid rocks. Its hardness is 4.5 to 5 and its specific gravity is about 3.2. There are many other phosphate minerals-for instance, amblygonite, autonite, collophane, lazulite, monazite, pyromorphite, torbenite, triphylite, turquoise, vivanite, wavellite-with base elements of calcium, magnesium, iron, aluminum, manganese, lead, cerium, lanthanum, lithium, or several of these in combination and with various amounts of water of crystallization. However, these phosphate minerals have but little value save as specimens for mineral collections, the exception perhaps being turquoise. While large deposits of igneous apatite are mined today for the production of fertilizers and chemicals, more than 90 pct of the world phosphate production is derived from secondary apatites, called phosphorites. These calcium (and magnesium, aluminum, and iron) phosphates have been formed in the alteration of sediments by solutions containing phosphoric acid. Such solutions have come from the weathering or leaching of igneous rocks containing apatite, bone deposits of prehistoric animals and marine life, or guano deposits (excreta of birds). These phosphorites are mostly alterations of phosphatic limestones that took place slowly over many years. Whether it is apatite or phosphorite that is being produced, it is known in the trade as "phosphate rock." The sedimentary phosphate deposits are of marine origin. Some, as laid down, were rich in phosphate or have become so by subsequent subaqueous action and have not been further enriched since emergence from the sea. Some were originally phosphatic limestones or marls from which calcium carbonate was dissolved away by waters containing carbon dioxide, in which calcium carbonate is much more soluble than

Jan 1, 1949

By C. W. Parmalee

The ideal fuel for burning ceramic wares is the one that, among other characteristics, has little or no sulfur. For that reason wood was long considered the most desirable fuel but its high cost has practically eliminated its use except in regions where it is very abundant or other circumstances warrant its choice. The fuels commonly employed are natural and producer gas, oil, and coal. The last named is undoubtedly the most important, not only in this country but abroad. Occasionally anthracite is used but the larger part is bituminous. The quality of the coal used depends on the kind and the value of the products being manufactured. For some materials, such as common brick, hollow block, etc., the quality of the coal is of less importance than in the manufacture of polychrome terra cotta, pottery, porcelain, etc. It is obvious that in burning clays that contain notable amounts of pyrite, marcasite or gypsum, as is the case with many common clays, the sulfur content of the fuel is of much less importance. The permissible amount of sulfur in the coal used will depend on the kind of product being manufactured. "A good sample1 of coal commonly used in pottery ovens has something like this quantity of sulfur: total sulfur, 1.20 per cent.; sulfur in ash, 0.11 per cent.; volatile sulfur, 1.09 per cent." This is based on English pottery practice. Correspondence with a. number of leading manufacturers in this country has produced the following figures: Sanitary Ware.—(a) 1.0 per cent, maximum, (b) 0.5 per cent. Sewer Pipe.—(a) 1.2 per cent., (b) 1.1 per cent, present run, 3.0 per cent, has been used. Terra Cotta.—(a) 0.5 per cent, basis of contract, 1.0 per cent, approximate. Pottery.—1.0 per cent, contract, 1.5 per cent, probable content. Enameled Brick.—1.3 per cent, maximum, (a) and (6) designate different firms giving figures. Since a pottery kiln for sanitary ware will require about 14 tons of coal for a bum, the quantity of SO2 that will be generated from a 1-per

Jan 1, 1920

Not many accounts of mining methods or labor conditions in the mines in early days can be found; all that have seemed to be authentic have been included in the preceding chapters. In practically every field opened before 18.10 the first coal was recovered by quarrying, or strip mining, the operations, of course, being by hand and on a very small scale; new places were opened as soon as mining became difficult or expensive. Where it became necessary to drift into the hillsides, the first coal was taken out by sleds, and later by small wagons, the power being supplied by the miner, often assisted by dogs. Wooden rails then came into use with small cars running on them. These cars were very small, usually holding only a few bushels; in the Belleville, Illinois field as late as 1863 cars having capacities of 10, 12, 14 and some as large as 30 bushels were in use. It was not until late in the 1830's that horses or mules came into use for haulage purposes in the Pittsburgh region, although some had been used in the Richmond, Virginia field at an earlier date. Coal was usually undercut by hand and wedged down, until explosives came into general use, and in nearly every field the coal was forked into the cars, the fine coal, or slack, being left in the mine. Hours worked were very long, but few records remain to show what they were at different places and times; there was undoubtedly a great variation between districts and mines. Roy stated a miner could often produce 5 or 6 tons in a day of 8 or 9 hours in a 5 or 6 foot seam, but in another seam possibly could mine only 3 1/2 to 4 tons. He estimated an average day's work in a four foot seam at about 3 1/2 tons.1 In a strike in the Belleville field in 1863 it was found that an average skilled miner in the field could probably produce 4 to 5 tons a day.2 This was only the coarse coal loaded and did not include slack. In 1873, a coal undercutting machine was tried at a mine about two miles from Brazil, Indiana, said to be the first coal machine ever invented in the United States. It was run by a 5 H.P. steam engine (later to be changed to compressed air), and consisted of an iron rim, four feet in diameter with movable steel teeth placed about one foot apart. The rim

Jan 1, 1942

By Paul L. Cloke, S. R. Titley

Paul L. Cloke (Assistant Professor, Dept. of Geology and Mineralogy, The University of Michigan, Ann Arbor, Mich.) - On examination of the paper by S. R. Titley it appears that several errors have been made, many of them of a fundamental nature. It is the intent of this discussion to point out these mistakes, and, insofar as possible with the data at hand, to correct them. EXPERIMENTAL WORK Ferrimolybdite-Limonite-Water: The results of the experiments described are very difficult to interpret; nevertheless, a correct interpretation is essential as is pointed out subsequently. The difficulties become apparent when reasons are sought for the continuing change of pH after conversion of ferrimolybdite to ferric hydroxide and for the change of pH when -stirring is stopped after a steady value has been reached. In seeking an explanation of this experiment it seemed to this writer that the (temporarily) steady pH values might be due to kinetic rather than equilibrium effects. In this view the steady pH on stirring might be explained as a balance of several reactions proceeding simultaneously. The first of these would be the slow solution of ferrimolybdite,

Jan 1, 1964

By Walter Crafts, C. M. Offenhauer

IN a previous study' of the carbide phase of chromium steels, it was shown that chromium carbide (Cr7C3) is a more stable carbide than cementite (Fe3C) at tempering temperatures above about 500°C. in quenched steels containing from about r1to 7.5 per cent chromium. Below this temperature the carbide phase was found to consist of cementite in all of the steels examined. In the present investigation it was found that a similar change in the carbide phase takes place in molybdenum and in chromium-molybdenum steels at approximately the same tempering temperatures. The approximate ranges of occurrence of the alloy carbides Cr7C3, Cr4C, and Mo2C have been determined in steels containing up to 5 per cent chromium, 1 per cent molybdenum and 0.60 per cent carbon. The experimental steels were made in small high-frequency furnaces from Armco ingot iron, low-carbon ferrochromium and calcium molybdate and were poured into 2-in. ingots, which were forged to 1-in. round bars. Sections of the bars were heated to the carbide-solution temperature and cooled rapidly to induce austenite transformation at relatively low temperatures. The steels were subject to some segregation, therefore comparatively high temperatures were necessary to dissolve all of the carbides. The microstructure of the quenched specimens was examined to verify the belief that the carbides had been dissolved, and also to determine the nature of the nonmartensitic structure of incompletely hardened steels. Ferrite and pseudomartensite were present in some of the specimens, but all specimens were quenched at such a rate that pearlite was not formed, in order to ensure the formation of the carbides at temperatures lower than about 500°C. It had been found previously' that in chromium steels the chromium carbide could be produced by transformation of austenite in the pearlite range above 500°C., and, in order to start the tempering treatment with the carbide present as cementite, the absence of pearlite was considered to be essential. The specimens were tempered for 64 hr. to permit the formation of the stable carbide. After tempering, the specimens were submitted to microscopic and X-ray examination. The carbides were separated either electrolytically in a 5 per cent HCl electrolyte or with an acidified copper-potassium chloride solution. The crystal structure of the carbide phases was determined by the Debye-Scherrer type of X-ray patterns. The X-ray patterns obtained from specimens near the boundaries between two types of carbides were generally weak and diffuse as compared with the other patterns. For this reason, and because the more soluble components may be masked, the boundaries between the phases could not be determined precisely. However, the X-ray analysis of electrolytically separated carbides has given consistently reproducible results

Jan 1, 1943