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CAPTAIN JOHN H. BALLAMY John H. Ballamy, Captain on the Regimental Staff of the 103d Engineers, was killed near Fismes, on August 9, 1918. Captain Ballamy was born at Plymouth, Pa., in 1886 and graduated from the High School of that town in 1903. In the following year he began work as a coal miner, and at the same time began the study of mining engineering in the evening and any other spare time. He continued this diligent and energetic practice for many years. For sixteen years he continued in the employment of the Delaware, Lackawanna & Western Railroad Company's coal mining department, first as a practical miner, next as rodman on the survey corps, of which he soon became chief, then draftsman and mining engineer in the company's office. In January, 1916, he was promoted to the position of district engineer in charge of mining engineering work. He became a member of this Institute in September, 1917. CAPTAIN JOHN H BALLAMY. Captain Ballamy's military experience began at the same time as his engineering practice. He was a charter member of Company A, Engineers, N. G. P., which was organized ten years ago. He saw service on the Mexican border as First Lieutenant of this Company, and remained in that position until November, 1916. In July, 1917, he was again called into service and was sent to Camp Hancock, Ga., with the 103d Regiment of Engineers. Before joining this regiment he attended a school for officers at Fort Sill, Okla., where he received special training in field fortifications. In December, 1917, he was commissioned Topographical Captain on the Regimental Staff, and in April, 1918, preceded his regiment by about six weeks to France, where he again attended an officers' training school. He rejoined his regiment on July 15, near Chateau Thierry and was killed near Fismes on August 9, 1918.

Jan 11, 1918

By C.P. Clingerman

The recent iastallation of a combination dust catcher, gas washer and precipitator at Carrie blast furnaces of the Homestead Steel Works has given very satisfactory results. The following description of the gas-cleaning installation is based on actual operating experience. When Carrie blast furnace No. 3 was rebuilt the decision was made to fine-clean all the gas produced by the furnace, because the management wanted fine-cleaned gas for fuel in the stoves and boilers. In order to keep the installation cost as low as possible, the gas-cleaning system was streamlined. All by-pass lines and goggle valves for bypass lines were eliminated from the system. A radical departure from the use of 8-in. tubes in the Cottrell-type precipitators was made by using 12-in. diameter tubes in a single compartment, the single compartment being divided into two units by a single partition wall in the top header. The Furnace Proper A few of the characteristic points and dimensions of the furnace proper are: Hearth diameter.............. 26 ft. o in. Bosh diameter............... 29 ft. 1 3/4 in. Stock-line diameter........... 20 ft. o in. Large-bell diameter........... 14 ft. 8 in. Small-bell diameter........... 6 ft. 6 in. Height of furnace from center of iron notch to bottom of large bell, closed............ 94 ft. 4 in. Height of furnace from center of iron notch to main platform ...................... 105 ft. o in. The furnace top is equipped with a MC-Kee revolving distributor, the operation of which is electrically interlocked with the skips, bell hoists, and coke-weighing and charging apparatus through a Freyn automatic charging control. The gas leaves the furnace first through four offtakes, 5 ft. 9 in. in diameter, then through uptakes that, are approximately 85 ft. high, which feed into a single down-comer that is 9 ft. in diameter. Dust Catcher The primary dust catcher installed (Fig. I) may be described as a single-cone type. The inlet pipe at the top of the dust catchei is 9 ft. in diameter; it is flared to a conical shape as it extends down inside the dust-catcher shell, which is 35 ft. in diameter, 37 ft. high on the straight side, and has a conical top and bottom. The slope of the bottom is so0 with the horizontal. The cone-shaped pipe inside the dust-catcher shell is 42 ft. high and is 19 ft. 7 in. in diameter at its bottom lip where it releases the gas inside the dust catcher. The ratio of the area of the inside cone at its bottom lip to the

Jan 1, 1943

By C. P. Clingerman

The recent iastallation of a combination dust catcher, gas washer and precipitator at Carrie blast furnaces of the Homestead Steel Works has given very satisfactory results. The following description of the gas-cleaning installation is based on actual operating experience. When Carrie blast furnace No. 3 was rebuilt the decision was made to fine-clean all the gas produced by the furnace, because the management wanted fine-cleaned gas for fuel in the stoves and boilers. In order to keep the installation cost as low as possible, the gas-cleaning system was streamlined. All by-pass lines and goggle valves for bypass lines were eliminated from the system. A radical departure from the use of 8-in. tubes in the Cottrell-type precipitators was made by using 12-in. diameter tubes in a single compartment, the single compartment being divided into two units by a single partition wall in the top header. The Furnace Proper A few of the characteristic points and dimensions of the furnace proper are: Hearth diameter.............. 26 ft. o in. Bosh diameter............... 29 ft. 1 3/4 in. Stock-line diameter........... 20 ft. o in. Large-bell diameter........... 14 ft. 8 in. Small-bell diameter........... 6 ft. 6 in. Height of furnace from center of iron notch to bottom of large bell, closed............ 94 ft. 4 in. Height of furnace from center of iron notch to main platform ...................... 105 ft. o in. The furnace top is equipped with a MC-Kee revolving distributor, the operation of which is electrically interlocked with the skips, bell hoists, and coke-weighing and charging apparatus through a Freyn automatic charging control. The gas leaves the furnace first through four offtakes, 5 ft. 9 in. in diameter, then through uptakes that, are approximately 85 ft. high, which feed into a single down-comer that is 9 ft. in diameter. Dust Catcher The primary dust catcher installed (Fig. I) may be described as a single-cone type. The inlet pipe at the top of the dust catchei is 9 ft. in diameter; it is flared to a conical shape as it extends down inside the dust-catcher shell, which is 35 ft. in diameter, 37 ft. high on the straight side, and has a conical top and bottom. The slope of the bottom is so0 with the horizontal. The cone-shaped pipe inside the dust-catcher shell is 42 ft. high and is 19 ft. 7 in. in diameter at its bottom lip where it releases the gas inside the dust catcher. The ratio of the area of the inside cone at its bottom lip to the

Jan 1, 1943

By C. J. Fleisch, C. P. Clingerman

THE recent installation of a combination dust catcher, gas washer and precipitator at Carrie blast furnaces of the Homestead Steel Works has given very satisfactory results. The following description of the gas-cleaning installation is based on actual operating experience. When Carrie blast furnace No. 3 was rebuilt the decision was made to fine-clean all the gas produced by the furnace, because the management wanted fine-cleaned gas for fuel in the stoves and boilers. In order to keep the installation cost as low as possible, the gas-cleaning system was streamlined. All by-pass lines and goggle valves for bypass lines were eliminated from the system. A radical departure from the use of 8-in. tubes in the Cottrell-type precipitators was made by using 12-in. diameter tubes in a single compartment, the single compartment being divided into two units by a single partition wall in the top header. THE FURNACE PROPER A few of the characteristic points and dimensions of the furnace proper are: [Hearth diameter 26 ft. o in. Bosh diameter 29 ft. 1 3/4 in. Stock-line diameter 20 ft. o in. Large-bell diameter 114 ft. 8 in. Small-bell diameter 6 ft. 6 in. Height of furnace from center of iron notch to bottom of large bell, closed 94 ft. 4 in. Height of furnace from center of iron notch to main platform 105 ft. o in.] The furnace top is equipped with a McKee revolving distributor, the operation of which is electrically interlocked with the skips, bell hoists, and coke-weighing and charging apparatus through a Freyn automatic charging control. The gas leaves the furnace first through four offtakes, 5 ft. g in. in diameter, then through uptakes that are approximately 85 ft. high, which feed into a single down-comer that is 9 ft. in diameter. DUST CATCHER The primary dust catcher installed (Fig. I) may be described as a single-cone type. The inlet pipe at the top of the dust catcher is g ft. in diameter; it is flared to a conical shape as it extends down inside the dust-catcher shell, which is 35 ft. in diameter, 37 ft. high on the straight side, and has a conical top and bottom. The slope of the bottom is 50° with the horizontal. The cone-shaped pipe inside the dust-catcher shell is 42 ft. high and is 19 ft. 7 in. in diameter at its bottom lip where it releases the gas in-

Jan 1, 1943

By U. N. Bhrany, N. Arbiter

Over the last one hundred years the problem of assessing the efficiency of energy consumption in producing rock breakage has been studied by many investigators from a variety of viewpoints. Their contributions fall into three main categories: (1) those characterized by a priori assumptions of a relation between the energy required for fracture and the degree of fragmentation produced (e.g. Rittinged(l), Kick(2) Bond(3)); (2) those studying simple systems with controlled energy applications (Gross(4)?(5) ); (3) those treating operating data to develop practical relative measures of efficiency (Coghill(6), Taggart(7)). Closely involved in the efficiency problem is the characterization of size distributions produced by comminution. Unless an experimental size distribution can be approximated by a distribution function and its constants, it becomes difficult if not impossible to develop quantitative energy-particle size relationships. While many such functions have been proposed, none are both completely adequate and convenient. The simpler fonns usually describe only a part of the sizing data, while those which apply to a greater extent are cumbersome(8). The thesis underlying the study here reported can be stated as follows: Size reduction in a tumbling mill subjects ore particles to a relatively constant fracture-producing environment. The intensity of the forces is determined by such factors as the mass of the breaking bodies (rods, balls, rock), and the speed and dimensions of the shell. For a particular ore the extent of size reduction will depend both on the intensity and on the duration of exposure. The latter is the residence time in a batch mill, or is determined by the effective volume and the flow rate in a continuous mill. Size reduction can thus be regarded as a rate process in which the weight of material in size intervals varies as a function of grinding time. Although both the rates of disappearance of coarser sizes and

Jan 1, 1961

By Edward W. Parker

(Toronto Meeting, July, 1907.) NOTE.-The material from which this paper has been prepared was collected for the U. S. Geological Survey Bulletin, Contributions to Economic Geology, 1906, and appears also, though in somewhat more extended form without illustrations, in that publication. ALTHOUGH the briquetting of coals and lignites has been carried on for many years in Europe, and has reached a particularly high state of development in France, Belgium, and Germany, it has made comparatively little progress in the United States. The causes for the backwardness of the United States in this regard are several, and first among them has been the abundant supply of cheap raw fuel with which the manufactured article has to compete. With our millions of acres of coal-productive territory, from which the product can in most cases be cheaply extracted, it has appeared in many districts more economical to waste the slack or culm, which con¬stitutes a considerable percentage -of the product, than to attempt to save it at the additional expense required for briquetting. It is for this reason that the view in all sections of the anthracite-region of Pennsylvania is marred by the unsightly culm-banks which encumber the ground, and that in some of the bituminous-districts one sees huge piles of unmarketable- slack allowed to burn up in order to get rid of them. When the coal is of a coking quality, or when the slack can be used for steaming-purposes, these losses are not sustained, but many thousands of tons of what might be converted into usable fuel have been wasted every year simply because of the increased expense involved in its preparation.

Sep 1, 1907

By K. P. Fournier

This report describes work on the problem of predicting oil recovery from a reservoir into which water is injected at a temperature higher than the reservoir temperature, taking into account effects of vis cosity-ratio reduction, heat loss and thermal expansion. It includes the derivation of the equations involved, the finite difference equations used to solve the partial differential equation which models the system, and the results obtained using the IBM 1620 and 7090-1401 computers. Figures and tables show present results of this study of recovery as a function of reservoir thickness and injection rate. For a possible reservoir hot water flood in which 1,000 BWPD at 250F are injected, an additional 5 per cent recovery of oil in place in a swept 1,000-ft-radius reservoir is predicted after injection of one pore volume of water. INTRODUCTION The problem of predicting oil recovery from the injection of hot water has been discussed by several researchers. l-6,19 In no case has the problem of predicting heat losses been rigorously incorporated into the recovery and displacement calculation problem. Willman et al. describe an approximate method of such treatment. 1 The calculation of heat losses in a reservoir and the corresponding temperature distribution while injecting a hot fluid has been attempted by several authors.7'8 In this report a method is presented to numerically predict the oil displacement by hot water in a radial system, taking into account the heat losses to adjacent strata, changes in viscosity ratio with temperature and the thermal-expansion effect for both oil and water. DERIVATION OF BASIC EQUATIONS We start with the familiar Buckley-Leverettg equation for a radial system:* This is sometimes referred to as the Lagrangian form of the displacement equation. If the injection water is at some temperature Ti which is above reservoir temperature TR, then lw, the flowing fraction which is water, is a function of both water saturation and fluid temperature at any given time and position in the reservoir. The temperature dependence of lw results from the temperature dependence of the viscosity ratio µo/µm of which lw is a function. Thus:

Jan 1, 1966

By V. F. STANLEY

THERE is not the slightest doubt that the present recoveries of valuable minerals by the Broken Hill mills could be improved, and that further machinery would be installed for the purpose if it could be assured that a high price of metals would be maintained; but concentration, like all other com¬mercial propositions, is a question of profit and loss, and no sensible metallurgist would carry out his work in such a manner as to obtain high recoveries at a financial loss to his company. There is a point in all concentration at which the added money obtained from the higher recovery of minerals is balanced by the extra cost of obtaining the last part of the recovery; a wise metallurgist will be content with his work before his working-expenditure reaches this point. To a casual observer, the work being done by the various Broken Hill mills is identically the same. A closer investigation will show that this is far from being the case, for it is a remarkable fact that, although the larger mines extend over a length of approximately only 2 miles, such a difference in the physical and chemical properties of the ores exists that, from the first reduction downwards, special treatment is necessary in each case, and hardly any two mills are identical in operation. The details of this paper refer directly- to the work of one . particular mill in Broken Hill, for the reason that the ore from each mine presents special difficulties and requires its own special treatment. I have selected the mill referred to because it has been constructed and modified under my supervision.

Sep 1, 1909

Discussion of "The Ordering Transformation in Titanium-Aluminum Alloys Containing up to 25 at. pct Aluminum" by F. A. Crossley, Trans, TMS-AIME, 1968, vol. 242, pp. 726-30. Dr. Blackburn, in his response, pointed out that the c/a ratio of the Ti-26 at. pct A1 alloy quenched from 900°C is 1.6006 and not 1.5965 as plotted in Fig. 19. An error in transcription was made in preparing Table V of Ref. 1 (F. A. Crossley, Trans. TMS-AIME, 1966, vol. 236, p. 1174). The corrected Fig. 19 strengthens rather than weakens the argument that a martensitic transformation occurs in alloys containing from 15 to 18 at. pct Al. A Study of the Factors Which Influence the Rate Minimum Phenomenon During Magnetite Reduction, by P. K. Strangway and H. U. Ross, Trans. TMS-AIME, 1968, vol. 242, pp. 1981-88. Page 1883 Reverse Figs. 5 and 6. Sigma—Its Occurrence, Effect, and Control in Nickel-Base Superalloys, by J. R. Mihalisin, C. G. Bieber, and R. T. Grant, Trans. TMS-AIME, 1968, vol. 242, pp. 2399-2414. Pages 2403, 2407, and 2414 Arrows should appear on Figs. 4 and 9; a sigma is missing from Table IV; line 2, page 2414, change "case" to "cast". These are published correctly in the bound volume. Index, December Issue There were a few omissions in the Author Index. These have been corrected and published properly in the bound volume.

Jan 1, 1970

By J. Gurland

A method of measuring the degree of contact between grains of 1 phase in multiphase alloys is derived. It is shown that the number and the areas of contacts between grains can be determined by metallographic analysis. Relations between the area of contact, the dihedral angle, and the mean free path of spherical grains are developed. It is often desirable to interpret the mechanical behavior of alloys in terms of structural features observed under the microscope. A quantitative description of microstructures of multi-phase alloys requires consideration of geometrical and topological factors such as the amount, size, and shape of grains and the continuity of the constituent phases. The number and dimensions of grains may be metallographically determined by applying the methods described by Smith and Guttman1 and by Fullman2 in their papers on the measurement of particle size and internal boundaries in opaque bodies. The present work intends to show that the same methods can be used to evaluate some of the parameters of phase continuity; namely, contiguity, number of contacts, and areas of contact of grains of 1 phase. Application of these concepts may be found in such problems as the calculation of electrical conductivity of 2-phase alloys, or the evaluation of the length of fracture path through grains of a brittle phase more or less dispersed in a matrix. I. DEFINITION The following definitions and derivations are based on 2-phase alloys, of which the units of the microstructure are grains. Contiguity is defined as the average fraction of surface area shared by 1 grain of a phase with all neighboring grains of the same phase. In a 2-phase alloy of constituents A and B, the contiguity of phase B is given by: where Cb is the contiguity of phase B aBß is the average area of contact between a grain of phase B and adjacent grains of phase B, aAB is the average area of contact between a grain of phase B and adjacent grains of phase A. The contiguity is a measure of the average degree of contact between grains of phase B in the 2-phase mixture. The ratio varies from 0 to 1 as the distribution of phase B changes from a completely dispersed to a fully agglomerated structure. Smith and Guttman1 have shown that the relative areas of 2 kinds of interfaces in a volume are directly proportional to the number of interfaces of each type intersected by a random line. The lines can be drawn on metallographically prepared surfaces as long as these are random sections through the sample. The contiguity ratio of phase B then becomes: Cb = 2NBB/NAB + 2N BB [2] where: NBB is the number of interfaces between grains of phase B intersected per unit length of random line. Nab is the number of interfaces between grains of phase B and grains of phase A intersected per unit length of random line. The factor 2 arises in Eq. [2] because each interface B-B is associated with 2 grains of B-phase, whereas the interface A-B involves only 1 grain of B-phase. The calculations of contiguity do not require any assumption as to volume fraction, particle size, or particle shape. Fullman2 gives an expression for the determination of the mean free path between particles. The mean free path PB between grains of phase A, i.e., through grains or agglomerates of phase B, may be found from the relationship: PB =2fB/NAB [3] where fB is the volume fraction of phase B. The calculation of the mean free path is independent of grain shape and grain size distribution. The calculation of the number of contacts and of contact area requires assumptions of grain shape, grain size, and contact shape. In the case of uni-

Jan 1, 1959

By R. M. Kuralt

CONCESSION from the Chilean government granting the company use of the Rio Salado water stipulates that a minimum of 35,000 metric tons of such salty water must be diverted from the Salado daily, and be disposed of in a manner not to contaminate the Rio Loa to the east of its confluence with the Rio San Salvador at Chacance, Fig. 1. Consequently, the selection of a suitable site for the sulphide plant tailing disposal had the additional problem of the disposal of excess brackish waters. Studies of the terrain in the area directly to the south of the plant indicated that construction of tailing ponds would be costly, and generally unsatisfactory due to excessive slope of the ground, lack of flanking hills, danger of raising the groundwater level thus causing springs along the Antofagasta-Bolivia railroad; and the possibility of creating a dust nuisance at the plant site or at Calama due to blowing sands. Further to the southwest the San Salvador river canyon offered some possibilities and surveys were made at sites A to D, inclusive, Fig. 1. Here again costly dams would have been necessary to impound the tremendous quantity of tailing produced by the new plant, and provision would have to be made to handle a considerable flow of water through the disposal area to the west of the Rio Opache. Additional surveys were made in the area to the southeast of the plant which brought to light the existence of a natural basin of about 14 1/2 square miles, flanked by low-lying hills, and requiring only the most rudimentary dams to close the outlets at the western end. It has been estimated that the Salar de Talabre, as this basin is called, has the ultimate storage capacity of 1 billion tons of tailing with dam heights of only 60 ft. Pipe Line Layout After the decision had been reached to use the Salar de Talabre for the main disposal area various methods were considered for transporting the tailing from the 300 ft diam thickeners to the disposal site, 17.7 miles away. The first plan was to construct a. concrete launder on a 11 pct grade but a survey of this grade line indicated it would not lead directly into the Salar but would terminate somewhat to the north, and from that point on the slope would be considerably less than 1 pct along a natural watercourse. A slope of 2 pct or even 2 ¼ pct in such a rough watercourse or ditch is the ideal, judging from the success of operations at Cananea. Since a simple ditch is the most economical vehicle for the transportation of tailing, it was decided to use one with a 2 ¼ pct grade for as much of the distance as was possible with the existing terrain and to construct a concrete headbox, or forebay, at the terminus of the ditch to supply a steel pipe line leading the rest of the way into the Salar. Referring to the maps of the Chuquicamata area outlining the tailing disposal system, Figs. 1 and 2, one can trace the flow of the tailing from the thickeners to the Salar as follows: The 8-in. spigot pipes from the thickeners (1) discharge into a concrete sump feeding a concrete launder with a 2 pct slope. The launder carries the tailing into a concrete junction box (2) located south of the sulphide plant. The purpose of the junction box is to provide emergency facilities for diverting the tailing flow from the main disposal area to the emergency area. It also provides a junction point for the launder from the sulphide plant's future east extension. The box is fitted with a pair of radial gates controlling flow into the main tailing ditch (6) or the emergency ditch (3) which terminates on the pampa. All ditches are of the same cross section, namely 4 ft wide at the bottom with 1 : 1 side slopes and with a minimum depth of 21/2 ft. Where the nature of the ground demands it, the ditches are lined with stone rubble. The overflow from the plant's 10 million gal reservoir (4) is led into the main tailing ditch by way of a spillway ditch (5). The main ditch terminates in a trashscreen fitted, concrete forebay (7) which serves as a headbox for a 24 in. O.D. x 3/8 in. wall, Dresser coupled, steel pipe (10) and a 12 in. x 8-gage spiral welded, wedge-lock coupled, steel pipe (8). The forebay is fitted with two compartments, one for each pipe, and flow into each is controlled by radial gates. Each compartment is fitted with a 6-in. drain pipe controlled by a rubber pinch valve. To protect the forebay from the abrasive effects of the tailing, the sides of the compartments will be lined with old

Jan 1, 1952

By L. F. Reinartz

ANOTHER year has rolled by. We are twelve months further away from the start of the depression and. therefore that much nearer to recovery. The accumulated needs and wants 'of our lame, virile population are ready to burst their bonds as soon as confidence returns and credit is stabilized. The steel industry has been harassed by uncertainty of the Government's attitude toward organized labor's de¬mands, and irregularity in volume of production. Steel labor, if left to its own devices, would in most instances, march shoulder to shoulder with the management of steel plants in their effort to pull the industry nut of the ?red.? Cooperation between management and men has improved considerably during; the year.

Jan 1, 1935

By R. S. Olson

An analysis of the wash thickener-two stage filtration circuit was undertaken to quantitatively describe the circuit performance. On the basis of information developed in this study, the factors of major importance in reducing soluble loss are: filter wash, wash zone size, filter capacity, filter feed pulp dilution, blow back, filter cloth, water to repulp, water to primary wash, water to feed pulp and thickener underflow. The wash thickener-two stage filtration circuit has been used in many hydrometallurgical plants to separate solute from leach solids. As many as 20 to 30 plant operating variables determine the performance of the wash circuit and hence the soluble loss. The complex relationships of these variables, with each other and with soluble loss, make the evaluation of their effect on mill performance difficult. An analysis of this type of circuit was undertaken to quantitatively describe the circuit performance. THE CIRCUIT The diagram in Fig. 1 represents the operation of the wash thickener-two stage filter circuit. This circuit provides for counter current, cross current or for a combination of countercurrent-cross current operation. The feed pulp stream, J, is intended to include any solution or water, other than primary and secondary filtrate, which enters the wash thickener. Except for filter washing, the operations in the circuit are simple mixing operations which can be described by material balance. The concentration of solute in the filtrate and cake solution after application of wash liquor cannot be described by simple material balance. The displacement washing equation, derived by Rhodes1 and extended by Choundhury and Dahlstrom,2 appears to best describe this case of miscible displacement in porous media. In a useful form, it is

Jan 1, 1963

By F. W. Bowdish

Analysis of data on the equilibrium exchange of sodium ions from salt solutions for calcium ions, from various fractions of an oxidized calcareous iron ore from Lorraine, France, strongly suggests that the calcium may be attached to the solid by only one of its bonds. Nearly equal values were found for the equilibrium expression from data covering concentration ranges of 1000 fold for the salt and 235 fold for the ion-exchangeable calcium. Differences indicated between the exchangeable calcium contents of the slime and the sandy portions of the ore, as well as the equilibrium relationship itself, are of importance in the flotation of the ore. During an experimental program on the flotation of oxidized calcareous iron ores from Lorraine, France, ion-exchangeable calcium was found in the ore. Because this source of calcium ions, if not controlled, could thwart all efforts at flotation, an extensive study was made of the phenomenon. The part of that study reported here was to obtain data on the equilibrium established between ore solids and solutions of sodium chloride, and to analyze the data to try to understand better the physical chemistry of the exchange. The ore contains oolites of goethite and quartz grains embedded in a matrix consisting of calcite and a brown cement believed to be the residue from oxidation of the silicate minerals found in fresh unoxidized ore. When ground, a substantial fraction of the ore becomes a fine slime in which is concentrated more than half of the alumina content of the ore. It is believed that the alumina occurs as unidentified clay minerals with which most of the ion-exchange capacity of the ore is associated. A preliminary experiment on the exchange of ions from the ore for sodium ions from solutions of NaCl showed that calcium was about ten times as abundant as magnesium in the exchange solutions. Thus calcium was the predominant ion in causing difficulties with flotation, and it was the only one for which data was obtained in this study. It is believed that the omission of magnesium ion exchange from consideration does not invalidate the results. It is possible, however, that inclusion of magnesium in the analyses and calculations or its prior elimination from the samples would have improved the consistency of the results. EXPERIMENTAL Data about the equilibrium exchange of sodium from solutions for calcium from the solids were obtained using two different samples of material. The solids

Jan 1, 1963

By Charles R. Fettke

In connection with the gathering of data for a report on the geology of the Humacao district of southeastern Porto Rico during the summer of 1916, under the auspices of the Scientific Survey of Porto Rico and the Virgin Islands, conducted by the New York Academy of Sciences, the writer had an opportunity to make a brief examination of the magnetite deposits of the eastern portion of the island. Inasmuch as these deposits have attracted some attention as possible future sources of iron ore, and as there is only scanty reference made to them in literature,' a short description based on this reconnaissance survey was deemed to be worthy of record. The magnetite deposits of eastern Porto Rico occur in association with a belt of limestone, trending approximately S 53° E, which begins 11/2 miles east of Juncos and can be traced for about 6 miles toward the coast. Four outcrops of magnetite were observed, which are designated as Nos. 1, 2, 3, and 4, respectively, on the map shown in Fig. 1. Others are probably present, but escaped notice on account of the thick mantle

Jan 1, 1924

By Grafton Thrasher

For the proper discussion of this subject it is necessary to incorporate in this paper the substance of part of an article published by rite in Metallurgical and Chemical Engineering for January, 1915. CAST-IRON TEST PLUG, SHOWING LOW MOTTLED AND MOTTLED FRACTURES. Several years ago, while I was thief chemist for a malleable-iron company, it occurred to me that the mere regulation of silicon was not sufficient to produce the desired degree of chill. But only after repeatedly encountering this fact in befit open-hearth and air-furnace, as well as cupola metal, was I led to regard variation in the total carbon as the principal cause of disturbance; and I resolved to determine the total carbon as well as the sill on ill every heat, and by a carefully prepared test piece to gauge the approximate degree of chill. The test piece selected was a round plug, 2 in. in diameter and 12 in. long, cast on end

Jan 10, 1915

By B. M. Larsen

Jan 1, 1948

By B. M. Larsen

Jan 1, 1948

By William B. Plank

IN this paper are presented the results of a complete statistical survey of the enrolment, courses and degrees, and the employment situation of recent graduates in all of the 46 institutions in the United States, including Alaska, and four of the six schools in Canada that grant degrees in mineral technology. The data were collected by the writer by means of a questionnaire in December, 1933, to which all the institutions, except one, replied. Included also are data of a similar study made by the writer in May, 1933, for the S. P. E. E. summer session of mining and metallurgical teachers at the University of Wisconsin and published in the Journal of Engineering Education for December. 1933. Some of the data of a paper of the writer's entitled, "The Trend of Mineral Engineering Student Enrolment in the United States," read at the A. I. M. E. meeting on Feb. 16. 1931, are given for comparison. Acknowledgment is made to all the administrative officers of the institutions who took the time from their busy lives to supply the in- formation asked for. Their fine cooperation has alone made possible these two surveys with a completeness of detail never before attempted

Jan 1, 1934

By Fred D. Wright, Philip B. Bucky

FOREWORD By E. D. Gardner[$] THE present known petroleum reserves are limited, and unless important new fields are discovered the Nation will be dependent, in the not too distant future, upon imports or upon synthetic liquid fuels to supplement domestic petroleum supplies. As a national defense measure, it would appear desirable to develop a substitute source of liquid fuels within the borders of the country if new discoveries of petroleum should prove disappointing. Congress undoubtedly had this in mind when it passed the Synthetic Liquid Fuels Act (78th Congress, approved April 5, 1944). The act directed the Bureau of Mines to construct and operate demonstration plants for producing synthetic liquid fuels from coal and oil shale, and from forestry and agricultural products. As a part of this program, the Bureau of Mines has built a demonstration plant at Anvil Points, near Rifle, Colo., for retorting the oil shale, and has developed a mine to supply the plant with oil shale. At the beginning of mining operations, the physical qualities of the oil shale, which would have a bearing upon the selection of mining methods and practices for exploiting the deposits, were largely unknown. A comprehensive mining-research program has been set up to determine the best methods and practices to follow to obtain the lowest overall practicable mining cost for producing the oil shale on a large scale. Unusually low mining costs will be necessary for shale oil to be produced commercially. Consideration now is being given to exploiting a series of flat-lying beds 70 ft thick by a room-and-pillar method of three benches with the advance heading at the top. This report is one of a series giving the results obtained in research on mining problems. INTRODUCTION The structure and physical qualities of the oil shale indicate that the 70-ft interval of the oil-shale measure can be safely mined in open stopes; studies indicate that this method would be the most economical one to use.1 As labor will be a major item of expense, the largest equipment that could be operated efficiently in the workings would be required to obtain the lowest practicable mining cost. It was therefore desirable to know the maximum safe span of unsupported rooms and the size of pillars necessary to support the overburden. This paper describes the work done at the Columbia School of Mines Barodynamic Laboratory in determining the maximum safe unsupported roof spans and the minimum safe pillar dimensions for the oil-shale mine at Rifle, Colo.

Jan 1, 1947