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By R. W. Brauchli

The author has made a statistical analysis of the relations between recovery and length of cores, diameters of cores and of drill pipe, and cost of coring under different contractual conditions. The data are presented in tabular form. The length of the core appears to play a minor part in the percentage recovered, with a tendency to lower percentages with increased core length; approximately 50 per cent of the length cored is normally recovered if the cores are more than 2 in. in diameter. The diarneter of the core has a more pronounced influence on the recovery; 1 3/8-in. cores average 29.1 per cent, 2 1/8 -in. core 53.4 per cent and 3 1/8-in. cores with 6 5/8-in. drill pipe 68.5 per cent. Where 31/8-in cores were taken with smaller diameter drill pipe, i. e., 4 1/2 - in. drill pipe in 8 1/4-in. hole, the recovery dropped to 49.8 per cent, indicating that the more rigid drill pipe favors good core recovery, a conclusion also borne out by the results obtained in using large drill pipe for smaller diame-ter cores. Surprisingly good recoveries were obtained from both the Johnson and Simpson sands; shale, limestone and dolomite each gave less then average recoveries. No serious fishing jobs resulted from coring operations. Attention is called to the need of a core drill which will make it possible to drill a full-sized hole while coring, and thus reduce the cost of cores. Data pertaining to core drilling should be consistently recorded, such as, diameter of drill pipe and collar, diameter of core, recovery, size of cutter heads, type of core-barrel, coring time, speeds of rotation and of pumps, weight on core-barrel, etc.

Jan 1, 1933

By Herbert Everest

THE idea of economical production is usually associated with large operations, tonnages, and mines, with even larger capital behind them. Nevertheless many small mines operate in the shadow of large competitors and make a good showing on the capital invested despite larger overhead expenses. For the purpose of this discussion I will divide the cost of production into classes, and specify opposite each the approximate percentage expended thereon: 1. Labor, including miners and company men; 60 to 75 per cent. 2: Development, including all necessary yardage, room turning, crosscuts, etc.; 7 to 12 per cent. 3. Deadwork, covering payment to miners for handling falls, drawslate, or faults, rock work, water, and in general, all nonproductive labor; 3 to 7 per cent. 4. Supplies: mine timbers, oil, brattice material, lumber, cement, and repairs to equipment; 2 to 6 per cent. 5. Expense: management, selling, office, taxes, etc.; 1 to 5 per cent. 6. Depreciation of coal reserves and royalties; 6 to 10 per cent. 7. Depreciation of equipment and interest on capital invested; 1 to 3 per cent. 8. Fuel; ½ to 1 ½ per cent. The labor cost for a small mine is relatively much lower than for a large mine, particularly the sum paid to company men. Idle-pay expense is cut to a minimum; the labor item being larger than all the others together, a saving in labor makes a decided showing on the total cost. The development charges in small and large mines are about the same. There should be a small showing on the Beadwork item in favor of the small mine. The unit cost for supplies is usually small in the small mine. The operator of the small mine finds, owing to a limited tonnage, that the expenses of management, selling, etc., are abnormally high.

Jan 1, 1916

[This paper was presented for the author, who is now in the National service, by H. M. Chance, and was illustrated by a working model built to the plans shown in the Bulletin, for February, 1918, p. 268. For more convenient operation, a small motor-driven centrifugal pump was attached to the overflow outlet (5, Fig. 1, p. 267) delivering to the inlet, 4. The pressure of ingoing water was then adjusted by a valve at 4, and was indicated by the column of water in a vertical glass tube connecting with the pressure chamber; an average pressure of 15 in. of water column was maintained during the demonstration. The sand used in the test was beach sand, which had simply been passed through a 20-mesh screen; about 45 per cent. was finer than 80 mesh. The circulation of water was at the rate of about 5 gal. per square foot per minute, and the rising velocity of the water in the box as a whole (not the velocity of the jets) was 6 in. per minute. At this consistency, the specific gravity of the fluid mass was between 1.55 and 1.60. Mr. Chance estimated that it would be possible to maintain a fluid mass of this consistency, 3 ft. deep, and with an area of 50 by 60 ft. with an expenditure of 15-hp. Mr. Chance then showed how a lump of coal would float on the fluid mass, and also made a separation of slate from coal on a small scale in a wire cage submerged momentarily in the fluid. The following demonstration relates to a matter not contained in the original paper, and was presented as of interest to the operators of cyanide and other lixiviation processes dealing with fluid mixtures of varying specific gravity.-Ed.]

Jan 4, 1918

By George A. Packard

The experiments here described were made under my supervision while temporarily acting as head of the Department of Metallurgy at the Missouri School of Mines, at Rolla. The work was done by Messrs. W. E. Brown, W. C. Richards, and F. L. L. Wilson, and the description of the results forms a portion of a thesis presented for the degree of Bachelor of Science. These results are submitted because I have seen no description of the application of pot-roasting to the treatment of a copper-concentrate. This work followed that already described in the discussion of the Lime-Roasting of a Galena-Concentrate,' and was similarly occasioned by the lack of a furnace of the reverberatory type in which the copper-concentrates could be smelted. The material for the tests was a mixture of three lots available. The first consisted of 8-mm. ,jig-concentrates and the second of 0.5-mm. table-concentrates, from the San Juan district, Colo. Both contained pyrite, chalcopyrite, and galena. The third was a lot of copper-ore, found in the laboratory, which had been crushed through 8-mm. screen, and had at some previous time, before the laboratory roasting-furnace was torn down, received a partial roasting. The analyses of these ores are given in Table I. In order to determine the applicability of pot-roasting as to each lot of ore, and the effect of varying proportions of lime, six preliminary tests were made in a " size N " Battersea crucible, 9 in. high by 6 in. in diameter at top, with a 5/8-in. hole bored through the bottom. The crucible was set in the upturned elbow of a 2-in. pipe, the joint being luted with clay; and air was obtained from a receiver supplied by a compressor. The

Jan 1, 1908

By Herbert A. Everest

The idea of economical production is usually associated with large operations, tonnages, and mines, with even larger capital behind them. , Nevertheless many small mines operate in the shadow of large competitors and make a good showing on the capital invested despite larger overhead expenses. For the purpose of this discussion I will divide the cost of production into classes, and specify opposite each the approximate percentagc expended thereon: 1. Labor, including miners and company men; 60 to 75 per cent. 2. Development, including all necessary .yardage, room turning, crosscuts, etc.; 7 to 12 per cent. 3. Deadwork, covering payment to miners for handling falls, draw-slate, or faults, rock work, water, and in general, all nonproductive labor; 3 to 7 per cent. 4. Supplies: mine timbers, oil, brattice material, lumber, cement, and repairs to equipment; 2 to 6 per cent. 5. Expense: management, selling, office, taxes, etc.; 1 to 5 per cent. 6. Depreciation of coal reserves and royalties; 6 to 10 per cent. 7. Depreciation of equipment and interest on capital invested; 1 to 3 per cent. 8. Fuel; ½ to 1½ per cent. The labor cost for a small mine is relatively much lower than for a large mine, particularly the sum paid to company men. Idle-pay expense is cut to a minimum; the labor item being larger than all the others together, a saving in labor makes a decided showing on the total cost. The development charges in small and large mines are about the same. There should be a small showing on the deadwork item in favor of the small mine. The unit cost for supplies is usually small in the small mine. The operator of the small mine finds, owing to a limited tonnage, that the expenses of management, selling, etc., are abnormally high. The item, depreciation of coal reserves and royalties, is usually about

Jan 1, 1916

By Fred Beasley

METALLURGISTS have, in the past ten years, overcome many difficulties of high zinc in lead blast-furnace slags. This problem was brought to the front at the close of the late war, by the price obtained for lead and the high cost of labor, fuel, and flux. The Port Pirie plant of the Broken Hill Associated Smelters, Australia, early in 1919 had increased the zinc oxide in its blast-furnace slags from 13.5 to 20 per cent., at the same time reducing the iron oxide in the slag from 33.5 to 25.6 per cent. These results were obtained by the following proceduret the addition of granulated slag to the roaster charge, 10 to 12 per cent., improved conditions of the crushing and mixing and roasting of the constituents of the sinter. By experiments and improve-ments the practice at this plant has developed a normal slag that carries 22 per cent. zinc oxide, and at times runs as high as 33 per cent. zinc oxide, with only slight furnace trouble. The granulated slag is used not, only as a diluent to the sulfur in the roaster charge, but also on account of its value as a slag-forming material and on account of the physical properties it imparts to the finished sinter. The ore is mixed with flux and given a double roast. All the first roasting is by sintering with the Dwight & Lloyd machines. The final roasting is performed partly by Dwight & Lloyd machines and partly by H. & H. pots. The charge to the blast furnaces consists chiefly of the final sinter, slag shells, siliceous and non-bearing sulfur, lead ores. High blast and ore columns are the practice. The greatest care is taken to obtain in the final sinter as low a sulfur content as possible.

Jan 2, 1925

By Tribhawan N. Srivastava

The Henderson Mine started production in 1976 utilizing a unique, push-pull, mine ventilation plan. As the mine expands in size, the mine ventilation requirements will increase and a more comprehensive system for maintaining ventilation will be needed. A computer simulation study was completed and has resulted in a revised, systematic, mine ventilation plan which is being implemented as mine development proceeds. Based on the computer simulation study, the new ventilation plan takes into account mining of the orebody simultaneously from both the 8100 and 7700 production levels. Present experience and empirical data indicate that volumes of air needed to adequately ventilate both production levels cannot be obtained utilizing the existing mine layout. The total volume of air needed for the mine excluding haulage tunnel will increase from 708 m3/s (1 500 000 cfm) to 1652 m3/s (3 500 000 cfm), thus a new No. 5 exhaust shaft and connecting drifts are being developed to expand the mine ventilation system to service both production levels. The original ventilation plan for the 7500 haulage level was to bring air from the No. 2 and No. 3 intake shafts and exhaust it through the 94X raise to the 8035 exhaust level and then through the No. 1 exhaust shaft to surface. Control of air volumes in loading drifts was initially planned to be achieved by air doors and air curtains. Air doors on the 7500 level require large cutouts, expensive installations and constant maintenance. The computer simulation study indicates that air- flow in the 7500-level production drifts can be con- trolled exclusively by air curtains which are less expensive to install and more effective in regulating airflows than air doors. Air from the 7500 level can be exhausted through a new raise to the new 7635 exhaust level and then through the new No. 5 exhaust shaft.

Jan 1, 1982

By C. W. Volek, J. E. Chappelear

The injection of a hot liquid into initially cool porous media, saturated with the same liquid and surrounded by two impermeable but heat conducting media (cap and base rock), has been studied both experimentally and theoretically. The temperature dependence of the viscosity was included in the theoretical model, but it was assumed that the specilic heats and densities of the various materials were independent of the temperature. Solutions to the theoretical model were approximated by numerical methods. Both theoretical and experimental results indicate that center-line temperatures are significantly higher than boundary temperatures. Comparison of experimental and theoretical results with a cold/hot viscosity ratio 01 19:1 were in reasonable agreement. Theoretical calculations show that the effect of the temperature dependence of viscosity was very significant at ratios of 100:l to 1000:1, which are typical of those that occur when injecting hot water to flood heavy oil reservoirs. INTRODUCTION We consider the problem of prediction of fluid flow and temperature distribution in an initially cold-fluid-filled reservoir on the injection of the same hot liquid by the use of mathematical and physical models. The results reported are for a two-dimensional rectangular section of the reservoir, as shown in Fig. 1. The injection and withdrawal faces are assumed to be equipotentials for fluid flow. The ultimate purpose of such models would be to predict hot-water injection performance. However, we note that in the work presented here, one of the most significant aspects of the problem — the instabilities resulting from two-phase, water and oil flow — is not included. We will not give a historical review, but refer instead to the paper of Spillette and Nielsen,1 which contains a rather complete bibliography and critical discussion. The physical problem is exactly the same as Spillette and Nielsen, except for certain simplifications in our assumptions. The mathematical details are somewhat different, and we will present the details of our method here. The mass flow equation, which is elliptic in character, is handled by successive overrelaxation. The heat flow equation, which is parabolic in character, is handled by a straightforward explicit approximation. Some difficulty arose in the over-all heat balance due to small errors in the solution of the mass flow equation, and we feel that a different formulation of the heat flow equation would be desirable for future work. In addition to the mathematical solutions for the temperature distributions in a porous medium due to the injection of hot liquids, experimental data are presented to check the validity of these solutions. A schematic diagram of the model is shown in Fig. 2. Certain qualitative physical conclusions were obtained from our numerical and experimental results. These are: 1. Assuming high (infinite) conductivity normal to the bedding plane in the reservoir is a poor approximation, and may lead to overestimates of the total heat losses (to cap and base rock) of as much as 50 percent. 2. More heat is retained in the reservoir (per unit of heat injected) for higher viscosity changes. 3. Any particular temperature isotherm moves more rapidly along the center line for higher viscosity changes. Consequently, the approximation of temperature independent viscosity is not suitable for obtaining quantitatively correct results. MATHEMATICAL MODEL Our mathematical model is that of a reservoir of thickness 2h. The problem is idealized from that of a linear hot-water drive, and we imagine that the input face is sufficiently far away from the injection wells that the stream lines enter it normally. Similarly the production wells are far enough from the outflow face that the flow lines leave it normally. The reservoir and surrounding cap and base rock

Jan 1, 1970

By C. C. Harris

This article presents a general equation describing the xYt surface in batch grinding. This equation contains as special cases most of the relationships which have been deduced from kinetic studies. In an earlier paper an analytical solution of the integro-differential equation of batch grinding was proposed1 as a first approximation of the xYt surface.' A more general equation will be presented here, which appears to combine in a single statement many conclusions reached by experiment. A possible underlying mechanism is that the changing product fineness leads to changes in the particle crushing strength and in the probability of encounter with the crushing media.

Jan 1, 1969

By AIME AIME

HOLDING the annual dinner-dance of the Institute at the Waldorf-Astoria had become such a tradition that there was widespread regret when it became known that the demolition of the building to make way for a huge office structure would perforce bring to an end the long and enjoyable series of gatherings our Institute had held there. This year the . Commodore was the place at which it was held and the commuters on the New York Central and the New Haven railroad systems, at least, found it a more convenient place for them. Except for the famous "Hoover" dinner, which will probably set a mark for all time, it was the largest dinner the Institute held, 1000 being present, and it taxed the facilities of the hotel to the utmost.

Jan 1, 1930

By Leo Behr

RECENT years have seen important practical advances in the construction of recording instruments for use with thermocouples. The difficulties of the problem will be appreciated when it is remembered that a 10° F. change in temperature of a base-metal couple at 1500° means a change in the electromotive force of about 0.00035 volt. In a circuit having a resistance of 350 ohms, this means a change in the current of 0.00000,1 amp.; and despite this requirement of great sensitivity the instruments must be sufficiently robust to withstand the rough usage which is so often their lot. Two methods are in common use for the measurement of the electromotive force produced by a thermocouple: 1. The milliammeter method, in which the current produced by the thermocouple in a circuit of known resistance is measured and from. this the electromotive force is deduced by Ohm's law. 2. The potentiometer method in which a known potential difference is opposed to the thermocouple electromotive force. The use of thermocouple and milliammeter for measuring temperature introduces some possible sources of error, for the reading depends on a number of factors, including the following: (1) Electromotive force set up at the hot junction, (2) electromotive force set up at the cold junction, (3) resistance of the circuit including thermocouple, lead wires, and meter, (4) field strength of magnets, (5) strength of controlling springs in meter, (6) friction at pivots, of pointer on paper, etc. It is obviously necessary to so design and maintain the installation that the reading is independent of all the factors except the first. The electromotive force set up at the cold junction can easily be taken care of provided it is constant. This has been approximately accomplished by burying the cold junction in the ground or actually secured by locating it in a container, the temperature of which is kept constant by means of a thermostat.

Jan 9, 1919

By R. W. Smith

IN the manufacture of portland cement, the basic and fundamental essential is a limestone uniformly rich in calcium carbonate and carrying less than 3 per cent magnesium carbonate. In searching for such a deposit, also having in mind quantity of rock, accessibility and quarrying facilities, the Pacific Coast Cement Co. discovered and developed its present quarry at View Cove on the east coast of Dall Island, Alaska. Dall Island is about 100 miles west of Ketchikan and 650 miles north from Seattle. It is 70 miles in length and 5 to 10 miles in width. The climatic conditions are much more favorable than would be expected. The temperature ranges from zero to 70 degrees. The precipitation is considerably more than at Seattle, although there was less snow during the last season than at Seattle. The island is rugged and mountainous; the vegetation is moss and grass and a large portion is timbered with spruce, hemlock and cedar. Some of this timber has market value. The limestone quarry is about 30 miles north of the extreme southern end of the island. The company acquired 240 acres which contain limestone of remarkable uniformity, the percentage of calcium carbonate being about 95. It is estimated that there is ample recover- able rock on the holdings to supply the cement plant at its maximum capacity for well over 100 years.

Jan 1, 1929

By J. L. Huitt, D. K. Lowe

This laboratory flow study covers propping agent transporl in horizontal fractures as influenced by the characteristics of the propping particles, fluid and fracture. Correlations are presented for the transport of particles, deposition and fingering through dunes. For particles moving individually, the article velocity in some cases may be low compared to the hulk average fluid velocity; in other cases, the particle velocity may exceed the bulk average fluid velocity. At high propping agent concentration or low fluid velocity, the likelihood of particle deposition and dune formation increases. Fingering and channeling accompany dune formation. Dune formation can be limited to the outer region of tile fracture by adjustment of pump rate, fluid viscosity and particle concentration. A method is presented by which pump rate and fluid vircosity may he selected to control dune formation for given propping agent concentrations. INTRODUCTION The area and flow capacity of a fracture predominantly control the stimulation derived from hydraulic fracturing. Generation of fracture area exposes the rock matrix to a flow channel; the flow capacity of the fracture controls the flow through the channel to the wellbore under a given pressure gradient. This paper pertains to an important factor that affects the flow capacity of a fracture: the transport of the propping particles in the fracturing during the hydraulic fracturing treatment. With a better understanding of particle transport, fracturing treatments may be designed to obtain more effective distribution of the propping agent. RELATED STUDIES In recent years, the approach to obtaining high fracture flow capacity has been partial monolayer propping. Three papers1-3 have dealt with various facets of this concept. However, in some papers dealing specifically with particle transport, the results generally indicate that packing the propping particles within a fracture is inevitable. In early work&apos; with a sand slurry flowing in a pie-shaped flow cell, qualitative results showed that particle deposition in a fracture during transport should be expected. In other studies"" correlations based on slurry flow in vertical linear flow cells showed particle deposition was primarily a function of fluid velocity. In a recent paper the rate of sand advance (movement of dunes) in a horizontal linear flow cell was discussed. Subsequently, a method was given for designing fracturing treatments, based on the extent of sand advance during a treatment. In this work, particles whose diameter was small relative to the fracture width (i.e., dl W< 0.2) were primarily considered. The particle transport approaches that of sand transport in a stream bed. The industry trend, however, has been to use larger size propping agents. That is, particles with diameters approaching fracture widths predicted by Perkins and Kern.&apos;&apos; This trend has led to generally better fracturing treatments with particle relative sizes in the range 0.2 <d/W < 1.0. In all of the particle transport studies previously cited, the fluid characteristics were discussed. However, little or no attention was given to the parzicle characteristics. In this study the particles. as well as the fluid and fracture, are considered. VARIABLES AFFECTING PARTICLE TRANSPORT In the exploratory work to set LIP a program for this study, it was noted that particles within a batch moved at different velocities in a horizontal linear flow cell. In some cases the particles moved independently. In others there was particle interference. Also, particle deposition and fluid fingering occurred in some cases (Fig. 1). As a result of observing the easily distinguishab&apos;le types of particle movement, the particle variables as well as the fluid variables were included. In addition, the fracture width was included because the fluid velocity profile surely must affect the movement of the larger particles differently than that of smaller particles. Particle variables were density, diameter (based on a sphere of equal volumetric displacement), long and short dimensions, drag coefficient and velocity. Fluid variables were density, viscosity, yield shear strength and velocity. Fracture variables were the width and the angle between the plane of flow and a horizontal plane. Two factors which may prove to be important were omitted because of the complexity resulting from the parameters cited above. These are the fracture-surface roughness and gross irregularities that cause changes in the direction of flow. EXPERIMENTAL EQUIPMENT AND PROCEDURE A linear parallel-plate flow cell (Fig. 2) was used in the study. Lucite plates 1 in. thick formed the test section and permitted a viewing section 11 ft long and 10 in. wide. Spacing between the plates was adjusted with shims to give fracture widths 1/36 to 1/2 in. The model could be tilted to an angle of 45" with the horizonltal plane. Fluid flow was maintained with a Moyno positive displacement pump. Single particles were placed in the flow stream with a particle injector. A .4 - in. diameter plunger moving at a constant rate in a

Jan 1, 1967

By J. P. Kerr, W. N. Logan

Lack of reliable and detailed information on many of the older fields in Indiana has necessitated the use of x and y in many instances. Even in the younger fields many data were lacking. It was thought that even though much was to be desired regarding the amount and accuracy of the information on the old Trenton field, it was too important to be left out. We trust that we are not perpetuating any of the inaccuracies that are to be found in the data for Indiana.

Jan 1, 1935

By AIME AIME

MANY of the important papers to be presented at the World Congress of Engineering at Tokyo, in November, 1929, are being furnished by members of A.I.M.E. and a list of them is given below: "Fifty Years of Physical Metallurgy,,-by H. M. Boylston, Case School of Applied Science, Cleveland, Ohio.

Jan 1, 1929

By E. G. Banks

THIS paper is presented in the belief that metallurgists and chemists will be interested in the practice of grinding in tube-mills in connection with stamps, especially since the records of working here given extend over a lengthy period of time (since May, 1905). The ore from the Waihi mine-more especially that produced in the upper levels-contains a large proportion of hard, chalcedonic quartz, and the gold exists in an exceedingly fine state, conditions which necessitate very fine crushing in order to ob¬tain a high extraction of the precious metal. Before the introduction of tube-mills at the 90-stamp Waihi mill, it was found necessary to stamp through 40-mesh (1,600 holes per sq. in.) woven wire-screens, having a fairly high discharge. The pulp then graded

Jan 1, 1907

By Arthur Baker

The orebodies are long bedding-plane lenses of chalcopyrite and sphalerite, associated with garnetite masses. Most of the orebodies are within a 50-ft thickness of Cambrian limestone; other Paleozoic limestones and dolomites are locally metamorphosed but only slightly mineralized. Pre-mineral faults are numerous, but shallow folds were the main ore-localizing structures.

Jan 12, 1953

By Sheldon P. Wimpfen

EVER since the Phoenicians roamed the known world in quest of metals to harden their helmets and precious metals and gems to adorn their ladies, many other nations have sought metals in the far corners of the earth. During the past 75 years American engineers have been among the foremost in seeking for and developing ore deposits. They have been responsible for many of the principal discoveries and for the application of modern methods to mines in the far comers of the earth. American mining engineers with a zest for travel and an insatiable appetite for metals have also introduced throughout the world many articles made of metal which we consider essential to everyday life. Production and consumption of these metals in many foreign countries has brought to these countries an increased standard of living and added tax revenue to their governments. But unfortunately engineers from this country are no longer given ready access to all the countries of the world. This is not only unfortunate for American engineers but also for the countries that have adopted the exclusion policies and especially so since time has proved that in countries where the American engineers have gradually been excluded there has been a significant decline in the progress of the mining industry. Removal of these barriers to professional men would aid immensely in the world-wide development of the mineral industries. New areas brought into production under the guidance of American engineers with their exceptional know-how would provide more jobs for nationals of the countries involved and a better standard of living for the nations as a whole.

Jan 1, 1946

By James W. Guider

Of all the technology available to the iron blast furnace operator, raw materials preparation [(Fig. 1)] is by far the most important. Superior raw materials have been basic to the success of the Japanese steel industry, and they are just as essential to the American steel industry in winning its competitive battle against foreign government subsidized steel imports. Therefore, the mine operator and the ore beneficiation engineer must strive to produce the highest quality raw materials possible to meet the precise demands of the modem blast furnace and steel plant.

Jan 1, 1982

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Jan 1, 1938