Search Documents

HARRY B. BARREN Harry B. Barren, born in Cleveland, Ohio, May 31, 1888, died in Indiana Harbor, Ind., on Mar. 18, 1918. After graduating from the Case School of. Applied Science of Cleveland, class of 1910, he spent a year at the Youngstown, Ohio, plant of the Carnegie Steel Co. and the following two years with the National Tube Co. in Lorraine, Ohio. In 1913, he went to Indiana Harbor as assistant superintendent of the blast furnaces of the Inland Steel Co. Four months later, he was made superintendent. On Feb. 11, 1914, he was married to Miss Anna E. Jones in Cleveland and is survived by his widow and two sons, and by his parents Mr. and Mrs. H. A. Barren, of Cleveland, and a brother Kenneth, a lieutenant in the 309th Engineers, stationed at Camp Taylor, Louisville, Ky. CHAS. H. JOHNSON Charles Harmany Johnson died Dec. 10, 1918, at his home in New Castle, Pa., after an illness of more than four weeks brought about by a complete nervous collapse. He was born near Clinton, Lawrence Co. Pa., July 9, 1870, a son of George W. and Elizabeth A. Johnson. His parents moved to New Castle when he was two years old, where he attended the public schools of New Castle. He then attended Allegheny College, and later Massachusetts Institute of Technology where he graduated as a mining engineer. For a year and a half after completing his work at Boston he was connected with the Mesaba Ore Range in Minnesota, and then returned to New Castle and became associated with his father in the operation of the old sheet mill, which was converted into a tin mill and later sold to the American Sheet and Tin Plate Co. Since that time he has been actively engaged with his father in the limestone business and other industries. At the time of his death Mr. Johnson was general manager of the Pittsburg Limestone Co., the G. W. Johnson Limestone Co., the Lawrence Limestone Co., the Keystone Limestone Co., and the Mahoning Limestone Co., all of which have quarries in Lawrence, Butler, Blair and Armstrong counties and also in Maryland and West Virginia. He was president of the Johnson Bronze Co., the National Stone Co., and the

Jan 3, 1919

By Eugene McAuliffe

THE Coal Division undertook its most ambitious program this year at the New York meeting, with six full' sessions and a luncheon on Monday which had a record attendance and at which the newly elected officers for the Division were announced. Eugene McAuliffe, president of the Union Pacific Coal Co., Omaha, Neb., was unanimously selected as Chairman; J. B. Morrow, preparation manager of the Pittsburgh Coal Co., as Vice-Chairman; and H. E. Nold, Ohio State University, Secretary. During the past year two new and interesting activities were conducted by the Division directed at stimulating the interest of junior and student members. Reports were presented at the luncheon meeting by the chairmen of these commit¬tees. J. B. Morrow. chairman of the paper prize awards, and H. N. Eavenson reported on the work of the mining machinery manufacturers fellowship, the work on which was completed the first of the year. Several excellent papers are expected from these two groups for the next meeting of the Division to be held in Pittsburgh some time in the fall. The regular six-session program was filled from start to finish with interesting papers and much discussion, but unfortunately not enough time for the latter. The papers covered the field of prospecting, mining methods, preparation, utilization, mechanization, and safety, and all through the program we were constantly reminded of the value and need of research in the coal industry.

Jan 1, 1936

By Dol Duwez, R. H. Willens

A technique is described by which metastable alloy phases can be obtained by very rapid cooling from the liquid state. The results obtained so far have lead to; 1) extension of solubility limits beyond their equilibrium values, 2) new phases not found under equilibrium conditions, and 3) amorphous alloys. MANY new nonequilibrium phases have been found in alloys rapidly cooled from the liquid state. The technique used in these investigations essentially consisted of propelling a small molten globule of an alloy against the inside surface of a rotating copper cylinder. Since these original experiments were performed, the technique has been greatly improved. The purpose of this paper is to describe in detail, for the first time, a simple and efficient apparatus for producing foils of alloys which are rapidly cooled from the melt and to summarize some of the observations made on these alloys. I) FAST COOLING APPARATUS In the original fast cooling apparatus used by Duwez, Willens, and Klement,' the alloy was melted in a graphite container having the shape shown schematically in Fig. 1. A shock tube was attached to one end of the graphite part to supply the pressure for ejecting the molten globule. The globule was received on the inside of a rotating cylinder located about 0.5 in. from the exit of the graphite container. The globule trajectory made an angle of about 15 to 25 deg with the tangent to the cylinder at the point of

Jan 1, 1963

New York Holds monthly meetings, except June, July and August J E SPURR, Chairman A D BROKAW, Vice-chairman SIDNEY ROLLE, Vice-chairman M H MERRISS, Secretary-treasures, Nichols Copper Co, 25 Broad St, New York M F CHASE KARL HOFFMAN Boston Meets first Monday of each winter month C R HAYWARD, Chairman E C GEITHER, Vice-chairman D H McLAUGHLIN, Secretary-treasurer, Rotch Bldg, Harvard Univ Cambridge, Mass GEO H GILMAN GEO A PACKARD

Jan 1, 1923

By Donald R. Rathburn, Jung K. Mok

An historical review and a survey of current practice in the application of fine dry grinding in the iron ore industry are given. Both closed and open-circuit applications are reviewed, and the advantages, disadvantages, and process considerations for these two systems are discussed. The ground ore temperature rise, and the cool-ing of ground ore and mill ventilation are also discussed, along with the agglomerating characteristics of dry ground ore. Screening during the production of high-grade lump ore for blast-furnace burden, particularly from recently developed ore deposits in Australia, South America, and Africa, has resulted in large tonnages of fine iron ore. Certain owners have decided to stockpile these fines until some later date. Others have either sold these fines "as is" for agglomeration at the steel plant or have established their own pelletizing plants to produce a premium quality agglomerate. In almost all cases some grinding is required prior to pelletizing. Most of these ores occur in the weathered state which may result in difficulty in the thickening and/or filtering stages when ground in a wet system. The resulting filter cake would probably require partial drying to provide a product suitable for the subsequent balling operation; however the partial drying and subsequent material-handling steps often produce random agglomeration. Normally, these agglomerates will cause problems in the green ball formation process. Thus dry grinding has developed as a solution to this problem. Most of the dry iron ore grinding installations utilizing the standard-type rod or ball mill are grinding high-grade natural ores that do not require beneficiation. Although power consumed per ton of final product is 120% or greater in dry grinding as compared with wet grinding, the additional power required for dewatering and handling a wet ground product can easily exceed the initial power savings. Dry grinding of raw feed and clinker has been practiced in the cement industry for many decades but its major application to the iron ore industry has taken place only within the past ten years. The successful application of dry grinding to iron ore has required the solution of several unique design and operating problems. The purpose of this paper is to review recent developments in fine dry grinding of iron ore for pelletizing and to discuss the various designs and operating parameters involved.

Jan 1, 1971

By James E. Reynolds, Duane N. Goens, Charles W. Kenney

A process for recovery of lead and silver from high-grade lead sulfide concentrates by dry chlorination and fused salt electrolysis was developed by Hazen Research, Inc., for Cyprus Metallurgical Processes Corporation (Cymet). A pilot plant was designed to handle mixed sulfide concentrates and to produce high purity lead in a 2,000- ampere capacity fused salt cell. Design of the dry chlorinator and fused salt cell is discussed. Integration of modern chlorine caustic cells for reagent generation was evaluated in the pilot plant program. Also included in the paper will be process chemistry; pilot plant flowsheet, equipment and data: by-product and impurity removal: and scaleup and economic considerations for the process.

Jan 1, 1977

By C. C. Koch, M. L. Picklesimer

PHYSICAL metallurgists interested in the alloying behavior and phase transformations in rare-earth metals and alloys require a strain-free polished and etched surface for both metallographic and X-ray examination. However, the rare earths have presented a problem in this respect since they are difficult to polish free from strain and are very reactive to the atmosphere. Their tendency to oxidize readily even at room temperature necessitates careful preparation and rather elaborate protection if their surface is to be examined for any length of time. The "light" lan-thanides (i .e., atomic numbers less than 63), especially lanthanum and cerium, are particularly bad in this regard; they oxidize very rapidly at room temperature. To eliminate the problem of plastic deformation remaining on the surface after mechanical polishing, electropolishingl and chemical polishing techniques2 have been developed. The electropolishing technique requires the specimen to be immersed in an electrolyte at -77°C and thus obviates a metallographic study of, for example, the fcc —double-hexagonal close-packed (dhcp) transformation in cerium which begins at -10°C. The chemical polishing solution has been used only on several of the "heavy" rare-earth metals. However, a convenient method for protecting the polished and etched surfaces from atmospheric corrosion has not been available. Our interest in the superconducting properties of lanthanum, and in the phase transformations in lanthanum and cerium, has led us to develop a technique for the metallographic preparation of lanthanum and cerium which clearly reveals the microstructure and preserves the etched surface for days. The specimens are chemically polished and anodized. The chemical polishing is performed at room temperature, the anodic oxide film provides the required protection from the atmosphere, and colors produced in the film provide a simple positive technique for phase identification. Lanthanum and cerium metal ingots obtained from Johnson Matthey and Co., Ltd., were remelted to suitable shapes and heat-treated to produce various microstructures. The samples were placed in epoxy mounts and, after a small hole was drilled in the back of the mount for electrical contact, were ground on silicon carbide

Jan 1, 1968

By H. Baldauf, H. Schubert

The structure of flotation reagents is important for their adsorption on mineral surfaces and thus for their flotation effectiveness. The influence of the structure of organic depressants on the flotation of fluorite and barite with anionic and cationic collectors was investigated using a modified Hallimond microflotation cell. In addition, adsorption, desorption, and electrokinetic results were employed to further elucidate the role and the correlation of the structural parameters for adsorption and the depressing effect. As hydrophilic reagents, different aliphatic and aromatic compounds capable of forming chelates with alkaline earth materials were used. For the depressing effect, the kind, number, and position of the polar groups in the molecule, as well as the structure of the reagent films play decisive roles.

Jan 1, 1980

By J. L. Bruce, G. W. Nicolson

THE country rock of the Mavrovouni mine of the Cyprus Mines Corp. is hydrothermally altered, disintegrated pillow lava, with very little tensile strength ("short" ground). In places, especially when wet, it is heavy ground, though not extremely heavy except where movement has been caused by underlying or nearby stoping operations. Damage by caving from roof or sides of galleries is not sudden but progressively faster when the size of the cave increases as a result of the fall of small pieces. CEMENT AND CONCRETE LININGS During early operations, galleries in the country rock were supported by customary mine-timbering methods. Available timber, however, was of poor quality and of relatively high price, and replacements were frequent and expensive. Furthermore, the massive sulphide ore bodies required large volumes of fresh air for ventilation and cooling, and it was therefore desirable to provide smooth-lined galleries with low resistance for the introduction of air to the ore bodies. It was also desirable to eliminate hazards of fire in shafts and underground workings insofar as this would be practical. Cement and concrete were relatively cheap, and plans for using these in support of shafts, underground galleries, raises, and chutes, were developed steadily. Circular shafts with diameters ranging from 10 to 15 ft were adopted for all permanent shafts and airways. In places these were lined with concrete poured behind sectional pressed-steel forms especially constructed for this purpose. In other places, they were lined with precast-concrete blocks. It was found that such precast concrete of good quality could be produced on mass production scale and delivered at the mine from most favorable point of manufacture, at a cost of about 25 cents per cubic foot of concrete. This refers to prewar costs. As the use of concrete blocks in circular lined galleries was developed, it was found that there were very great advantages in the reduction of air resistance as compared with customary timber supports. After best methods had been developed, it was found that breakage and replacements were extraordinarily low, averaging only about 1 ½ pct per year during a period of about 15 years. Furthermore, it has been possible to reclaim for re-use a very large percentage of the concrete blocks from abandoned galleries. During the history of the operations precast-concrete blocks have been used for lining shafts, raises, ore chutes, galleries, and shaft stations, ranging from 3 to 16 ft id. The thickness of lining has varied from 5 to 10 in.; with I2 in. thickness in a few cases. Junctions of two galleries at angles ranging from about 30 to 60° are usually made of poured concrete behind sectional forms, or of a combination of poured concrete and precast blocks. In nearly all cases it has been found advisable to make the point of the wedge at the junction, of

Jan 1, 1947

By C. A. Queener, W. L. Mitchell

In recent years bismuth telluride has received considerable attention, primarily due to its value as a thermoelectric material. Since there seems to be little dependence of thermoelectric effect upon crystal orientation1 many of those who have previously worked with this material have been unconcerned with knowing its crystallographic angles to any degree of accuracy. However, when one studies deformation modes, lattice stresses, or the aniso-tropy of mechanical properties of a material, knowledge of the interplanar angles and their relationships as depicted in a stereographic projection becomes essential. Lange2 originally published crystallographic data which described the structure of bismuth telluride as rhombohedral with the space group A more recent work by Francombe3 presented data which corroborated and refined the original findings of Lange. In addition, Francombe established that there is no coexistence of rhombohedral and hexagonal lattices in bismuth telluride as was suggested by Vasenin and Konovalov,4 and that the structure can be treated as a pseu do hexagonal structure with an axial ratio, c/a, of 6.95 for convenience of presentation and study. The interplanar angles and standard projections for several unalloyed hexagonal metals have been previously published,5-9 but the present work appears to be unique in that it presents data for a rhombohedral compound in pseudohexagonal form. As a pseudohexagonal structure, this material has an extremely large axial ratio as compared to the usual range of approximately 1.5 to 1.9 for the hexagonal metals. Since the interplanar angles are a function only of c/a, the angular relations between various planes for Bi,Te3 will be grossly different

Jan 1, 1965

By B. T. Brady

Experimental data 1,2 on the energy dissipation in polycrystalline materials over a large frequency spectrum (10-2 cps to 106 cps) suggests that the quality factor [(e= Q )] is independent of the frequency providing both fluid-saturated and ferromagnetic materials are excluded from the test data. Additional data 3 suggest that the quality factors are less at high pressures than at low pressure, i.e., Q increases as the pressure increases. Observations on the dissipation of single quartz crystals with their polycrystalline counterparts indicate that the energy dissipation is less for the single crystal than for the aggregate .2 Both of the above observations, particularly the latter, are indicative of the fact that the dissipation increases with increasing surface contact within the polycrystal, i.e., a frictional attenuation scheme may be operative. Knopoff and MacDonald 4 have shown that a friction model gives rise to a quality factor frequency independence property. A partial listing of some values of Q, observed for both metals and non-metals, is given in Tables I, II, and III. In Tables I and II, the data all show Qs>Qp, i.e., the energy loss per cycle for compressional body waves is greater than that for shear waves of the same frequency. Most samples in Table III show the same effect. It is objective of this paper to suggest: (1) a frictional attenuation scheme may account for the observed losses incurred during the propagation of low amplitude stress waves in dry brittle rock and, in particular, that such a scheme can account for the observed relative magnitudes of the quality factors for P and S wave motion; (2) a friction scheme can readily explain the difference between the dynamic and static moduli in dry brittle rock; (3) a friction scheme, while in itself an intrinsically

Jan 1, 1968

By A. J. Sinclair

Sampling methods for feasibility studies are dependent on geological characteristics of the deposit being evaluated. A preliminary geostatistical study is required, taking into account autocorrelation aspects of variables under study in order that sample plan design can be optimized to produce grade and tonnage estimates within stipulated error limits. In general, regular sample distribution is to be aimed for, and regular sample lengths (from drill core or cuttings, for example) are preferred in order to facilitate statistical analysis. Sample plans are modified and improved (either technically or economically), as a sampling program proceeds and additional information is obtained. Bulk samples for metallurgical testing can be evaluated in terms of the variance shown by all such samples in a deposit and the internal variability of grade. Attention to these sources of variability will ensure proper evaluation of pilot test results from bulk samples. In addition, a general mineralogical study of a deposit is an important, often ignored, requirement in a more general but semi-quantitative evaluation of the representativeness of a bulk sample; and, moreover, provides information useful in optimizing plant design.

Jan 1, 1978

By J. R. Campbell

It is the purpose of this paper to review somewhat in detail the literature on the subject that is extant, which ought to provoke considerable beneficial discussion. The value of 1 per cent. ash in blast-furnace practice was brought to the attention of the writer when he was connected with the U. S. Steel Corpn. Various groups of blast furnaces in the Pittsburgh district compiled a statement in which the average saving was about 6 c. per ton on pig iron for every 1 per cent. of ash reduction. Later, the same groups of furnaces, when materials had reached a higher price, compiled a statement in which there was a 30-cent saving on pig iron for every 4 per cent. ash reduction in the Pittsburgh district, which was equivalent to 7½ c. for every 1 per cent. ash reduction. In the Chicago district, on long freight hauls of coal and coke from the Pocahontas district, the set-up was 30 c. per ton of pig iron for every 2 per cent. ash reduction, or 15 c. for every 1 per cent. Some years ago, Ralph H. Sweetser, of the American Rolling Mill Co., presented a paper which provoked considerable comment.' To get the matter again before this body a résumé of his computations is given below: Comparison of Pig Iron Costs Ash in Coal. Ash in CoKe, Fixed Carbon in Coke peR Ton OB PIG Iron peN Day, Per Cent. PeR CeNt. Coke. Per Cent. PIG IRon. Lb. Tons 6.00 8.70 89.80 1800 365.0 7.00 10.15 88.35 1860 353.2 Coke with 8.70 Per Cent. Ash Ore mixture, 4144 It). = 1.85 tons @ $5.50.............................. $10.17 Coke, 1800 lb. = 0.90 ton @ $6.50..................................... 5.85 Limestone, 1000 lb. = 0.50 ton @ $1.50................................. 0.75 Labor, $365 per day on 365 tons....................................... 1.00 Supplies and service, $250 per day...................................... 0.68 Overhead, $200 per day............................................... 0.55 Reserve for relining and depreciation................................... 1.00 Total............................................................. $20.00

Jan 1, 1931

By Bryan Denby

The curvilinear nature of shear strength envelopes for rock discontinuities may be simply and precisely represented by a power curve. The development of shear strength criteria applicable to rock discontinuities is reviewed and related to a theoretical basis for curvilinearity. An approach is reported to enable the quantification of the two power curve indices using a nomogram based on empirical studies. Using a knowledge of discontinuity basic friction angle and peak dilation angle the relevant indices ranges may be predicted for mine slope studies.

Jan 1, 1984

By L. Ibagos S., T. Balberyszski

The volumetric flow rate and the initial concentration of the copper sulfate solution influence the rate of reaction and, hence, the precipitation efficiency in a continuous launder cementation system. Using a model of such a system it was found that the precipitation efficiency is related to the two process parameters by the approximate expression [ log [1-n] = -[0.072 co 2.91 e – 0.511c o Q 1.15 ] N Q] where c is the initial concentration, Q, is the volumetric flow rate and N is the number of launders in the system. This expression agrees with an equation developed for the rate of the cementation process if it is assumed that the exponential factor represents the change in the surface area with concentration and the exponents 2.91-and 1.15 relate the hydrodynamic conditions and the surface concentration to co and Q. The experimental results are used to determine the values of a Precipitation Coefficient for any volumentric flow rate and initial concentration of solution. Knowing the Precipitation Coefficient, the total volume of launders that will result in a given precipitation efficiency can be calculated. It is established that maximum precipitation efficiencies are obtained for initial concentrations between 6.0 and 7.0 gpl and volumetric flow rates of approximately 350 m3 /hr Good agreement between calculated and actual launder volumes is obtained for a number of industrial operations for which a comparison is made.

Jan 1, 1973

By R. V. Ramani, R. D. Zigmond, R. L. Frantz

In this paper, a computer program for determining the capital and operating costs for coal preparation plants is described along with its applications. The logic of the computer program is based upon the results of a survey of existing preparation plants, operators, plant contractors, equipment manufacturers, and consulting firms. This survey provided information on plant performance, order-of-magnitude costs, and costing procedures. Additionally, an extensive cost data file for the major capital equipment in coal preparation plants has been developed. Using the washability data and the computer program out- puts, the cleaning costs per million Btu can be calculated. Details about the program and its applications can be found in a report by the authors (Zigmond et al., 1979).

Jan 1, 1983

By K. V. S. Sastry

First attempts to apply computers for the modeling, simulation and analysis of mineral processing unit operations date back to almost twenty years. Since then substantial efforts have been directed to the development of computer software for mineral processing. Current literature indicates considerable activities in the development of computer packages for interactive and batch simulation of individual unit operations and integrated plant operations. This paper presents a comprehensive review of various simulation systems that have been described in the literature along with a discussion of present trends in the development of mineral processing simulation software. Finally, the paper includes a detailed discussion of the desirable features in future simulation packages.

Jan 1, 1984

By M. L. White, J. L. Shafer, C. L. Caenepeel

Often an in situ leach is the only practical economic method for copper recovery from small low grade oxide deposits. The decision to develop a copper property by an in situ blast and leach is strongly dependent on ore grade, tonnage, hydrology, and the copper extraction and acid consumption rates. The inherent leachability of the ore will in part determine the desired amount of fracturing of the ore deposit by the blast in order to obtain a copper extraction rate that is economical. For the last several years Occidental Research Corp. (ORC), has undertaken a modest effort to address the question of how to best identify and determine the information necessary to predict and control the operation of an in situ leach for the economic recovery of copper.

Jan 2, 1979

By Malcolm J. McPherson

INTRODUCTION The use of computers to assist the mine ventilation engineer began to grow in the early 1950s when electrical analogue devices were developed for the analysis of mine airflow systems in Europe and the United States (l,2,3). The need for quantitative planning of airflows in mines had become pressing and analogue computers fulfilled a useful purpose for over a decade. From about 1964 onwards, digital computer programs for mine ventilation network analysis became available (4,5) and, superseding the analogues, spread rapidly throughout the major mining countries. The successful implementations of such programs resulted in further research being directed towards the development of simulation programs to model the quality as well as the quantity of air flowing through underground workings. It is no understatement that digital computers have revolutionised environmental planning techniques in modern mining industries. This paper discusses the complex interaction of natural and mining factors which governs environmental hazards in underground mines and further indicates some of the simulation techniques which have entered practical usage or are at an advanced stage of development. THE PROBLEM OF ANALYSING THE MINE ENVIRONMENT The traditional methods of approaching environmental problems underground were based largely on empirical or intuitive techniques. The reason for this lay in the intricate relationships which exist between the many variables that influence the mine environment - and which limit the use of purely analytical methods to the simplest of cases. In order to illustrate the need for simulation techniques, Table 1 shows the three most widespread mine environmental hazards

Jan 1, 1977

By N. De Nevers

A calculation method has been developed for carbonated water flooding. This method takes into account the effects of oil viscosity reduction and oil swelling, due to carbon dioxide transferred to the reservoir oil from the carbonated flood water. It shows the effect of changes in carbonation pressure and carbonated water slug size. The method is based on a Buckley-Leverett-type linear flow model. The mathematical approach is similar to that developed by Welge, et al, for enriched gas drives. However, it contains a very important advance over the previous papers in its solution to the problem of injection of a "slug" of carbonated water rather than continuous injection of fluid of constant composition. The method also shows the chromatographic nature of the transport of Co 2 in a carbonated water flood. Sample calculations indicate that viscosity reduction is the most important source of incremental oil recovery. Swelling contributes to a lesser extent. Raising the carbonation pressure increases recovery, but this effect diminishes as pressure is increased. increasing the slug size also increases recovery, but again this effect diminishes as slug size is increased. INTRODUCTION In recent years there has been considerable interest in carbonated water flooding as a secondary recovery method. Several carbonated waterflood field tests are now in progress.1'3 Yet there is no published method for predicting the results of a carbonated water flood. This paper presents a method for predicting the oil recovery for ; carbonated water flood. In a carbonated water flood. a carbonated water slug is injected into the reservoir. Then this slug is driven forward by injecting plain water behind it. The field operator must choose the carbonated water slug size and the CO2:water ratio in the slug. The calculation method presented here shows the effects of both slug size and CO2:water ratio on oil recovery. The calculation assumes that the addition of carbon dioxide to flood water increases the oil recovery by two mechanisms: 1. Carbon dioxide transfers from the flood water to the oil, thereby reducing the viscosity of the oil. At the lower viscosity, the oil flows more readily and is more easily swept out by flood water. 2. The carbon dioxide which transfers to the oil phase causes the oil phase to swell. Assume that the residual oil saturation (volume per cent of pore space) is the same whether the oil is carbonated or not. Then the same volume will contain less stock tank oil if it is carbonated, due to oil swelling. The plain water driving the carbonated water slug forward will strip the CO2 out of this residual oil., causing this oil to shrink. This swelling-followed-by-shrinking leaves behind a lower residual oil saturation than would have been left behind by a plain water flood. Since some oils swell to almost twice their original volume at reasonable CO2 pressures, this effect may be large. Two additional benefits sometimes claimed for carbonated water flooding are: 1. Adding CO2 to flood water greatly increases the water injectivity in some cases.1 (Injectivity = BPD of water injected/ft of sand x differential bottom-hole pressure, psi.) CO2 forms a weak acid in water; injecting carbonated water is, therefore, somewhat like acid treating. If a formation responds to acid treating, it will probably have a higher injectivity for carbonated water than for plain water. Although the increased injectivity increases the profitability of a carbonated water flood over a plain water flood, it does not change the production curve based on total fluid injected. Therefore, the method presented here does not show any effect of increased injectivity. 2. It is claimed in one patent4 that carbon dioxide will react with some constituents of the oil to form detergents. These detergents are expected to increase the oil recovery. There is little data available to support this idea. The calculation method presented here takes no credit for extra oil recovery due to the formation of any such detergents.

Jan 1, 1965