Search Documents

By C. H. Mahoney, P. E. LeGrand, A. L. Tarr

MAGNESIUM, it is now generally realized, differs in some important aspects from most other structural metals, not excepting even its close neighbors, the aluminum-base alloys. This is particularly true with respect to the effects of elevated temperatures on the two light metals when in the molten state. Aluminum, on the one hand, should not be raised above its melting point any higher than is absolutely necessary if one is to avoid danger of gas absorption and a tendency toward grain growth in the resultant casting.1 Magnesium alloys, in sharp contrast, actually benefit by being heated to temperatures considerably above their melting point. It is a commonly accepted practice to establish the desired fine grain in the industrial magnesium alloys of the aluminum-bearing group by raising the temperature of the melt after the refining treatment to some 25e C (482°F.) above the melting range. To recall briefly the typical practice for magnesium alloys, the molten metal is refined by fluxing at about 730° C. (1350 F.), and after the melt has been covered with a suitable flux, the temperature is raised to between 870°C. (1600°F.) and 930°C. (I700°F.). It is then cooled as quickly as possible to the casting temperature. This superheating, as it is called, results in a fine-grained cast metal structure which has superior mechanical properties, enhanced amenability to solution heat-treatment, and better machinability than cast metal that has not been superheated. The minimum temperature of superheating varies to some extent with the composition of the alloys to be treated and usually increases with the percentage of magnesium in alloys containing the same constituent metals.2 It should be noted that the grain-refining effect of superheating is markedly apparent only in connection with the alloys containing aluminum as an alloying constituent-the binary alloy containing r.5 per cent manganese, for example, is not refined in grain to any noticeable extent by a superheating effort. While the phenomenon of superheating is well known in the magnesium industry, the mechanism of the process has remained a mystery. During the superheating cycle something occurs that forces a fine grain upon the solidifying metal. The process has been controlled in an empirical manner, and occasionally, especially on large-scale melts, and for no apparent cause, no grain refinement is obtained from a superheating treatment. Only a few papers of importance have been published on the subject. Of these, one of the most helpful is the paper by Achenbach, Nipper and Piwowarski,3 which gives considerable data concerning the effect of superheating on various properties such as fluidity, shrinkage and hot crack formation, as well as grain size, and a discussion without any definite conclusions on the theoretical aspects of superheating. A short article by Murphy, Wells and

Jan 1, 1945

By W. O. Philbrook, K. M. Goldman, G. Derge

T. Rosenqvist—The most interesting point in this paper is the observed transfer of iron into the slag in the initial stage of the desulphurization process, after which the iron again is reduced to the metallic state. The authors interpret this observation as showing that the sulphur enters the slag as an iron-sulphur compound which subsequently is decomposed by the slag. The present writer has previously suggested the following equation for the desulphurization process: S + O2- ? S2- + O For equilibrium in the blast furnace the oxygen potential is defined by equilibrium with graphite and CO of 1 atm pressure: C + O ? CO [2] During the desulphurization process the reactions proceed in the direction of the arrows. If one assumes eq 2 to be significantly slower than eq 1, the transfer of sulphur into the slag, in accordance with eq 1, will build up a local oxygen potential at the metal-slag interface very much higher than that corresponding to the value defined by eq 2. This is possible because the equilibrium oxygen potential in eq 1 is high as long as the sulphur content in the slag is low. This oxygen potential will again be able to oxidize some iron: Fe + O ? Fe2+ + O2- and an increase in the iron content of the slag will be observed. Adding up eqs 1 and 3 one obtains: S + Fe ? S2- + Fe2+ The net effect is thus in harmony with the experimental observation but is obtained without assuming any close ties between the sulphur and iron atoms during the process. Furthermore, it follows from eqs 1 and 2 that when the sulphur content in the slag increases, and equilibrium with C and CO is finally approached, the local oxygen potential at the metal-slag interface will decrease, and the iron in the slag will be reduced back into its metallic state. C. E. Sims-—The data and conclusions presented in this paper are thoroughly convincing in establishing the mechanism of sulphur transfer from iron to slag as in a blast furnace. The evolution of gaseous CO in step 3 of the reactions given on p. 1112 makes the process virtually irreversible. Assuming that the process is similar in slag-metal systems other than in the blast furnace, it is readily seen why free CaO and re-ducing conditions so greatly favor desulphurization. On the other hand, the very effective desulphurization obtained in oxidizing slags when strongly basic, must be attributed to the relatively high stability of CaS as compared to FeS. The ease and simplicity with which the reactions of classic chemistry agree with the experimental data and explain the mechanism is noteworthy. The concept of molecules of FeS, soluble in both phases (metallic iron is not soluble in the slag), migrating from the iron to the slag and there reacting with CaO, which is soluble only in the slag phase, is clear and uncomplicated. This is likewise true for step 3. Those who would deny the existence of molecules or molecular-type combinations in liquid iron, must strain to provide a mechanism so lucid. In the absence of molecules, the Fe and S exhibit a remarkable collusion. L. S. Darken—The investigation and interpretation of rate phenomena in the range of steelmaking temperatures is a difficult task. Most of the laboratory investigations of steelmaking reactions have been concerned with equilibrium. Having determined the equilibrium, our attention naturally focuses next on the mechanism and rate of approach to equilibrium. The authors seem to have contributed substantially to our understanding of these factors for the case of sulphur transfer. I should like to ask the authors whether they consider that the sulphur transfer reaction is diffusion controlled as many high-temperature reactions seem to be. If so, it would seem reasonable to suppose that the slow diffusion step of the process is the transfer across a pseudo-static layer or film similar to that considered in heat flow problems. As the diffusivity and fluidity are smaller for the slag than for the metal, it may tentatively be assumed that the sulphur gradient exists in a thin layer in the slag adjacent to the slag-metal interface and that the metal and the main mass of slag are each maintained uniform by convection. On this basis the amount of sulphur transferred across unit area per unit time is D p (?S%)/100 ?1, where D is the diffusivity, p the density, (?S%) the difference in percent sulphur on the two sides of the layer, and ?l is the layer thickness. At the beginning of the experiment the main body of the slag and hence one side of the layer contains no sulphur; therefore (?S%) may be replaced by (S%), the sulphur content of the slag at the slag-metal interface, which in turn is equal to L[S%] where [S%] is the sulphur content of the metal and L is the distribution coefficient. The rate of transfer thus becomes DpL[S%]/100 ?l, which the authors designate K[S%]. Equating these two quantities and setting D = 10-6 cm2 per sec, p = 3 g per cm3, L = 40, and K = lo-+ g cm-2 sec-1, it is found that ?l, the film thickness, is about 0.01 cm—a value of the order of magnitude of that found in heat transfer problems in liquids. The uncertainty of the numerical values used leaves much to be desired, but at least it can be said that this calculation tends to support the proposed model involving diffusion through a film. Although this does not seem to affect the general argument, I should like to call attention to the fact that the diffusivity3 of sulphur in hot metal is found (on conversion of units) to be about 10-4 cm2 per sec rather than 104 cm2 per sec as stated by the authors. The three equations written by the authors to express the steps in the overall process of sulphur transfer may alternatively be written ionically as only two Fe + S = Fe++ + S-- Fe++ + O-- + C (graphite or metal) = CO (gas) + Fe where the underscore is used to designate the metallic phase; ionic species are slag constituents. After the authors have so neatly demonstrated that iron and sulphur transfer together (at least initially), this fact seems almost self evident; from eq 4 it is seen that if sulphur acquires a negative charge during transfer

Jan 1, 1951

By John M. Cummings, Warren. Harry V.

THE Geology Department of the University of British Columbia has undertaken the task of examining the ores from as many as possible of the gold mines of British Columbia. The object of this work is two- fold. First, it is known that detailed mineralographic studies are useful in attacking ore-dressing problems, and these studies are intended to be helpful to mill operators. Second, it is hoped that mineralogical and mineralographical studies will prove useful adjuncts to those of lithology and structure in throwing more light on problems related to the origin of the deposits. The paper shows that gold, for all practical purposes, is not present in an ore mineral in solid solution nor as a eutectic with any mineral; if the latter were true the gold particles would be distributed equally throughout that mineral. The gold occurs either as native gold or as a gold-silver telluride and is not distributed regularly through any mineral or minerals. The relationship between the gold or gold-bearing telluride and the associated minerals affords a means of classification that is not only simple but also appears to be directly related to the geological nature of the deposit, which may be summarized as follows: 1. Disseminated Gold Type.-Precious minerals in extremely small grains, which are distributed without regard to grain size, fractures or grain boundaries: (1) Hedley variety-gold in native form; (2) Surf Point variety-gold predominantly as telluride. 2. Veining Gold Type.-Precious minerals in grains of various sizes, which occur mainly along fractures and grain boundaries: (1) Cariboo gold quartz-goldin native form. (2) Waneta variety-gold as telluride. Application Te Ore Dressing These mineralographic studies have proved to be of value to mill operators, and future studies of a similar nature should prove increasingly useful because every year provides a further opportunity to correlate

Jan 1, 1937

By T. M. Chance

The first bituminous coal cleaning-plant to use the sand-flotation process1 was placed in operation on Oct. 1, 1925, at the tipple of the East Broad Top Railroad & Coal Co., at Mt. Union, Pa. The general principles governing the design of plants of this type and typical operating results of anthracite sand-flotation units have been previously outlined.2 I will, therefore, confine myself to a description of this particular plant and to some of the elements of design it incorporates. The Original Mt. Union Tipple The plant was installed to transfer coal.from steel narrow-gage cars of 60,000 and 70,000 lb. capacity to standard-gage equipment. The coal was hauled about 30 miles from the East Broad Top field. This coa1, mined by the Rockhill Coal & Iron Co. and other operators, is a low-volatile slow-coking coal with ash of high-fusion point. It is normally a high-grade coal that does not contain an excessive amount of free impurities, but it requires careful handling to secure a fair domestic yield, and, naturally, the need for narrow-gage transshipment increases the difficulty of the problem. The narrow-gage cars were discharged into an 8-car track hopper equipped with two feeders passing the coal to two 42-in. apron intake conveyors. Twin units were installed throughout the original plant, two lump-and-egg shaking screens, two lump picking-tables, two egg picking-tables and two stoker conveyors. The stoker coal was loaded without cleaning. The loading tracks and booms were in duplicate and, with the layout as originally installed, either unit could be run independently, each unit shipping a different class of coal or both units making the same product.

Jan 1, 1927

By D. P. Koistinen, K. E. Beu

THE application of the integrated intensity X-ray diffraction method to the measurement of retained austenite concentrations in hardened steels has been fully described.'-' In developing this particular method, the assumption was made that the samples used would have flat surfaces which were free from voids and in which austenite was homogeneously distributed. The original calculation procedure' was modified to include the constant factor G which relates the intensity of a given diffraction line to the concentration of the component giving rise to that line. In order to include A (0) in the constant factor G, it was further assumed that the geometrical conditions relating the X-ray beam, camera cylinder, and sample surface were precisely met.V has been demonstrated that these geometrical conditions can be met experimentally without difficulty in many cases;' however, for those samples which have irregular surfaces or where it is impossible to locate the sample in the proper position with respect to the X-ray beam and film cylinder, it is desirable to evaluate experimentally what effect these nonideal conditions have on the reliability of the austenite determination. This note decribes the results of such experiments. SAE 1095 bar stock was used throughout and all samples were heat-treated as follows: 1600°F for 1 hr in a neutral atmosphere, brine quench, and 300°F temper for 30 min. To control heat-treatment conditions as closely as possible, all samples were plated with 0.002 in. copper to protect the immediate surface from the heat-treating atmosphere. After heat treating, the samples were deplated electroly-tically so that only the copper was removed. The heat-treated surfaces thus obtained were nearly as bright and polished as the original surfaces before copper plating although no mechanical work was done on the samples after heat treating.

Jan 1, 1954

By Emerson F. Heald

A systematic study was made of new and old data on chemical activities in Fe-Pt alloys at elevated ternperatuves. Experimental results may be expressed in terms of the excess free energy using Least-squares analysis of the data gave the following values for the constants: for the temperature range 1130° to 1350°C, and tentatively to 155O°C, B = -3.326564 and C = 0.221051; for the temperature range 650" to 850°C, B = -2.555690, C = 1.762735, and D =0.097196. In the study of iron-containing silicate systems, it is sometimes desirable to have a direct experimental measure of the activity of iron in the system. The well-known solubility of iron in platinum, often a headache in experimental work on iron compounds under reducing conditions, can be used to advantage in this respect. If the activity of iron in an Fe-Pt alloy in equilibrium with the silicate is known as a function of the composition of the alloy, chemical analysis of the alloy will give a knowledge of the activity of iron in all of the phases in the system. The present study was undertaken in order to elucidate the characteristics of Fe-Pt alloys as iron activity indicators. This work is intended to tie together some previous work, which may be summarized as follows. Larson and Chipman1 determined the activity of iron in Fe-Pt alloys at 1550°C by equilibrating platinum metal with calcium oxide-iron oxide-silica melts of known iron activity. Compositions of the resulting alloys were determined by chemical analysis. A similar study was carried out by Taylor and ~uan,' who worked at 1300°C. They brought the Fe-Pt alloys into equilibrium with iron oxide under conditions of known partial pressure of oxygen, and thus, from the work of Darken and ~urr~,~ conditions of known iron activity. Compositions were determined indirectly, by following the change in weight of the sample. Sundaresen et el* used the electromotive force of cells in which the alloy formed one electrode in order to measure the activity of iron in the alloy at 650" and 850°C. These temperatures were chosen to be above and below the first-order phase transition which takes place upon the ordering of Fe3Pt and the second-order transition which occurs upon the ordering of FePt3. EXPERIMENTAL 1) High Temperatures. The starting materials used were thin platinum foil, about 0.002 mm thick, and Fisher Certified reagent ferric oxide, Fez03, which had been heated for 24 hr at 1000°C. An intimate mixture of 80-mesh Fez03 and platinum platelets was placed in a thin platinum foil envelope. The latter was suspended from thin platinum wires in the hot zone of a vertical-tube, platinum-wound furnace of the type described by Muan and ~sborn.~ A capillary gas mixer similar to that used by Darken and Gurry3 was used to prepare a precisely known mixture of carbon dioxide and hydrogen, which was allowed to flow upward through the furnace tube. The partial pressure of oxygen in contact with the sample was thereby fixed at a value which was calculated from the charts prepared by porter? Temperatures were measured with a Pt-10 pct Rh-in-platinum thermocouple, which was calibrated using the melting points of gold (1062 .@C) and diopside, CaMgSizOB (1391.5"C). Temperature control was maintained to within i3"C with a Geophysical Laboratory proportional controller, using the furnace resistance as the sensing element. Samples were quenched by passing a small current through the platinum suspension wires, allowing the sample to drop into a bath of dibutyl phthalate at the bottom of the furnace tube. Prior to chemical analysis the samples were washed with acetone and dried. 2) Chemical Analysis. It proved possible, in almost all cases, to separate the Pt-Fe platelets physically from particles of iron oxide. The platelets were dissolved in a small volume of aqua regia, evaporated to dryness, and redissolved to 0.1 M HC1. In order to determine iron in the platinum alloy potentiometrically, it is necessary first to remove the platinum. A 10-cm column of Amberlite IR-120 cation exchange resin in the hydrogen form provided separation quickly and quantitatively: The mixture of iron and platinum in 0.1 M HC1 was added to the top of the column, and washed with about 100 ml of 0.1 M HC1. Under these conditions, the iron, principally in the form of cations such as FeC1" and FeCl;, is held quantitatively in the uppermost centimeter of the column. The platinum, in the form of anions such as PtC&- , is washed through without being adsorbed. After a qualitative test with stannous chloride indicated all of the platinum was removed, the iron was

Jan 1, 1968

By F. M. Jardine

I1944 the cry was for higher production more tons, more metal. New plants were built, capacity of old plants was increased and millmen all over the country were treating tonnages far above normal, sacrificing recoveries to vet that extra metal from extra tons. The list of "critical minerals," at first long, has been scratched until there remains, not a list, but only an item or two. Stocks of all metals and minerals, with possibly one or two exceptions, are now entirely adequate and numerous plants have curtailed or suspended operation.

Jan 1, 1945

By S. S. Solomons, V. H. Tillmann, L. Smaniotto

INTRODUCTION The Zinc Corporation, Limited and New Broken Hill Consolidated Limited (ZC/NBHC) mines are situated at Broken Hill, New South Wales, Australia (refer Fig. 1). Both companies are wholly owned subsidiaries of Australian Mining and Smelting Limited (AM&S). Annual production from both mines is 2.3 million tonnes of lead, silver and zinc ore mined by a variety of underground methods. From 1911 to 1968, conventional overhand cut and fill stoping (both open and square set timber) was used almost exclusively, though some trials with shrinkage and sublevel stoping were undertaken. The trial stopes used short holes in conparison with today's sublevel open stopes. Mechanised cut and fill methods with regular rockbolt support were introduced in the early 1960's and undercut and fill stoping replaced square set timber stoping for pillar extraction in the late 1960's. Since 1968, increasing emphasis has been placed on longhole open stoping (LHOS) although some areas of the orebodies are comrnrnitted to the abovemehtioned methods since existing stope development makes it impracticable to change. In 1979, approximately half the total production was mined by LHDS and it is expected that the LHOS proportion will increase in the future. A vertical crater retreat method of longhole mining has also been successfully introduced to replace the undercut and fill stoping of some pillars and it is expected that this stoping method will be used in some future primary ore extraction. This paper reviews the progress of LHOS technology at ZC/NBHC with emphasis on the stope planning and design process and the importance of rock mechanics applications. The objective has been to maximise the primary extraction of the orebody giving due consideration to stability and safety. This has led to the design of a novel sloped wall longhole stoping method which should eliminate the formation of pillars by mining the orebody in a "continuous" manner incorporating a sloped sand interface between the blocks of ore being mined and the previously mined sandfilled stope. The expected results include the complete extract of the orebody without the formation of pillars and with minimal dilution of ore with sandfill. The evolution of the method can be best illustrated by following the mining progress of the relatively new Southern 'B1 Lode orebody at the NBHC mine. GEOLOGY Southern '8' Lode is the southern extension of the 'Bt Lode orebody, which is currently being mined further to the north in ZC/NBHC (refer Fig. 2). The ore block being developed and stoped extends for a vertical height of 90 metres from No. 17 level down to approximately ten metres above No. 19 level. The ore minerals are sphalerite and galena with the zinc:lead ratio averaging approximately 3:l with a minor amount of silver associated with the galena. There is a sharp contact between the ore and the country rock on the footwall and keel, while the hanging wall gradates into medium and low grade material. The gangue is primarily quartz together with garnet, biotite, feldspar, gahnite, sericite and minor apatite. In various horizons within the orebody, grey feldspar, quartz and garnet pegmatites occur as stratiform layers. The ore and country rock are competent with some inconpetent boundaries.

Jan 1, 1981

INTRODUCTION TO MINERAL INVENTORY The first explanations to be made must answer the questions: What is a mineral inventory, and how is a mineral inventory different from an ore reserve? The term mineral inventory is a relatively new designation-born of the computer age and grown to common usage with the development of computer techniques for extending drill-hole sample values to regularly gridded data blocks representing in situ mineral values. The term has been generally accepted by geologists and engineers because it has a precise meaning which avoids many of the ambiguities of ore reserve. A mineral inventory is an estimated inventory of mineral in place and is usually achieved by dividing the deposit into regularly spaced gridded blocks to which are assigned estimated values using various extension techniques. A mineral inventory makes no presumptions about the mineability of the blocks or their cutoff grade. Such parameters are implied with the use of the term ore reserve. Ore has a specific definition (see Chapter 1) and ore reserve must reflect that definition and thus has economic and mining constraints. An ore reserve will be different under different economic and mining conditions and is subject to reclassification by many factors. The term mineral inventory avoids all such ambiguities. A mineral inventory estimate may be good or poor according to how much is known about the deposit, how extensively it has been sampled, and what techniques have been employed to assign estimated grade to the blocks. Such an inventory, however, does not change with time, economic conditions, or mining technology, but only with new and better data and methods. Thus, the mineral inventory is an excellent designation for describing mineral value distribution within a deposit's geological setting. The techniques for assigning values and tonnage to mineral blocks or panels are the subjects of this chapter. Whether a summation of the blocks be called a mineral inventory or mineral reserve is not important. Many conventional methods for computing reserves of minerals refer to the methods in terms of ore reserve calculations. OBJECTIVES OF MINERAL INVENTORY OR RESERVES Mineral inventories are computed for the purpose of determining the quantity, the quality or value, and the spatial distribution of potentially economic minerals within the earth. Such computations are made at all stages in the life of a mineral extraction process, from the early exploration phase to the final year in the life of a mine. To quote Popoff (1966, p. 2), "They are the most responsible and irreplaceable tasks in the valuation of a mineral deposit. Efficiency in extraction and productiveness is impossible without accurate reserve computations." The accuracy of estimate required for each of the three parameters of quantity, value, and location is dependent upon the objective for which the reserve estimate is being made. Again quoting Popoff (1966, p. 2): Reserves are computed to determine the extent of exploration and development; distribution of values; annual output; probable and possible productive life of the mine; method of extraction and plan design; improvements in extraction, treatment, and processing; and requirements for capital, equipment, labor, power, and materials. Such computations are used to assist development

Jan 1, 1980

By C. F. Rust

Experimental evidence is presented showing that reproducible formation resistivity factor measurements and resistivity index determinations on reservoir core samples map be made utilizing either the two- or four-electrode methods. Equipment is described which permits the application of either technique to the core specimen without loss of time and with a minimum amount of effort. Since electrical resistivity measurements on reservoir rocks are important not only in electric log interpretations, but also in the study of fundamental rock parameters, it was considered desirable to compare the applicability of both techniques of resistivity measurement. INTRODUCTION The techniques employed in the measurement of electrical resistivity on reservoir rocks, as practiced in the petroleum industry, may be reduced to two fundamental methods: one utilizing two electrodes only, usually referred to as the two-electrode method; the other using two additional electrodes and commonly known as the four-electrode method. In the former, the core sample whose resistivity is to be measured is mounted between two electrodes, usually metal discs, which serve to pass a current of known magnitude through the saniple. The potential drop across the sample, measured between these same two electrodes, and the core dimensions furnish the necessary data to compute he resistivity of the sample. In the four-electrode technique the end plates, again usually ol metal, still serve as current electrodes. Additional probes or electrodes dispersed along the sample are utilized to measure the potential drop. Both methods are used in practice1,2 and it is the purpose of this paper to compare the two techniques as applied to the determination of formation reiistivity factors and the resistivity index exponent n, in Archie's empirical equation I = p/po = Sw 11. To evaluate the two techniques it is desirable to point out the experimental precautions which must be taken to insure reliable results. It is evident from the electrode arrangements of the two- and the four-electrode methods that, for a given sample, the volume of sample included by the four-electrode method is less than the volume included by the two-electrode method. One of the prime difficulties associated with the two-electrode method is the possible occurrence of an abnormally high resistance between the electrodes and the brine-saturated sample, which results in erroneous resistivity values for the core sample itself. Attempts to alleviate this difficulty meet varying degrees of success. Methods used in the past employ such absorbent materials as brine-saturated Kleenex: chamois, or felt. A technique described by Morgan, Wyllie, and Fulton,1 appears to be more satisfactory. Their method consists of spraying or painting the end areas of a core sample with conductive silver paint. This technique reduces contact resistances between the porous medium and the electrodes to a value which is generally negligible as compared to the resistance of the sample itself. In the four-electrode method these requirements are not as stringent since separate probes are used to measure the potential difference acrozs a portion of the sample. In this case, contact resistance occurring at the current electrode; are of no consequence. and those appearing at the potential

Jan 1, 1952

By M. H. Harris

A solution has been obtained to the problem of calculating flow into a cased and perforated well. Equations describing the idealized system were solved by numerical analysis techniques on a high-speed digital computer. Computer experimentation indicated that the results are within 3 per cent of the exact solution. Computed results are presented as an apparent skin effect, on a series of dimensionless working curves. Productivity ratios can be rapidly calculated from these data for the wide range of wellbore and reservoir properties actually encountered in field practice. Any reasonable vertical unisotropy ratio can be included in the computation. Productivities can be calculated for wells perforated in a regular pattern, a single horizontal plane or a vertical line on one side of the casing. Solutions for asymmetrical perforation patterns can he easily obtained from the working curves. Solutions for flow into sells completed with a horizontal notch are also provided. Results show that productivity can be increased by extending perforation penetration. This is especially true for shallow penetrations. The highest productivity is obtained when perforations are arranged in horizontal planes, as opposed to a vertical line on one side of the casing. Four or five perforations in one plane will approach the maximum productivity. The effect of perforation hole diameter is small, and is rioticeable only in shallow penetrations. INTRODUCTION The perforating method of well completion has been successfully and widely used in most producing areas. In the past, the desired perforation results were achieved by simply shooting numerous large holes in the casing. A perforating scheme for an area, or particular well, was established by evaluating two types of experimental data: the well flow index and the productivity ratio.' 3 The well flow index test is a means of determining the relative flow capacity of perforations in a linear system. This test also gives information on the perforation size, shape and hole damage which can be expected under well-bore conditions. However, relative flow data from this test cannot be accurately tarnsformed into the radial wellbore system. Thus, no true indication of the productive capacity of a perforated well can be obtained from the well flow index test. A well productivity ratio is derived from electrolytic models built to simulate perforated wells. Available productivity ratio data 2,3 are limited in that all important variables are not extensively evaluated. For example, data are limited to well radii of 3 and 6 in., and to maximum perforation penetration of 1 ft. Practically no information is available on the pattern of flow into a perforation. The effect of various perforation patterns on well productivity has not yet been determined and the effect of perforation diameter has not been thoroughly studied. Also, it will be shown that this type of electrolytic model data is probably in error by as much as 10 per cent or more. Modern completion techniques such as limited-entry perforating, single-plane fracturing, steam and hot air injection and chemical consolidation of incompetent formations all rely on injection and/or production from a limited type of completion. These completions may consist of several high-capacity perforations or a horizontal notch which is equivalent to an infinite number of perforations in one plane. Proper design of these new completion techniques requires, in part, accurate estimates of a well's productive capacity. Data on perforated and notched well productivity have been determined in this investigation. Results are presented as a function of the pertinent variables. Equations describing the idealized system were solved by numerical analysis techniques on a high-speed digital computer. DESCRIPTION OF PROBLEM PRACTICAL APPLICATION The computed results can be applied to wells which meet several general specifications. An external drainage radius of the well must be known or assumed, but the well need not be drilled on a regular pattern. The well must be cased and cemented through the entire producing interval. And, the casing must be perforated with one or more horizontal planes of perforations. The perforation pattern itself must conform to three geometrical requirements. 1. The perforations in a plane must divide the plane into equal angular segments (Fig. 1). Four perforations per plane would be 90" apart. 2. Perforations in a plane must be directly above or beneath perforations in adjacent planes. For two shots per plane, perforations would lie in the BB' plane (Fig. 1). 3. Horizontal planes must have equal vertical spacing and be separated from the top and bottom of the formation by one half the spacing interval. Single planes of perforations should be in the middle of the formation. This is shown in Fig. 2 where h is the formation thickness in a single plane conlpletion or the spacnig interval in a multiple plane completion. The spacing interval at the top and bottom of the formation may vary somewhat without causing an appreciable error in the productivity calculation if the formation is reasonably thick and there are many horizontal planes of perforations. A special provision has been made to calculate the productivity ratios of wells with uneven spacing intervals, or for the case where a single plane of perfora-

Jan 1, 1967

By H. E. Elliott, R. S. Busk, A. T. Peters

One of the problems of the fabricator of metals and alloys is the propensity of some composition rarnges toward abnoermal grain growth during certain stages of fabrication. In this respect magnesium alloys are no exception among alloys. Fortunately, convenient foundry control methods have been developed by which the abnormal grain growth possible in certain magnesium-alloy castings can be suppressed. An unusual characteristic of magnesium alloys, however, is that castings of some composition ranges are subject to local grain growth during heat-treatment without the necessity of intentional mechanical stressing of the part prior to heat-treatment. This effect occurs only under certain critical casting conditions, and only in certain alloys of greater than about 8.5 per cent aluminum content, such as C alloy (Mg, 9 per cent A1, 2 Per cent Zn, 0.2 Per cent Mn) and G alloy (Mg, 10 per cent Al, 0.2 per cent Mn). The first observation of this phenomenon in cast magnesium alloys occasioned much surprise, since it is not observed in other meta1s- Jeffries and Archer1 have stated that "the grain size of cast metals, provided they have not been plastically deformed, cannot be changed by heating below the melting point." The study of the conditions leading to abnormal grain growth in cast magnesium alloys, and of the mechanism of this phenomenon, has led to some very interesting observations. Perhaps the most significant of these was the observation that stresses originating from cooling contraction in magnesium-alloy castings are the only stresses necessary, under certain conditions, to cause local recrystallization and subsequent grain growth in the casting during heat-treatment. Before recent developments, this phenomenon of abnormal local grain growth Was a definite problem in the production foundry on C alloy sand castings. Illustrated (Fig. I) is a picture of a casting scrapped because of this defect. A mottled effect is produced in coarse-grained areas by suitable etching solutions. The properties of metal in the affected area are much impaired by such grain coarsening. Ultimate tensile strengths of half the normal value have been measured in metal of this nature. In this paper are described the techniques that have been developed by The DOW Chemical Co. for studying, and in large measure solving, this production problem. Also discussed are studies of the mechanism by which excessive grain coarsening occurs, and the fundamental conditions that produce this effect. The abnormal grain growth discussed in this paper will be referred to as "germinasium since this term has been definedl as any process leading to an abnormally coarse grain size. This paper deals only with abnormal local grain growth that may occur during the solution heat-treatment of certain magnesium-alloy castings, as dis-

Jan 1, 1945

By C. R. Breck

THERE has been a running discussion over the-past several years with respect to the life and adequacy of our natural gas reserves.. Some of the experts agree on one phase of the subject at least--that eventually it will be necessary to augment deliveries of natural gas in the northern markets by the development of a more or less comparable synthetic gas made from coal. The purpose of this paper is to point out: 1. It may be later than a good many people think-the demand for additional domestic and commercial pipeline gas in at least one section of the North will probably be felt rather keenly before an adequate synthetic pipeline-gas project, developed on economic lines, could be completed, even if ' initiated in the near future. 2. The evidence is considerable that such a project would be technically sound and economically self-supporting, although the margin of profit at first might be so small and the magnitude of an economic enterprise so great that financing would be impossible without Government aid in some form or other. The part of the country referred to is the heavily populated section in the northeast of the United States. To simplify an examination of the situation, let us take the potential marketing areas within range to the north and east of the three pipelines: 1. Tennessee Gas Transmission Co. as it enters Erie County, New York state. 2. Texas Eastern Transmission Corporation just before the tap that supplies the general Philadelphia area. 3. Transcontinental G& Pipe Line Corporation as it enters the state of New Jersey The sources of gas for these three pipelines are in Texas and Louisiana. The ultimate maximum combined capacity of the three lines at the points stated is approximately 1,200,000,000 cu ft per 24 hr at the maximum load factor probable of attainment in normal operation. The areas supplied by the three lines north and east of the points stated, or assumed as eventual markets, are: 1. Central New York state, from Buffalo to Albany (both inclusive).

Jan 1, 1953

By H. A. Wentworth

The Huff electrostatic plant of the United States Smelting CO., operated in conjunction with its wet concentrator at Midvale, Utah, was the second plant of substantial size installed using the Huff process, and the first plant to be put in operation on the so-called Western complex ores. , In spite of the machinery being of early design and consequently embracing many of the mechanical weaknesses incident to a pioneer plant the mill has operated steadily and uniformly since the summer of 1909, about five years now, without any material change other than some rearrangement of the machinery for better handling of the sizes. The ore, at present coming almost entirely from the company's mines in Bingham, is of practically the same composition as that upon which the plant was first put in operation. The crude ore, consisting of the sulphides of copper, lead, zinc, and iron, and containing small amounts of gold and silver, is brought by train and delivered to the hoppers of the wet concentrator, where by jig and table treatment, there are produced a shipping lead concentrate, a tailing, and a middling product, the latter being conveyed in push cars to the adjoining electrostatic mill. The crude ore delivered to the wet concentrator contains from 6 to 9.5 per cent. zinc. At times the plant is used also for the treatment of custom ores from the Bingham district and elsewhere, and at such times the results of the plant vary according to the material in use, but as by far the larger portion of the product is that from the company's own mines, the results given below are fairly indicative of the general work obtained. The diagram, Fig. 1, illustrates graphically the flow of the ore through the electrostatic mill. The middling coming from the wet mill, containing from 15 to 18 per cent. moisture, is hoisted while on the cars by an elevator, and delivered from the top of the elevator to a hopper over the drier. The drier first installed was of too low capacity to take care of the moisture present in the tonnage to which the mill was later

Jan 1, 1915

By D. G. Russell

A simple method has been developed with which flowing bottom-hole pressure data from two-rate flow tests in oil or gas wells can be analyzed to estimate the formation permeability, skin factor and average reservoir pressure. The required pressure data are obtained by observation of the transient bottom-hole pressure behavior after the stabilized producing rate of the well is changed to another, higher or lower, rate. The new method yields the same information as a conventional pressure buildup analysis, but eliminates the need for closing in the well. The analysis of a two-rate flow test is of the same degree of difficulty and requires about the same engineering time for application as a conventional pressure buildup analysrs. The extended closed-in periods experienced with conventional buildups because of long, low-rate afterproduc-tion periods are eliminated by fIow tests. Other anomalous pressure buildup effects, such as "humping" due to weli-bore phase segregation, can be successfully eliminated with the new method. Generally, flow tests of about 24 hours' duration run with conventional pressure measurement equipment are sufficient for interpretation purposes. Field testing of the two-rate flow test method has established that it is a reliable and economical method which can be used in many instances to complement or even replace conventional pressure buildup methods. INTRODUCTION The principal method for estimating formation permeability and well damage, or skin factor, in a producing oil or gas well is the analysis of shut-in bottom-hole pressure buildup data.' This familiar method has been used quite successfully by reservoir engineers for many years. It is based on the solution of the radial flow equation for constant rate conditions, and requires that the well be closed in for a sufficient period of time to obtain a clearly defined linear portion on the plot of observed t + ?t bottomhole pressure vs log t + ?/?t(where At is shut-in A? time, and t is producing time to the instant of shut-in). From the slope of the plot and other normally obtainable data, the permeability, skin factor, and reservoir pressure at infinite shut-in time (if the reservoir were infinite) can be estimated. Over the years several drawbacks have become apparent in the use of conventional shut-in pressure buildups for determining permeability and skin factor. The conventional pressure buildup interpretation theory assumes that a well is closed in at the sand face and that no production into the well occurs after shut-in. In practice, of course, the well is closed in at the surface, and inflow into the well continues until the well fills sufficiently to transmit the effect of closing in to the formation. This adjustment period is commonly referred to as the "after-production" portion of the pressure buildup. In the tight reservoirs, long, low-rate after-production periods frequently occur, and the well must be shut in for several days or, in some instances, even weeks to obtain an interpretable buildup curve.' Obviously, such long shut-in times can cause loss in current income, both from reduced oil production and from the fact that personnel and pressure measurement equipment are occupied with a single well for too long a time. In other cases, even long shut-in periods do not seem to be of much aid in obtaining an interpretable buildup." If there is considerable phase redistribution (liquid fallout or bubble rise) after a well is shut in, then curves with no interpretable portion are often obtained during the buildup. In addition to instances in which wellbore effects cause trouble, there are also cases in which the major objection to use of the closed-in pressure buildup is simply the fact that the well must be shut in. When there is no proration and when the well has limited producing capacity. closing in the well means loss of income. From the foregoing discussion it is apparent that it ib desirable to have an alternative method of obtaining the same information as that derived from a conventional buildup without the need of closing in the well. One possible solution which has been offered for this problem is the use of a bottom-hole shut-in tool' which isolates the major portion of the flow string from the formation face during the buildup. In this paper an alternative method which is frequently successful in avoiding wellbore effects and which does not require the use of special equipment is presented. A new, simple method has been developed with which the flowing bottom-hole pressure data from flow tests in oil or gas wells can be used to estimate permeability, skin factor and the average reservoir pressure. The required pressure data are obtained by observation of the transient bottom-hole pressure behavior after the stabilized production rate of the well is changed to another, higher or lower rate. The need for closing in the well is eliminated, and pressure measurement periods of only 18 to 24 hours arc usually sufficient, even in tight reservoirs. Thus, the new

By H. L. Riley, B. W. Gandrud

AS far as the present writers know, this system of flotation has not been used elsewhere in this country, but in the last couple of years it has been introduced, with minor variations, at one plant in England and one in Wales.' The system has been described and discussed in a number of publications.2-5 The following is quoted from an abstract of the latest of these,5 a paper presented at an International Conference on Industrial Combustion in 1952. On the basis of experience to date with the commercial plants, it is believed that the kerosene-flotation process incorporates all the necessary elements to make it greatly superior to anything else now available for treating of fines in wet processes of coal preparation. Additional study and investigation are still needed, however, to determine if certain phases of the process can be improved to such an extent as to make it generally satisfactory and acceptable to the industry. Further improvements will be needed with respect to the capacities of the flotation cells and the reagent consumption. The situation referred to above explains why an investigation is being made of the possibilities of achieving better cell capacities. Results obtained from this investigation, which is still in progress, are believed significant with regard to both cell capacity in general and the relation of cell design to cell capacity in particular. Commercial equipment now being used in a laboratory-type investigation should have performance characteristics similar to those of the larger machines. Equipment and Procedures: All flotation tests have been made in a standard Denver sub-A 24x24-in. unit cell of 12-cu ft volume. Cell modifications to make it more suitable for the tests were an adjustable front-wall section for varying cell depth and a perforated scraper-drag assembly for removal of the float product. There is also an apron dry-coal feeder, a gravity-feed water supply, reagent feeders, and a centrifugal pump that feeds the mixture of coal, water, and reagents into the flotation cell. A wattmeter connected into the drive-motor circuit records the power requirements of the impeller throughout each run. Dry coal, water, and reagents are all fed through a pan-type intake to the feed pump. A Sturtevant blower was set up to furnish air for supercharging. A centrifugal pump with a garbage-can intake provides for disposal of refuse flow to an outside settling tank. Figs. 1 and 2 show the flotation cell; Fig. 2 also illustrates the blower for supercharging. For purposes of this investigation, the percentage by weight of the feed coal recovered in the float product under a standard set of conditions has been considered as the criterion of cell capacity. The authors realize that such a criterion may be somewhat unorthodox, as the term cell capacity is usually understood to refer to feed input and ordinarily takes into account the ash analyses of the float product and refuse. However, the word capacity is flexible enough so that Webster gives one definition as maximum output, a definition which seems to justify, at least partly, acceptance of the above criterion. It has been the authors' experience in the Birmingham district that the ash-reduction efficiency of the coal-flotation process is generally satisfactory and that the only real problem is to increase the rate of float recovery so that the feed rate to any given bank of cells can be increased without undue loss of coal in the refuse. Originally it was planned to operate the flotation cell to simulate continuous operation during sampling periods. It was assumed that operating for reasonable time with feed coal, water, and reagents turned on would stabilize conditions so that the weight of float coal discharged during a fixed time interval would be an accurate measure of the rate at which the coal was being floated. It developed, however, that this supposition was erroneous. The float coal, caught for fixed time intervals and weighed, gave widely varying results in duplicate runs. Efforts to correct this trouble failed, and it was decided to try to operate on a batch-test basis, whereby all the float coal produced during a run on a known weight of feed coal would be caught in tubs, dewatered, and weighed. This method gives consistent and reproducible results, with total float product weight rarely varying by more than 3 or 4 pct on duplicate runs. The standard test procedure is as follows: A 132-lb sample of dry feed coal is weighed and placed in the feed hopper. The feeder is adjusted for a rate of 800 lb per hr. Feed water and reagents are turned on, and the feed and refuse pumps are started. One minute later the impeller is started. Six minutes are allowed for the cell to fill up with the water-reagent mixture. The feed of dry coal is started at the end of this 6-min period. One minute later the float-coal removal drag is started. The float coal is caught in one tub for the first 6 min after the flow of feed coal starts. Tubs are then changed, and the float coal is caught in a second tub until the feed coal runs out, when the tubs are again interchanged to catch the float coal for the remainder of the run in the first tub. The cell is kept running for 3 min with the water and reagents on after the feed stops to allow residual float coal to be removed. At the end of a test the wet float coal in both tubs is weighed and the total weight recorded. The product in the second tub is used for moisture determination and screen-size analyses. When the

Jan 1, 1956

By Allen D. Garrison

The chemistry of clays and shales has been assuming increasing importance in the petroleum industry, and two factors have greatly influenced this trend. The first has been the growing evidence that the marine shales are source beds, and that the petroleum compounds have migrated to the sand reservoirs overlying or underlying these shales: the second has been the demand for drilling fluids better adapted to penetrate these marine shales without difficulty and expense. Origin and General Composition of Clays and Shales Clays, shales and surface soils are all closely related materials. They are mixtures of finely divided compounds formed during the weathering of the solid crust of the earth. Their nature is essentially colloidal; that is, their behavior is determined by the state of division of the materials and the surface reactions of the smallest grains. Many of their chemical changes affect only the exposed areas and do not involve the interior of the grains of matter. The extent of these reactions is more nearly proportional to the area of the substances than to their mass, therefore it is possible for a small amount of a material in an extremely fine state of division to completely dominate the character of the mixture. More than 83 per cent of the crust of the earth is composed of three elements—oxygen, silicon and aluminum. Consequently, the corrosion products of the crust are predominantly aluminosilicates, or compounds of aluminum, silicon and oxygen. If we add the compounds of six elements—iron, calcium, sodium, potassium, magnesium and hydrogen—more than 98 per cent of the crust has been accounted for. These six elements, together with the first three, also appear in soils, clays and shales. But direct chemical analysis is of little value in describing the nature of such materials. Many chemical individuals may be present in variable amounts, and the extreme subdivision of any one of the individuals may contribute qualities that dominate the mixture.

Jan 1, 1939

By W. E. Coxe

In 1857 the Philadelphia and Reading Railroad Company, whose main line extended from Philadelphia to Pottsville, Pennsylvania, with branches into the coal regions of Schuylliill County, made a contract with the Fairmount Rolling Mill, at Philadelphia, for the rerolling of some four thousand tons of iron rails. The essential features of the agreement were that the old rails should be piled with puddled iron and rolled into flats for the rail pile, which latter was • to be of a section seven inches square, and after being heated, and reduced, by rolling, to a bloom of a section about fire inches by six inches, was to be reheated, before being finished into the rail of the T pattern, four inches high, weighing sixty-four pounds per yard. Great care was exercised in the execution of this contract, and the rails being distributed over all parts of the road, gave general satisfaction by their excellent wear. Five years afterward, the principal proprietor and manager of the Fairmount Rolling Mill was elected President of the Reading Railroad, and very many of the rails made under his supervision were still in use in the tracks of the company, of which he had just assumed the management. His first efforts were, therefore, directed to the procuring of more rails of the same quality wherewith to replace those worn out. Here he encountered some trouble, makers refusing to bid because of the details in the specification, the required reheating being the objectionable feature, for which it was generally intimated they would not be sufficiently remunerated for the increased labor, extra coal consumed, and the changes necessary from the established methods of working. Most of the new rails were, therefore, bought, and the old rails rerolled without regard to any particulars as to the manner of piling, heating, etc., each manufacturer furnishing what he deemed the best article possible for the money. As a consequence, rails of all imaginable grades were placed in the tracks, some giving out in six

By Allen D. Garrison

The chemistry of clays and shales has been assuming increasing importance in the petroleum industry, and two factors have greatly influenced this trend. The first has been the growing evidence that the marine shales are source beds, and that the petroleum compounds have migrated to the sand reservoirs overlying or underlying these shales: the second has been the demand for drilling fluids better adapted to penetrate these marine shales without difficulty and expense. Origin and General Composition of Clays and Shales Clays, shales and surface soils are all closely related materials. They are mixtures of finely divided compounds formed during the weathering of the solid crust of the earth. Their nature is essentially colloidal; that is, their behavior is determined by the state of division of the materials and the surface reactions of the smallest grains. Many of their chemical changes affect only the exposed areas and do not involve the interior of the grains of matter. The extent of these reactions is more nearly proportional to the area of the substances than to their mass, therefore it is possible for a small amount of a material in an extremely fine state of division to completely dominate the character of the mixture. More than 83 per cent of the crust of the earth is composed of three elements—oxygen, silicon and aluminum. Consequently, the corrosion products of the crust are predominantly aluminosilicates, or compounds of aluminum, silicon and oxygen. If we add the compounds of six elements—iron, calcium, sodium, potassium, magnesium and hydrogen—more than 98 per cent of the crust has been accounted for. These six elements, together with the first three, also appear in soils, clays and shales. But direct chemical analysis is of little value in describing the nature of such materials. Many chemical individuals may be present in variable amounts, and the extreme subdivision of any one of the individuals may contribute qualities that dominate the mixture.

Jan 1, 1939

By Allen Garrison

THE chemistry of clays and shales has been assuming increasing importance in the petroleum industry, and two factors have greatly influenced this trend. The first has been the growing evidence that the marine shales are source beds, and that the petroleum compounds have migrated to the sand reservoirs overlying or underlying these shales: the second has been the demand for drilling, fluids better adapted to penetrate these marine shales without difficulty and expense. ORIGIN AND GENERAL COMPOSITION OF CLAYS AND SHALES Clays, shales and surface soils are all closely related materials. They are mixtures of finely divided compounds formed during the weathering of the solid crust of the earth. Their nature is essentially colloidal; that is, their behavior is determined by the state of division of the materials and the surface reactions of the smallest grains. Many of their chemical changes affect only the exposed areas and do not involve the interior of the grains of matter. The extent of these reactions is more nearly proportional to the area of the substances than to their mass, therefore it is possible for a small. amount of a material in an extremely fine state of division to completely dominate the character of the mixture. More than 83 per cent of the crust of the earth is composed of three elements-oxygen, silicon and aluminum. Consequently, the corrosion products of the crust are predominantly aluminosilicates, or compounds of aluminum, silicon and oxygen. If we add the compounds of six elements-iron, calcium, sodium, potassium, magnesium and hydrogen-more than 98 per cent of the crust has been accounted for. These six elements, together with the first three, also appear in soils, clays and shales. But direct chemical analysis is of little value in describing the nature of such materials. Many chemical individuals may be present in variable amounts,. and the extreme subdivision of any one of the individuals may contribute qualities that dominate the mixture.

Jan 1, 1939