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By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By N. A. MacLeod

Ball wear in wet grinding mills has been the subject of considerable discussion in the last few years. Its importance to millmen is obvious in view of the fact that ball wear may cost from 2 to 4 per ton of ore for the more fortunate operators and many times this cost when a refractory ore must be reduced to a fine state of comminution. Consequently, any factual information that may be obtained on the nature of ball wear may be applied to introduce changes leading to lower ball consumption and therefore also to lower costs. In 1940, while employed with the Britannia Mining and Smelting Co., Britannia Beach, B.C., I had occasion to investigate factors that might affect ball consumption. We performed a number of tests, which will be briefly described. These tests originated as a result of changing from a 3½-in. ball to a 2-in. ball of different type of manufacture and were motivated by the desire to obtain the best possible type of 2-in. ball. Our testing was helped by the fact that the company operated a one-ton electric furnace, so that it was quite easy to make steel changes. It should be emphasized at the beginning that these tests produced results that in their general application to milling should be interpreted qualitatively rather than quantitatively. As will be evident in the description of the method used in testing, the balls were subjected to an intermittent batch grinding treatment, which may have produced results in rate of steel wear that may not have an accurate correlation with steel consumption in full scale mill operation. However, with this bed and rather arbitrary method of testing, results were obtained that may be of some interest and may stimulate further work on this subject. Unfortunately, owing to pressure of war work, our tests, which we had intended to repeat on a continuous grinding operation, had to be discontinued. Method of Testing Tests Were made in a Small cylindrical mill, rotating at 72 r.P.m. Internal dimensions of the mill were 10½-in. dia. by 9 in. The cylindrical inner surface was lined with rubber and the two ends were formed by steel plates. A charge of 2000 grams dry feed and 2000 C.C. of water was added evev 2 hr., the mill being drained of pulp before the addition of each charge, by tilting and sluicing with water. The mill was run On afternoon and night shift for daily periods of about 15 hr. total grinding time, and each test lasted from 60 to 150 hr. This batch method of testing involved a higher average pulp temperature in the mill, which undoubtedly affected comparative results in Some of the tests. Hardness oF Ball vs. Wear In these tests, special "eutectoid" steel balls were used, some of which were chill-Cut and quenched, some sand-cast and quenched, and some sand-cast and annealed to a dead soft condition. The results obtained, using as mill feed an ore containing considerable free quartz and

Jan 1, 1947

By N. A. MacLeod

Ball wear in wet grinding mills has been the subject of considerable discussion in the last few years. Its importance to millmen is obvious in view of the fact that ball wear may cost from 2 to 4 per ton of ore for the more fortunate operators and many times this cost when a refractory ore must be reduced to a fine state of comminution. Consequently, any factual information that may be obtained on the nature of ball wear may be applied to introduce changes leading to lower ball consumption and therefore also to lower costs. In 1940, while employed with the Britannia Mining and Smelting Co., Britannia Beach, B.C., I had occasion to investigate factors that might affect ball consumption. We performed a number of tests, which will be briefly described. These tests originated as a result of changing from a 3½-in. ball to a 2-in. ball of different type of manufacture and were motivated by the desire to obtain the best possible type of 2-in. ball. Our testing was helped by the fact that the company operated a one-ton electric furnace, so that it was quite easy to make steel changes. It should be emphasized at the beginning that these tests produced results that in their general application to milling should be interpreted qualitatively rather than quantitatively. As will be evident in the description of the method used in testing, the balls were subjected to an intermittent batch grinding treatment, which may have produced results in rate of steel wear that may not have an accurate correlation with steel consumption in full scale mill operation. However, with this bed and rather arbitrary method of testing, results were obtained that may be of some interest and may stimulate further work on this subject. Unfortunately, owing to pressure of war work, our tests, which we had intended to repeat on a continuous grinding operation, had to be discontinued. Method of Testing Tests Were made in a Small cylindrical mill, rotating at 72 r.P.m. Internal dimensions of the mill were 10½-in. dia. by 9 in. The cylindrical inner surface was lined with rubber and the two ends were formed by steel plates. A charge of 2000 grams dry feed and 2000 C.C. of water was added evev 2 hr., the mill being drained of pulp before the addition of each charge, by tilting and sluicing with water. The mill was run On afternoon and night shift for daily periods of about 15 hr. total grinding time, and each test lasted from 60 to 150 hr. This batch method of testing involved a higher average pulp temperature in the mill, which undoubtedly affected comparative results in Some of the tests. Hardness oF Ball vs. Wear In these tests, special "eutectoid" steel balls were used, some of which were chill-Cut and quenched, some sand-cast and quenched, and some sand-cast and annealed to a dead soft condition. The results obtained, using as mill feed an ore containing considerable free quartz and

Jan 1, 1947

By H. J. Hager, P. H. Haskell

Fox the past 20 years Alabama has probably led all other coal-producing districts in the proportion of coal prepared by wet washing. All rail mines, with one exception, and a high percentage of truck mines now wash all coal under 3 or 4 in. Most of this coal has been cleaned by jigs, though within the past few years one or more wettable plants have been installed for cleaning some of the coal from practically all the major seams. These tables supplement either single or multicell jigs usually for one of the following purposes: (I) to wash the finer sizes of raw coal, (2) to rewash the finer sizes of washed coal after they leave the jigs, (3) to retreat jig middling products. The main purpose of this paper is to present cleaning data and operating factors involved in the preparation by tables of the finer sizes of raw coal. The specific information and data pertain to the Mary Lee seam of coal as extracted at three of the mines of the Alabama By-Products Corporation, which produces coal for both by-product coke-oven consumption and the general steam market. The Mary Lee seam has long been recognized by coal-preparation authorities as presenting a very difficult ash-reduction problem. Washing characteristics and physical structure of the seam vary widely. Its thickness ranges from 32 in. with no rock parting to l00 in. with from one to three partings. Over the entire field the bed contains bands of bone and rash. These are distributed throughout the seam cross section. A portion of this bone or bony coal occurs as definite bands 1 in. or more thick; however, some bone is interlocked as layers within the coal and consequently breaks into flakes when the coal is broken. This material is relatively high in ash and its specific gravity varies widely. It has been observed that in most cases in the washing of Mary Lee coal from 10 to 18 per cent of the raw coal is within ±0.10 specific gravity of the attempted point of separation. Therefore with the average wet jig, only with extreme difficulty can a low-ash product be obtained without loss of much marketable coal in the refuse. Without entering into the controversial question as to whether modern washing equipment can efficiently handle simultaneously all sizes up to a given maximum, suffice it to state that the comparatively old-type jigs in service at the Alabama By-Products Corporation's mines could not clean the fine sizes satisfactorily. Therefore, in 1931 that company installed six diagonal-deck tables at the Barney mine, which were installed as an addition to a complete plant cleaning all coal below minus 3 in. The primary washed product is shipped as steam coal and another portion may be prepared as washed nut for domestic trade. These tables are washing 9 tons per hour of raw coal sized to ¼ in. to 0. The washed table coal is mixed with the 3 to ¼ -in. jig product to constitute a 3 in. to o steam coal.

Jan 1, 1941

By R. C. Amero

FULLERS earth is a general name applied to claylike minerals that have high natural ad-sorptive powers. They are usually distinguished from ordinary clays by a higher content of combined moisture and a lower apparent density. The definition is based on the ancient use of the material for fulling, or cleansing woolen cloth of oil and grease; a more modern definition should mention the ability to decolorize oil and should differentiate between fullers earth, which is naturally active, and certain bentonites which only develop decolorizing power after being leached with strong mineral acids. The distinction between fullers earth and activable bentonite is not a sharp one, since all types of intermediate clays can be found that are more or less naturally active, yet more or less responsive to acid treatment. The differences seem to depend somewhat on the amount of leaching that has been effected by natural ground waters. Physical Properties The principal mineral constituent of activable bentonite is montmorillonite, a hydrous aluminum silicate to which the following formula is generally assigned: Al,Si,O,, (OH), • 12H,O Fullers earth, as well as bentonite, is usually composed of some variety of montmorillonite, but in the United States the largest tonnage of fullers earth is mined from deposits of attapulgite. Attapul- gite is a clay mineral composed of monoclinic crystals of colloidal size, having a fibrous shape similar to that of asbestos. A sectional model of the attapulgite "molecule" is shown in Fig. 1. A single unit is about 18.0 A wide and 12.9 A high. The length of the molecule is variable, but this same molecular pattern is repeated to form needlework crystals, rather than flakes, as is characteristic of the montmorillonite crystals. Attapulgite has been assigned the following formula,' although actually, the magnesia has been largely replaced by alumina: (0%) . (OH), • Mg, . Si, • Olp . 4H20. The fibrous nature of the mineral is clearly shown in the electron micrograph, Fig. 2. Although the deposits contain 5 to 30 pct quartz and carbonate minerals, impurities in the commercial products are held at a low level, principally by selective mining on the basis of oil decolorizing efficiency. The particle size distribution of this fullers earth has been estimated from settling rates, using sodium citrate as a dispersing agent.' Since Stokes' law only applies to spherical particles, the equivalent particle diameter calculated below is probably more nearly an index of diameter rather than length of the needlelike crystals: Eqnivalent Arbitrary Particle Distribution, Size Diameter Pci Classil3catian 0.05 mm 0.80 0.80 pct sand 0.05-0.02 mm 10.96 0.02-0.01 mm 9.65 42.10 pct silt 0.01-0.005 mm 11.40 0.005-0.002 mm 10.09 2-0.5 micron 10.61 0.5-0.2 micron 29.09 56.58 pct clay 0.2-0.05 micron 14.86 0.05 micron 1.96 The differential thermal analysis shown in Fig. 3 is unique and supports the theory that this is a distinct mineral species.' ~ Other general properties of the mineral as shown in Table I.

Jan 1, 1952

MARSHAL Foch has been made an honorary member of the four national American societies of civil, mining and metallurgical, mechanical, and electrical engineers. The governing bodies of these societies, aggregating 45,000 members, conferred this signal honor in expression of the appreciation of American engineers for the unmatched services of this master of engineering principles. At Engineering Societies Building, in New York City, on Tuesday afternoon, Dec. 13, the presidents of the four societies presented one beautifully engrossed certificate of honorary membership in the four societies. The auditorium was filled so that not even standing room was left, more than a half hour before the cere-monies began, and the main entrance on the street had to be closed. J. Vipond Davies, president of the United Engineering Society, made the opening address. Col. William Barclay-Parsons, 'who was commander of the first engineer regiment to go abroad, addressed Marshal Foch in French. Colonel Parsons said: The art of engineering was defined a long time ago as "the art of directing the great sources of power in nature for the use and convenience of man." No better definition can be found today. Of all the sources of power in nature, the greatest, the most valuable, and at the same time the most difficult to direct, is the energy of man himself. He who can direct human energy and turn it to the service of mankind is a great engineer. You, Marshal, have directed a greater mass of human energy than any other man has ever done. And you have successfully directed this mass for the highest uses of mankind, in that you by its aid have preserved for him one of the most precious of human possessions-liberty-liberty not only for your own illustrious country, but for all the nations of the world. The four national engineering societies, of the United States now desire to make record of their appreciation of this fact and to convey to you an expression of their most profound admiration for the great leader of men by conferring on you honorary member-ship in all the societies, the highest honor in their gift and one hitherto never conferred on a single individual. Four thousand members of these societies were enrolled in the armed service of the United States, the greater part of whom had the glorious distinction to serve the common cause in France under your orders. They heard the voice and they saw the hand of the master as he led them through battle to victory. Now we desire that you will still continue to lead us, but in peace, by permitting us to inscribe your name at the head of our roll of honor, where it will be, as your deeds have been, an example to us to do better work, and where it will remain forever a noble inspiration for all future generations. Marshal, in the name of those members of the societies who served under your orders and particularly .in the name of those who fell on the field of honor in France, I salute you.

Jan 1, 1922

By A. Lee Barrett

MINE power systems are quite different in many respects from those usually found in industrial plants. Wide areas are served, usually by a circuit which is connected continuously throughout the mine. In some mines one circuit may run for several miles. There are usually a small number of branch circuits involved considering the areas covered. In all mines circuits are sectionalized at approximately 2500 ft by means of switches. All, or at least most branch circuits, are sectionalized by manual switches. In some mines automatic sectionalizing gear is used to give automatic overload protection at various strategic points in the system. However at the present time this practice is by no means general. Circuit hazards in mine electric systems are much greater than those encountered in industrial installation. Falls of roof or coal form a prominent hazard which may break the wire and may cause ground faults or partial ground faults. Ordinarily electric cables in mines cannot be insulated, there- fore n fall can easily cause a short circuit. The humidity in most mines is greater than 95 pct and at many points the floor is wet and water may be dripping from the roof over insulators, increasing the difficulties at that point. All electrical circuits must be contained in the mine entries and space must also be provided for moving equipment and men. Consequently, because of the proximity of the power circuit to the locomotives and other moving equipment a considerable hazard is introduced. Mine haulage track is not always good so that wrecks are not unusual. A wreck frequently does consider- able damage to the power circuit. A ground fault frequently results in heavy arcing which may be sustained. If the damage to the circuit is mechanical it may mean that a trip of coal has wrecked and a great deal of coal dust could be thrown into the air as a result of the wreck. The arc resulting from the ground fault could ignite the coal dust and a mine explosion would result. A continued ground fault, even if high in resistance, also may ignite the coal which may be on the bottom of the entry and a mine fire may result. The fire hazard is one of the most serious hazards in any coal mine. Since the mine circuit is completely interconnected, any power interruption resulting from a fault seriously affects the whole operation. Unless the mine is completely sectionalized the whole mine will be shut down until the fault is located and the proper switches are opened to isolate the fault; even then large areas may still be without power if they are supplied by a single power circuit which may be opened in isolating the fault. Further, mine power is usually supplied from a relatively small number of substations and the fault may be on a stub feed circuit, in which case power cannot be restored until the fault has been cleared. Practice varies widely with respect to placement of substations. Usually they will be spaced from I to 2 miles in 500V operations, and at 1 mile intervals or less, in 250 V operation. Actual mining operations govern substation placements and no general rules can be followed.

Jan 1, 1948

By G. W. Healy, N. R. Griffing

The activity coefficient of sulfur in Cr-Fe-S melts was determined by measuring the values of Ph3Rh,in equilibrium with such melts. The results showed that chromium has a pronounced negative effect on the activity coefficient of sulfur, which up to about 20 pct Cr may be represented by the equation: log f; = - 0.019 pct Cr It was also found that the influence of chromium was not altered by the presence of appreciable amounts of carbon or silicon. The influence of alloying elements on the activity of sulfur in molten iron has been the subject of many investigations; however the effect of one important alloying element, chromium, has not been reported, and the object of the present investigation was to determine its effect on the activity coefficient of sulfur in liquid Fe-Cr-S alloys both in the absence and in the presence of carbon and/or silicon. EXPERIMENTAL PROCEDURE The general method employed consisted of passing mixtures of hydrogen and hydrogen sulfide over the sulfur-containing melt of interest and analyzing the outgoing gas to determine the equilibrium value of which resulted from the reaction shown in Eq. 111. H, (gas) + S (dissolved in melt) = H,S (gas) 111 The activity coefficient of sulfur, fs, was then calculated with Eq. [2] in which K, is the known equilibrium constant of Eq. [I] for dilute solutions of sulfur in iron. Raw materials used in the preparation of the charges included commercial grades of electrolytic iron and chromium and carbon-reduced, acid-treated silicon of 99.9 pct purity. Oxygen in the former presented no problem since it was removed by hydrogen in the normal course of an experiment. Carbon was added in the form of premelted iron-carbon and chromium-carbon master alloys. Sulfur was charged as iron sulfide prepared by first heating powdered iron and sulfur in air in an alumina dish until reaction occurred, and then deoxidizing by melting in a hydrogen atmosphere. Commercially bottled hydrogen and hydrogen sulfide were used after being passed through purification systems containing palladium, "Drierite," and magnesium perchlorate in the case of hydrogen, and "Drierite" alone in the case of hydrogen sulfide. Lamp-grade cylinder argon served as a flushing agent without further purification. Charges weighing from 300 to 500 g were placed in a high-purity alumina crucible of 0250-ml capacity and heated in an argon atmosphere in the induction-heated carbon-tube furnace illustrated in Fig. 1. During heating, a dry hydrogen atmosphere was substituted for argon in order that deoxidation of the charge and flushing out of argon would be accomplished sooner. After the charge melted, a Pt-Pt 10 pct Rh thermocouple contained in an alumina protection tube was immersed in the melt, and the desired melt temperature (usually 1600°C) was established. Higher temperatures up to 1760°C were required for some chromium-rich alloys because of their higher melting points, and in these cases a Pt 6 pct Rh-Pt 30 pct Rh thermocouple was used. The thermocouple output was fed to an automatic controller which provided straight line temperature

Jan 1, 1961

By Dennis J. McCarthy

TAX TREATMENT OF EXPLORATION AND DEVELOPMENT EXPENSES Exploration Expenditures General Principles. Exploration expenditures are expenses incurred within the United States or the Outer Continental Shelf for the purpose of ascertaining the existence, location, extent or quality of any deposit of ore or other mineral. Section 617 of the Internal Revenue Code permits a taxpayer to elect to deduct these expenditures currently. In default of election, exploration expenditures must be capitalized and included in the basis of the mineral property for purposes of determining cost depletion. Entitlement to Deduction. In order to qualify for current deduction, exploration expenditures must be incurred prior to the commencement of the development stage, i.e., when mineral deposits are shown to exist in sufficient quantity and quality to reasonably justify commercial exploitation. Generally, expenditures for geological and geophysical investigations, reconnaissance, surveying, testpitting, trenching, drilling, driving of exploration tunnels and adits and similar types of work constitute exploration expenditures if incurred prior to the commencement of development. However, expenditures to further delineate the extent and location of an existing commercially marketable deposit are not exploration expenditures. Exploration expenditures may be incurred from within a producing mine to ascertain the existence of what appears to be a different ore deposit. Expenditures for property which is subject to an allowance for depreciation are not deductible under §617. However, depreciation properly allowable on assets used in exploration is an exploration expenditure. Election to Deduct Exploration Expenditures. The election to deduct exploration expenditures is made by deducting these expenditures in the return for the first taxable year for which such treatment is desired. The election applies to all properties of the taxpayer. Once the election is made all exploration expenditures, in the year of election and thereafter, must be deducted, unless the election is revoked with the prior consent of the Commissioner. The Tax Equity and Fiscal Responsibility Act of 1982 (TEFRA) made substantial changes in the treatment of exploration expenditures. For exploration expenses incurred after 1982 and before 1984 by a corporation, the amount currently deductible is limited to 85% of the expenditure. The remaining 15% must be capitalized and recovered over a five-year period. Although exploration costs capitalized under this provision are not part of the basis of the property for cost depletion purposes, investment credit may be claimed with respect to these expenditures. Further changes have been added by the Tax Reform Act of 1984 with respect to expenditures made after 1984. The Act reduces the amount of otherwise currently deductible costs from 85% to 80%. Additionally, it limits the availability of the investment credit for capitalized mineral exploration costs to deposits located in the United States. TEFRA also allows individuals a special method of recovery under which exploration expenditures may, at the election of the taxpayer, be amortized ratably over a 10-year period beginning with the year in which the costs are paid or incurred. If taxpayer does not use this method, the excess of the exploration expenditures actually deducted over the amount that would have been deducted if 10 year amortization had been elected is an item of tax preference subject to the alternative minimum tax.

Jan 1, 1985

By H. L. Gilbert, H. A. Johansen, R. G. Nelson

High purity electrolytic chromium plate has been hydrogen-reduced and arc-melted under inert atmosphere to give sound ingots. These ingots may be hot forged to break the as-cast structure and then worked at 500°C to finished form. An intermediate and final high temperature anneal gives best ductility and grain refinement. HE increasing interest in refractory metals for I use in high temperature gas turbines, jets, and rockets has led to great efforts in exploiting the fifth and sixth groups of transition elements. Much of the interest has centered upon production of metals that can be worked and fabricated by normal means to produce various engine parts or components having the high temperature characteristics of oxidation resistance and high creep strength required for such uses. If these pure elements do not exhibit all of the desired characteristics, it may be possible to obtain a high purity alloy that will. Since, in all cases, development of these transition elements has depended to a large extent upon removal of oxygen and nitrogen as a means for producing ductile metal, the hypothesis has been accepted that oxygen is the normal embrittling material in a high grade electrolytic chromium metal. Various methods have been employed in an effort to remove this oxygen by direct or indirect attack.' Many workers have endeavored to produce pure chromium through reduction of the oxide by carbon or active metals, while others have recommended fusion electrolysis of chromium salts or reduction of these salts by active metals, such as sodium or calcium. Hydrogen reduction of the oxide or chloride and thermal dissociation of the iodide have been reported to give high purity metal. Experimental Work The U. S. Bureau of Mines, Northwest Electro-development Laboratory, has systematically re-examined the more promising methods and developed two which produce metal having ductility at temperatures lower than heretofore described. It has been found' that oxide-free chromium chlorides in a NaCl-KCl carrier salt may be mag- nesium-reduced and the byproduct salts distilled off in vacuo to give a high purity chromium sponge. Fusion analysis conducted under the supervision of W. W. Horton of Knolls Atomic Power Laboratory showed the oxygen and nitrogen both to be below 10 parts per million on a selected sample. A small amount of carbon is carried over in the final metal from the chlorination and reduction steps which are carried out in graphite crucibles. This trace of carbon is not detrimental to hot ductility. A more practical method for producing ductile metal is through hydrogen reduction of high purity electrolytic chromium. Normal electrolytic chromium will, if "developed" for a few minutes at 1100°C in high vacuum, be found by the method of Adcock:; to contain about 1 pct chromium sesquioxide (Cr2O3). This small amount of oxygen may be effectively reduced by hydrogen.2, 4 The method employed was to crush electrolytic chromium plate in a gyratory crusher to pass 60-mesh; finer crushing appears to offer no advantage. A nitric acid leach removed the iron picked up during crushing. The leached powder was rinsed until neutral with distilled water and vacuum-dried.

Jan 1, 1954

By G. H. Lair, B. J. Langford, S. H. Rockwood

Volumetric reservoir analysis is dependent upon a firm relationship between porosity, connate water, and net pay. Capillary presswe data on core samples interrelate these three factors. It is shown that other reservoir problems may be resolved: (I) whether water production is indigenous or extraneous to the oil producing interval, (2) an oil-water contact is being approached, and (3) whether the oil-water contact is remotely situated. INTRODUCTION Volumetric analysis of carbonate reservoirs challenges the ingenuity of the reservoir engineer due to an absence of data on the minimum oil-bearing and oil-producing pores of a given reservoir rock. Many studies with this as the ultimate objective are available on the physical characteristics and fluid-flow behaviors of the carbonate reservoir. Some of these publications are: Hassler, Brunner and Deahl's1 study of the role of capillarity in oil production; Jan Law's' and A. C. Bulnes3 statistical treatment of core analysis data; Chalkley's method4,5 or estimating specific surface areas and porosity; Walter Rose's9 report on porosity, reserves, and recovery; Archie's78 tex-tural classification of carbonate rocks; Stewart and Spurlock's9 large core analysis; Purcell'sl0 mercury injection method in studying capillary pressure phenomena; Stewart, Craig, and Morse's" model multiple-phase flow test in investigating the relative permeability effect; and Fatt's12 illumi-nating network model of porous me-dia. The present study shows that the static reservoir parameters of porosity, connate water, and net productive thickness may be interrelated by utilizing a statistical arrangement of mercury capillary pressure data in addition to routine core analysis. This report summarizes the techniques that are followed to obtain these relations which are applicable to carbonate and sandstone reservoirs that do not contain clays. Capillary pressure statistical studies may be extended to include clay-bearing rocks by utilizing water capillary pressure data. POROSITY It is necessary that the laboratory determine the porosity value that includes the pores that are filled with hydrocarbons and water in the reservoir. This is the reservoir's total porosity. When true total porosity values are not available, correlations between porosity and capillary pressure characteristics lose their significance since they cannot be related through oil-base mud core data to the porosity and connate water values found in reservoirs. Methods used to obtain reliable porosity data will not be discussed; however, care should be exercised in selecting samples for measurement. In homogeneously porous formations, small core plugs sampled frequently will furnish adequate information. Where heterogeneity, particularly vugginess, is present, large-core porosity data may be required; in such reservoirs a suite of capillary pressure samples may be selected from small plugs. CAPILLARY PRESSURE Capillary pressure data provide a method for relating two carbonate reservoir parameters, porosity and connate water saturation, which in turn permit delineation of productive rock types necessary for selecting the third parameter, productive thickness. Specific applications of capillary pressure data to petroleum reservoirs were suggested by Archie8 and Pur-cell."' Their work demonstrated that in preferentially water-wet rock, porc

Jan 1, 1958

By George C. Lipsey

FOR many years the mining of coal and other minerals occurring in sedimentary formation has been planned according to information obtained by diamond drilling. Placer mines have used churn drills for sampling and estimating deposits, and some metal mines have used the method either wholly or in part. The diamond drill method, therefore, although not entirely new, has come into more prominence in the last few years. Several large-scale developments now being planned or in the production stage have used this method in varying degree. This paper will describe the operations of the Howe Sound Exploration Co., Ltd. at Snow Lake, Manitoba, where a gold mine was developed and equipped for mining and milling wholly on the basis of diamond drilling results. The property has been in operation nearly 3 years, so that results can be compared with the estimates and assumptions made from the diamond drilling. In developing the subject it is necessary to outline as briefly as possible the normal method of mining, from discovery to production stage. Table I sets forth these various stages, the methods of financing, the percentage of total cost, and the relative time interval required for each section of the work. The plan presented in this table cannot be applied to all mining developments, but it gives a basis on which to work. In Stages 2 and 3 the investment is moderate, and if the property does not respond favorably the options or interests can be terminated. Stage 4 represents a sizeable investment, and it is more difficult to drop the property after this part of the operation is completed. Stages 2, 3, and 4 may be repeated several times by diflerent groups or companies before Stage 5 is reached, but when Stage 5 has been started the total capital investment must be provided and there can be no turning back. It will be noted that there is a total of 116 units of investment in the overall cost if one company proceeds with Stages 2, 3, 4, and 5 in order. Stage 4, however, consumes time which is important when markets are being considered. Bringing a property into production when metal prices are low is much more difficult than when the prices are high, and if this stage can be set aside production is attained more quickly. When Stage 4 is eliminated the total becomes 86 units of cost for Stages 2, 3, and 5, but an adjustment of 14 units must be added to Stage 5. These 14 units represent the useable value of Stage 4. Thus Stage 4 represents an investment of 30 units if carried on before Stage 5, but 14 units if included with Stage 5. This is accounted for by the fact that development work can be accomplished more cheaply under Stage 5, since everything that is done is planned on a production basis with permanent services. Stage 4 must of necessity be governed by expenditures, and a great many temporary services, including housing and power plants, must be provided. The prospect shaft is a small one, and drifts,

Jan 1, 1953

By H. Siebel, N. A. D. Parlee

MIZIKAR, Grace, and parlee' studied rates of absorption of oxygen in stagnant liquid silver in a new type of open-ended gas-liquid metal diffusion cell, where the capillary diameters used were 1.5, 2, and 3 mm. They reported that the rates appeared to be controlled by the rate of diffusion of oxygen atoms in liquid silver and they used their data to arrive at the following equation as a reasonably good representation of the diffusivity (D) of oxygen in liquid silver: D = (14.7)10-4 exp (-7100 ± 3200)/RT cm2/sec Since this type of diffusion cell offers a promising method of measuring diffusivities of gas atoms in liquid metals, the senior author of this note and his students are carrying on a critical examination of the method, and are attempting to improve and extend the technique. The research reported briefly here was designed to examine more carefully the earlier conclusion1 that the absorption rate is diffusion-controlled, and to find out whether the diameter of the cell could be greatly increased and thus extend the method to systems where gas solubilities and diffusion rates are much smaller than in the case of oxygen in silver. For the purpose of these experiments, the open-ended capillary cell used by Mizikar, Grace, and Parlee was replaced by a small cylindrical crucible (1 to 1.7 cm in diam) containing pure silver (to a depth of 1.3 to 3 cm) resting in the bottom of a closed-ended vertical quartz tube, with an evacuated quartz filler tube just above the crucible to reduce dead volume. The method of measuring the rate of absorption was otherwise very similar to that of Mizikar, Grace, and Parlee and all the same precautions were taken. A much smaller and shorter tube furnace was used for heating, but simi- lar precautions about temperature gradients were observed. This technique was nothing more than the application of a Sieverts apparatus to the measurement of rates of absorption in small crucibles of stagnant liquid silver. In each run the volume of gas absorbed (V) was plotted against the square root of time (vt) and apparent diffusivities were calculated from the slope of the line, as described by Mizikar, Grace, and parlee.' Care was taken to use only the data obtained during the early part of the total absorption period while the cell functioned as one of infinite depth. The results at 1000°C are summarized in Table I and Fig. 1. The following conclusions may be drawn: 1) Diffusion cells up to 1.7 cm in diam yield, within experimental error, the same values for the diffusivity of oxygen (D) as the capillary cells of 1.5, 2, and 3 mm in diam employed by Mizikar, Grace, and Parlee. Cell diameters as large as 1.7 cm are probably not desirable for the most precise work. Intermediate sizes, perhaps 0.5 cm, will be most practical. 2) All absorptions measured were consistent with a model where the rate is controlled by the diffusion of gas atoms in the metal. The fact that runs at absorption pressures ranging from 0.1 to 1 atm (at any one temperature) give

Jan 1, 1965

By Tracy L. Atherton

THE mapping of slant-hole oil fields is complicated by the fact that relationships between wells are subject to variations in three dimensions and are not readily adaptable to representation on a plane surface. Formalized perspectives usually expand into an undecipherable maze of lines and peg models with the necessary supporting dividers serve only to awe the beholder with their perplexity. The usual compromise arrived at is a plotting of the projection of the courses of all of the wells in the field onto a horizontal plane (Fig. I). These complicated plots, when used in conjunction with the individual vertical projections of the wells, heretofore have been the most practical foundation for the study of general field problems. This method has many shortcomings, however, which are not found in the recently developed straight-hole equivalent projection method. The new method supplies a series of maps similar to the type now in common use in straight-hole fields. It is remarkably simple to use and supplies a ready check on clearances between wells. Perhaps the method can best be understood if the steps that led to its development are described. Common practice of oil-well surveyors is to supply the owner with a tabling of the survey and a plan view plotted against the true meridian. If a vertical projection is plotted, it is usually plotted separately and shows vertical depths against the projection of the stations along a section line joining the top and bottom of the hole on the plan view. A more practical method would be to plot the section line of the plan parallel to the abscissa of the vertical projection (Fig. 2). If this is done, any line parallel to the "base line" will pass through the points on each projection that represent a single point in the hole. Another very useful graph, which the well surveyor could provide, would be plotting of vertical depths against measured depths. Fig. 3 illustrates the next step facilitated by this form of plotting. The lines designated "well A" represent the proposed course of a hole to be drilled. Superimposed on the drawing, in its proper relative position, is the plan view and a portion of the vertical projection of a present producer, "well B." From an inspection of the plan view we arrive at the correct conclusion that the only indicated possibility of collision is at the point of crossing of the wells on the plan view. Referring to the vertical projection, note that the vertical depths of stations on B, in the vicinity of the crossing, are plotted on the same plane of projection as A. The clearance is the indicated distance between the two vertical sections measured along the parallel to the "base line," which passes through the point of intersection of the plan view. Repeat this process for the other wells crossing the proposed course of A on the plan view and you have a preliminary check of clearance for the proposed course. A similar comparison of crossings on the vertical projections of all

Jan 1, 1941

By George C. Lipsey

FOR many years the mining of coal and other minerals occurring in sedimentary formation has been planned according to information obtained by diamond drilling. Placer mines have used churn drills for sampling and estimating deposits, and some metal mines have used the method either wholly or in part. The diamond drill method, therefore, although not entirely new, has come into more prominence in the last few years. Several large-scale developments now being planned or in the production stage have used this method in varying degree. This paper will describe the operations of the Howe Sound Exploration Co., Ltd. at Snow Lake, Manitoba, where a gold mine was developed and equipped for mining and milling wholly on the basis of diamond drilling results. The property has been in operation nearly 3 years, so that results can be compared with the estimates and assumptions made from the diamond drilling. In developing the subject it is necessary to outline as briefly as possible the normal method of mining, from discovery to production stage. Table I sets forth these various stages, the methods of financing, the percentage of total cost, and the relative time interval required for each section of the work. The plan presented in this table cannot be applied to all mining developments, but it gives a basis on which to work. In Stages 2 and 3 the investment is moderate, and if the property does not respond favorably the options or interests can be terminated. Stage 4 represents a sizeable investment, and it is more difficult to drop the property after this part of the operation is completed. Stages 2, 3, and 4 may be repeated several times by diflerent groups or companies before Stage 5 is reached, but when Stage 5 has been started the total capital investment must be provided and there can be no turning back. It will be noted that there is a total of 116 units of investment in the overall cost if one company proceeds with Stages 2, 3, 4, and 5 in order. Stage 4, however, consumes time which is important when markets are being considered. Bringing a property into production when metal prices are low is much more difficult than when the prices are high, and if this stage can be set aside production is attained more quickly. When Stage 4 is eliminated the total becomes 86 units of cost for Stages 2, 3, and 5, but an adjustment of 14 units must be added to Stage 5. These 14 units represent the useable value of Stage 4. Thus Stage 4 represents an investment of 30 units if carried on before Stage 5, but 14 units if included with Stage 5. This is accounted for by the fact that development work can be accomplished more cheaply under Stage 5, since everything that is done is planned on a production basis with permanent services. Stage 4 must of necessity be governed by expenditures, and a great many temporary services, including housing and power plants, must be provided. The prospect shaft is a small one, and drifts,

Jan 1, 1953

By R. C. Rumble

To /neet the need for a subsurface flowmeter capable of measuring low rates of flow, a new type flowmeter has been built and tested in the field. This instrument is self-contairzed, can be run on a wire measuritzg line, and may be run through 2-in tubing. It consists of a rotameter flow-measuring assembly used in conjuncfion with a conventional subsurface instrument clock and chart. Integral with the tool is an inflatable packer which, by effecting a seal between casing and tool, diverts all fluid through the metering assembly. Packers of different size adapt the tool to different sizes of casing or liner. The flowmeter is capable of measuring rates of flow as low as 4 BID. The flowmeter has been run experimentally in seven wells. It has yielded satisfactory resrrlts in five of the wells, and in the other two it failed to operate because of loose sand in the hole. INTRODUCTION One of the oldest problems facing the producing branch of the petroleum industry is the location of depths of entry of fluid into a wellbore and the determination of the amount of fluid entering at each depth. This problem is often encountered in oil and gas wells completed in single producing horizons as well as in multiple pays; it is a common problem in oil wells which produce excessive volumes of gas. Several subsurface instruments designed to detect and meter flow in the producing interval of wells have been described in the literature.1,2,3,4 Most of these, however, were apparently intended for use in wells producing at relatively high rates of flow, and consequently lack sensitivity to low flow rates. Because of this and of other possible disadvantages inherent in these instruments, development of the subsurface flowmeter described in this report was undertaken. DESIGN OF FLOWME'I'ER Design Requirements In considering the requirements for a flowmeter that would not be subject to the limitations of devices existing at the time work was started, attention was first given to the size of the tool. Such an instrument should be of sufficiently small diameter that it could be lowered readily in 2-in tubing in a producing well. The flowmeter should contain a recording unit so that it might be run on a wire measuring line with conventional reeling equipment. It should not be too long to be accommodated by a portable lubricator of the type used for running conventional subsurface instruments. The metering mechanism of the flowmeter should not be affected appreciably by the viscosity or density of the flowing fluid. Such an instrument should record flow rates in the range from less than 10 B/D to over 100 B/D. To achieve this, a packer should he provided to divert through the instrument all fluid flow in the wellbore. This packer should be of such nature that in the event of excessive flow the fluid would bypass the instrument instead of forcing it up the wellhore. General Description A flowmeter incorporating most of the desired features outlined above was developed and built. The instrument is an elongated tubular device terminated on

Jan 1, 1956

By John B. Campbell, George G. Binder, Loren H. Jenks

Conventional methods of producing crude oil generally leave in the depleted reservoir 1 to 3 bbl of oil for every barrel brought to the surface. This paper describes a field test which explores the possibility of using gas-driven liquid propane to improve oil recovery efficiency. Laboratory theoretical and experimental studies leading to the field test indicated that low residual oil values could be obtained using limited amounts of propane driven by gas. The field test involved (I) repressuring the reservoir to the vapor pressure of liquid propane, (2) injecting the liquid propane, and (3) injecting dry natural gas to drive the propane and oil to the producing wells. Considerable preliminary work, including new drilling, work-overs, and a pressure survey was necessary. During the test, oil production rates have averaged about nine times the final primary rates. The field average GOR is still below the pre-test value. Oil recovery to date is 112,000 STB. About one-third of the 143,-000 bbl of propane injected has been produced to date. Results of the C-2 block test are encouraging and larger, more definitive field projects ore indicated. INTRODUCTION The cost of finding new crude oil reservoirs continues to increase. Today the oil industry is drilling more and deeper wells and making fewer big discoveries. Nevertheless, the demand for oil and oil products increases each year. It is natural that oil companies, faced with this situation, concentrate on recovering more of the oil existing in our known crude oil reservoirs. There is room for much improvement. For every barrel of oil brought to the surface by present day methods of producing crude oil, generally 1 to 3 bbl are left behind in the depleted reservoirs. The "unrecoverable" oil in petroleum reservoirs is largely "trapped" by capillarity. Since flushing with a solvent is exceedingly effective in combatting oil loss due to capillarity, much effort in this direction has been made in recent years1,2,3,4,5. Liquefied petroleum gas (LPG) has been suggested as a relatively inexpensive. abundant, easily recoverable solvent for crude. Some operators have observed striking responses in oil recovery at production wells offsetting input wells in LPG storage reservoirs6,7,8. In spite of the low unit cost of LPG, however, it is uneconomic in general to inject a volume of LPG into a reservoir to remove a like volume of oil. A modification of solvent flushing in which a limited quantity of liquid propane is gas driven through a reservoir appears capable of recovering much more than 1 bbl of oil per barrel of propane injected. Also, most of the propane injected should be recoverable for sale or re-use. Since such a process could well have widespread and profitable application, The Carter Oil Co. is making an extensive evaluation of the technique's possibilities. Laboratory development of the gas-driven propane recovery technique was sufficiently encouraging to justify testing in the field. Since 1952, the Production and Research Departments of The Carter Oil Co. have cooperated in a field test in the C-2 Block leases of the

Jan 1, 1958

By C. E. Westley, R. R. Beebe

Recent works by Plaksh et al.,1,2 Glembotsky,3 and others have renewed interest in the effects of oxidation on the surface activity of the sulfides. Since considerable supporting data are available on galena-potassium ethyl xanthate, this system is being studied at South Dakota School of Mines and Technology by contact angle measurement, using the technique of McGlashan and McLeod.4 The contact-angle method is well adapted to investigation of the influence of small amounts of oxygen on the establishment of a stable contact angle on galena in potassium ethyl xanthate solution. Samples of massive galena containing less than 0.001 pct Cu, 0.001 pct Ag, and 0.01 pct Zn are mounted in castable plastic and ground and polished manually in a glove box containing the desired proportions of oxygen and purified nitrogen. Oxygen and nitrogen contents of the test atmosphere are determined with an Orsatt analyzer. In this way a reproducible surface is created in contact with whatever oxygen concentration is desired. Water and all reagents are prepared and stored in the test atmosphere. An absorbent is used to control carbon dioxide that might be present. Final polishing is done with levigated alumina on a glass plate. The prepared sample is placed in the test solution and transferred to the contact angle apparatus in a sealed box, and the test is conducted while both sample and solution are maintained under a slight positive pressure of the test atmosphere. Table I gives the results of contact angle measurements made at various 02-Nz concentrations.5 Each value of contact angle is the average of at least six individual measurements made with different bubbles at different locations on the sample surface. A period of 20 to 25 min has been found by preliminary tests to be sufficient to permit develop- ment of maximum angle, and all measurements are made after at least this much time has elapsed. Exposure of prepared samples to pure water under the test atmosphere before exposure to the xanthate results in a small but measureable and consistent reduction in contact angle. This is particularly true in the oxygen-deficient atmospheres; for example, at 2.2 pct O2 such water contact reduces the average angle about 8" from that obtained by immediate testing in xanthate. The general trend of increasing contact angle with increasing oxygen concentration corresponds to the results of Plaksin et a12 The absence of diminish-ment of contact angle at very high oxygen concentration has been attributed to the high concentration of xanthate per unit of galena surface: At oxygen concentrations below 1 pct (the limit of accuracy of the Orsatt analyzer), no contact angle is obtained. Furthermore, no contact angle is obtained at any oxygen concentration in the absence of xanthate. This result seems to be in disagreement with those cited by Plaksin.1 Further work on the interaction of galena and potassium ethyl xanthate or a similar collector as influenced by oxygen is contemplated. The true nature of the interrelation is as yet unknown. For example, it is not known whether the governing oxygen reaction is with the mineral, with the reagent, or with both.

Jan 1, 1961

By J. Trevor Eyton

This paper addresses some of the financial risk hedging mechanisms available to a corporation contemplating entering the mining business at a significant level of investment. The approach which I have adopted is that of a case study relating to the risk hedging aspects of Brascan Limited's acquisition of 46% of Noranda Inc., formerly Noranda Mines Limited, followed by some remarks on options open to mining companies for reducing financial risks. To appreciate the rationale for Brascan committing over (Can.) $800 million to the natural resource sector, some knowledge of Brascan's parent company, Edper Investments Ltd., is required at the outset. Edper Investments, prior to acquiring a 48% interest in Brascan, controlled Trizec Corporation, one of the world's largest property companies. Edper also was involved in the financial services sector through Continental Bank of Canada and Hees International Corporation. For some considerable time prior to Edper's investment in Brascan, Edper felt a need to provide a better balance to its portfolio of property and financial services. Brascan was rich in cash, relatively unleveraged, and with important existing holdings in natural resources and consumer products. The acquisition of a controlling interest in Brascan provided an opportunity to balance Edper's portfolio of property and financial services with world class acquisitions in natural resources, supplemented by consumer products and financial services. The end result of Brascan's activities over the last few years is illustrated by the following analysis of the book value of Brascan's assets at the end of December 1983. [Can.$% Millions _ Natural Resources81748 Consumer Products 654 38 Financial Services 157 9 Other Operations895$1,717100% ] Source: Brascan 1983 Annual Report. The large emphasis on natural resources came about from the fact that for Canadians operating mainly in Canada, natural resources were attractive. We were impressed by Canada's position in the world natural resource industry as illustrated by some of the following statistics: [CommodityCanadian Estimated Production In 1983 As % of World Total Newsprint 30 Potash 21 Uranium 18 Molybdenum 17 Zinc 17 Pulp 16 Nickel 16 Saw Logs 11 Silver 8 Copper 7 Lead 6 Gas 5 Oil3] Source: U.S. Bureau of Mines and Brascan estimates. In addition, the quality of Canada's reserves looked good to us. The polymetallic nature of many of Canada's orebodies offered a hedge against risk relating to any one metal. Put another way, the by-product metal credits offered a net operating cost advantage with respect to many of the metals mined. Our energy costs also appeared to provide Canada's natural resource industries with an advantage. We considered Canadian mining expertise as amongst the world's finest, as were the exploration equipment and technology developed in Canada. In addition, Canada offered a stable political climate and one we knew. Further, we felt the Canadian Government's future role in the industry would, on balance, be benign, partly because the Canadian content of the mining industry was very high in contrast to the petroleum industry. Granted, we could find cheaper sources of copper; however, to do so would involve venturing into regions where country risks were higher.

Jan 1, 1985

By P. W. Bridgman

FLOW and fracture are admittedly complicated phenomena of which we are yet only partially masters. There is not even universal agreement as to the details of the language best adapted merely to describe the phenomena, to say nothing of any theory. Certain points of view, however, are widely held, and there are certain general expectations as to what sort of procedures will prove permissible. A complete description of any situation involving fracture would demand a complete history of all the variables in the control of the experimenter. The experimenter endeavors so to choose his independent variables that they are adequate to determine completely the situation in the sense that when an experiment is repeated on the same initial material and with the independent variables running through the same sequence, all the other measurable properties of the system will also run through the same course. It is, I think, the universal conviction that when we are dealing with phenomena of flow and fracture, the most important of the necessary variables are the external forces acting on the various parts of the body; when the force history repeats, with certain obvious restrictions such as that temperature must be constant and non- mechanical forces, such as magnetic field, must be small enough to be ineffective, then the history of all measurable properties will repeat; in particular the alteration of geometrical shape due to flow will run through the same course, and when fracture comes, the character of the fracture will be reproduced. Complete mastery of merely the descriptive aspects of flow and fracture would demand the ability to predict the entire course of flow and complete characterization of the final fracture for every possible combination of forces, varying in time in all possible ways, when applied to bodies of all possible shapes. This full situation is obviously of prohibitive complexity. The first stage in reduction to manage- able simplicity is to resolve the total situation into the sum of contributions made by small elements. Our thesis is that if each of the elements is adequately characterized, the behavior of the whole system is fixed. The element is usually taken of a simple geometrical shape, suggested by the mathematics best adapted to the particular situation. The forces acting across the faces of the elements are ultimately determined by the forces acting across the free faces of the whole system. The detailed process of finding the precise forces on the faces of all the elements in terms of the external forces involves the use of a special mathematical invention, the stress, a constructional quantity consisting of six components. The stress at any point in the body is determined by both the external force system and by the physical properties of the body at all points; quite different stresses prevail at the interior of bodies of the same shape under the same external forces in the elastic and plastic ranges. The forces on the faces of any element are, however, uniquely determined by the stress system at the element. Our fundamental thesis

Jan 1, 1944