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By W. F. Campbell

One of the deepest coal operations today is the Springhill No. 2 mine of Cumberland Railway & Coal Co., subsidiary of Dominion Coal Co. Ltd. Mining is now conducted at a slope distance of 14,000 ft, with 4400 vertical ft of cover. The record of Springhill No. 2 can be said to contain the history of bumps in the Province of Nova Scotia. The Springhill coal field forms part of the Cumberland field of Carboniferous age. There are seven mineable fields in the area, numbered in the order they were discovered. Mining has been carried on in all but the No. 4 and 5 seams, but present operations are confined to the No. 2. Fig. 1 shows a vertical section through the seams. Opened in 1873, No. 2 mine was first worked from parallel slopes driven from the outcrop of No. 2 seam down to the 7700 level. As the mine went deeper, a two-place auxiliary slope was driven from the 6900 level to the present workings. A transfer level at the 7800 connects the main haulage with the auxiliary haulage slope. The No. 2 mine plan is shown in Figs. 2 and 3. The seam is bituminous, averaging 8.5 to 9 ft thick. There is a well defined parting 14 to 16 in. from the roof, and this roof coal is harder than the rest of the seam. Average pitch at the outcrop is 30°; at the 6500 level, 20°; at the 7900 level, 16°; and at the 13,800 level, 10°. Immediate roof and floor strata consist of beds of variable thicknesses of shales, grading to arenaceous shales to shaly sandstones to sandstones. A characteristic of the strata is the appearance and disappearance of sandstone bands, of considerable thickness, over distances of several hundred feet. MINE HISTORY Entrance to No. 2 mine is obtained by three parallel slopes, separated by 100-ft pillars except in the upper portion of the mine where the pillars are smaller. Main haulage levels were originally driven 600 ft apart for room and pillar extraction. A parallel drainage and intake airway was driven below each haulage level and a parallel return airway (counter level) above. Pillars 80 to 100 ft wide separated these two servicing entries from the center haulage level. Inclines up to 700 ft apart were driven off the levels up pitch to the upper levels, and at 40-ft centers level rooms 12 ft wide were driven off each side of the inclines, separated by crosscuts every 50 to 100 ft. The rooms were driven up to 350 ft or until they were holed into the room from the adjoining incline—a plan followed until the level reached its boundary. When the entire area between two levels had been divided into pillars, the pillars were extracted from the boundary back to the slope pillars. As depth of cover increased, mining conditions required larger pillars, and at the 3300 level, where NECESSARY PRECAUTIONS FOR MINING AT DEPTH In Room and Pillar Extraction: 1) Pillars must not be too small. 2) Pillar size must be uniform. 3) Extraction lines must be kept as straight and as uniform as possible. If they are irregular they should be brought back into line slowly. 4) Peninsulas of coal surrounded by gob must be avoided. 5) Pillars should not be disturbed after they have been formed. If it is necessary to increase the width of an opening, stripping or ribbing should be done gradually over a long distance. If increased width of openings is necessary at certain locations and can be anticipated, the openings should be widened during development. 6) Under no conditions should small pillars be left in the gob that would interfere with caving. In Long wall Operations: 1) Pillars formed during development for retreating longwalls should be uniform in size and shape. 2) Pillars should not be split after development, but if pillar splitting does become necessary it should not be carried out within the zone influenced by the working faces. 3) Width of levels should not be increased after the levels have been developed. If it does become necessary to widen a level it must be done gradually over a long distance by stripping or ribbing the coal ribs. 4) Extraction lines should remain uniform. If several walls are worked together—one behind the other—a uniform distance must be maintained between them and the wall faces must be kept as straight as possible. If the walls do get out of line, they must be brought back into line gradually to allow for a gradual change in the stress distribution over the area. 5) Stumps of coal, roof supports, etc. must not be allowed to remain in the gob to interfere with caving. 6) Midwalls should be built as rigidly as possible and properly maintained.

Jan 1, 1959

By George C. Branner

STATE geological surveys have had an interesting development in this country. They first appeared more than a hundred years ago. The fact that they have persisted and are now an important part of most state governments indicates the basic need for state departments con-cerned with mineral problems. The number of state surveys in 1941 (43) is at an all-time high, as is the aggre¬gate appropriation for their support for the fiscal year 1940-'41 ($1,600,708). The first surveys, those of North and South Carolina (see Fig. 1), made their appearance in 1820-'30, but neither survived that period. During the next two decades surveys were created in twenty states, but only three survived. In fact, it was not until the decade 1900-'10, when eighteen surveys were created and only five abolished, that these agencies became firmly established.

Jan 1, 1941

I HAVE hitherto shown you as well as I knew how the procedure for those operations which are needed in melting on either a large or small scale, as you might need them. Now I wish to pass on to some other operations which, while related to the operations and power of fires and also to certain kinds of melting, are not wholly connected with the art of casting. First I wish to begin with the actual substance of the art of alchemy, which is sought and desired with so much zeal. Then, as a useful and ingenious thing, I wish to inform you of the art of distilling, and, afterwards, how you should operate a mint in an orderly fashion and without burden on the conscience. Then I shall tell you of the ironworker, the goldsmith, the coppersmith, and some others, as you will be able to see in their places.

Jan 1, 1942

By S. R. Seagle, R. Schuhmann, N. A. Parlee

Rates of CO evolution and CO absorption were measured for liquid-iron alloys containing from 0.15 to 4.4 pet C, using a modified Sieverts apparatus. The alloys were held in alumina crucibles, so that both crucible-metal and gas-metal reactions occurred simultaneously. The data are interpreted on the hypothesis that the C-O reaction rate is controlled by oxygen diffusion across the boundary layers at the gas-metal and crucible-metal surfaces. CARBON-monoxide evolution from liquid Fe-C-0 alloys is a key reaction in steelmaking processes, both in the steelmaking furnace and in the ingot mold. Also, this reaction probably is an essential step in the desulfurization of iron under blast-furnace conditions, for which carbon is the principal reducing agent or deoxidizer present in the iron. The thermodynamic properties of the solutions of C and 0 in Fe and the equilibria of these solutions with gaseous CO and CO, have been investigated in some detail and are reasonably well understood. However, while these equilibrium data have accumulated in the laboratory, other data have accumulated to show that equilibrium conditions are often not achieved under plant operating conditions. Thus, an understanding of the rate and mechanism of the carbon-oxygen reaction not only has theoretical interest, but ultimately may assume considerable practical importance. Reaction Mechanisms and Rate Equations The over-all reaction under consideration is C (in liquid Fe) + 0 (in liquid Fe) + CO (gas) [I] Although this appears to be a straightforward heterogeneous reaction, involving just two phases (liquid metal and gas), a wide variety of reaction mechanisms has been proposed. Different authors have expressed divergent views as to the nature of the rate-controlling steps. Accordingly, a brief discussion of the possible reaction steps will be given. 1) Homogeneous Reaction Within Liquid Iron— Feild' and Jette2 have applied chemical reaction-rate theory as if the C-0 reaction were a homogeneous, second-order reaction occurring within the liquid-metal phase. This mechanism necessarily produces CO "molecules" dissolved in liquid iron. In order to yield CO gas, the dissolved CO must nucleate CO gas bubbles, and the dissolved CO also must be transported to the bubble surface where it enters the gas phase. However, this mechanism is inconsistent with present concepts of the structure of liquid-metal solutions. 2) Reaction at Gas-Metal Interface—At the surface of contact between gas and metal, a second-order reaction between dissolved C and 0 may occur to produce a CO molecule which enters the gas phase. This reaction presumably involves an "activated complex" structure in the interface. Some such surface mechanism appears essential because the carbon and oxygen do not occur in the same species in the two phases and therefore must react in some way at the surface. At liquid-iron temperatures, calculations based on reaction-rate theory indicate that the surface reaction must proceed extremely rapidly and thus cannot be a rate-determining step in the C-O reaction under ordinary conditions.3 ccordingly when the over-all reaction is measurably slow, the surface reaction may be considered at equilibrium; that is Cc* Co* = m'Pco [2] in which Co* and Co* are the concentrations of carbon and oxygen (weight per unit volume of metal), respectively, in the liquid-metal phase at the interface, Poo* is partial pressure of CO in the gas phase at the surface (atmospheres), and m' is the mass-action constant for Reaction [I]. Since activities of C and 0 are not strictly proportional to concentrations, m' is not a true thermodynamic equilibrium constant but varies with carbon content.' 3) Mass Transport—If the reaction is to proceed at the gas-metal surface, as described above, dissolved carbon and oxygen in the liquid iron must come to the surface and gaseous carbon monoxide must move away from the surface. Details of the mass-transport mechanisms depend on the kind of system under consideration. For a stirred metal bath in contact with a CO gas phase, the transport of C and 0 to the surface can be described by 1 Do Mols CO evolved per sec = — 1/16— Do/do Agm (CO-Co*) = 1/12 Do/do Agm *(Co-Co*) [3] In this equation, Do and D, are the diffusion constants for oxygen and carbon in liquid iron, 6, and 6,. are the boundary-layer thicknesses, A,,,, is the

Jan 1, 1959

By Albert J. Phillips

IT is a privilege and a pleasure to deliver this, the Twenty-sixth Annual Institute of Metals Division Lecture. Eleven years ago C. A. Edwards addressed this audience on the subject "Gases in Metals." In spite of this relatively recent presentation, I propose to, present to you some quantitative considerations of industrial significance on the separation of gases from molten metals because recent observations and measurements, presented in the literature, have made it possible to evaluate mathematically the gases or gas-forming elements and to develop an understanding of the mechanism of gas separation that should permit, in many cases, the control of this baffling phenomenon. DEFINITIONS Before becoming involved in the subject it seems desirable, for the sake of clarity, to state certain definitions and axioms: I. We are dealing only with gases that separate from the molten metal either during cooling or in the transition from liquid to solid. Gases injected by the liquid stream and trapped in the solidified metal, or those forced into the solidifying metal from the mold due to either thermal or cher4ical effects are not considered. Neither are we concerned with the separation of gases within the metal after solidification is complete for the data on this subject are at least complete enough to permit the practicing metallurgist to satisfactorily control the phenomenon. 2. Only where there is a decrease in solubility or a change in equilibrium of either a gaseous element or an element that reacts to form a gaseous compound, can there be a separation of a gaseous phase from the liquid or solidifying metal. 3. The amount of gas separating from the liquid during solidification can be computed directly from the amount of the element contained in the liquid when freezing commences less the amount of the element found in the solid either as a separate solid phase or in solid solution. 4. There can be no separation of gas bubbles unless the told pressure of the gases in the liquid meld exceeds atmospheric pressure plus the hydrostatic pressure of ' the liquid metal head. (In addition, pres- sure resulting from surface tension in- creases the total pressure requirement for initial gas separation. Probably the remarkable differences in appearance caused by different gases in the same host metal are influenced greatly by the rate of diffusion of the gas in the liquid metal balancing the decrease in pressure as the gas bubble size increases and the surface tension effect decreases. Thus hydrogen with its high diffusion rate in liquid copper separates in the form of relatively coarse cavities for a given rate of heat extraction.) 5. Gas separation can take place at less than atmospheric pressure in the voids created by the shrinkage from liquid to solid, but since the voids are not caused by the gas the presence of gas in the voids is of importance only where it

Jan 1, 1947

By R. K. Adams

The availability of digital computers costing less than $15, 000 makes them economically attractive as versatile research and experiment tools. Several such installations at Oak Ridge National Laboratory have shown that improvements in research productivity and experimental control accuracy can result from the proper application of a small digital computer. However, careful attention must be paid to several potentially expensive problems in 1) interfacing the experiment to the computer and 2) programming the computer. Two specific applications of very small computers to metallurgical testing and research are described. One is a system for data acquisition and precise temperature control associated with a thermocouple testing program, and the other is a system for pulse calorimetry determination of the specific heat of metals over wide temperature ranges and in the neighborhood of metallurgical transformations. The data acquisition system which uses a PDP-8/S computer monitors signals from up to 264 low-level thermocouple signals at two scanning speeds: 10 readings per second for long-term drift data, and 500 readings per second for transient data. The pulse calorimetry system which uses a PDP-8 computer accepts up to 6 signals of 0- to 10-v range and because of a larger analysis program, uses a bulk memory (magnetic disc) in addition to core memory.

Jan 1, 1969

AS you have seen, I have hitherto demonstrated to you the methods for bringing to their proper and pure bodies all metals of any kind of ore that you have found and mined. If these did not now serve human needs all these operations with so much expense and labor would have been profitless and in vain. Therefore I recognize that it is necessary in speaking to touch upon all their operations and to teach you to ma& use of the metals according to the works needed. There are, for instance, some metals that would serve poorly for certain works when reduced to their fineness and perfection. Also, many artificers, in order to avoid expense, are content with having a larger quantity of metals and do not care so much about their perfection. For this reason they willingly accept mixtures of one metal with another, for example, of silver or copper with gold, of copper with silver, and likewise of tin or lead with copper. Tin is accompanied by lead to avoid spoiling its nature completely. All this, however, is done with a definite proportion by weight and not by chance. These admixtures, as I have told you in another place, are called in gold, carats, and in silver, leghe. This, in fact, means nothing but the association of one metal with another, indeed, a soiling of their purity and fineness, to which they have been brought with so much hard effort and labor. But now leaving this discussion, I shall tell you in the present book as briefly as I can, in order, first of the alloying of gold with silver or copper, and then of that of silver with copper.

Jan 1, 1942

By H. C. George

IN 1901, the United States produced 69 million barrels of crude oil, which was 41.4 per cent of the world production. By 1931, these figures were 850 million barrels and 62.1 per cent respectively, showing over this period of thirty years an increase in the world production of 718 per cent. Over this same period, the increase in the demand for technically trained men in the industry was much greater than the increase in crude-oil production. This was because much of the increase in production was secured by the application of geology and physics to the finding of new oil fields and by .the development of engineering methods and equipment incident to drilling and producing from much greater .depths.

Jan 1, 1934

By S. Harries Daddow

IN order to get an idea as to the strength of steel rails, it will be well to review the tests to which iron rails have been subjected. In England, Mr. Ashcroft found that the best 80 pound rails broke under a 300 pound weight, falling 15 feet. In Germany the Society of Railway tyfanage1-s determined 011 and have long applied a test of 1000 pounds falling 101/2 feet, as the standard which all first-class iron rails must reach. In this country no inspection nor test is applied, but tests made show that iron rails fro; our most reliable makers, break under a G foot fall of a 1500 pound drop as an extreme test, most of those tested breaking under a far less test; some breaking with less than a 3 foot fall of the same weight. Everywhere where steel has been used, engineers have come to the conclusion that some test is required to show the regularity and strength of the product. As compared with iron, the tests which steel will stand are wonderful. After numerous experiments partially based on the experience of the rail-mill at Graz, belonging to the Southern Railway of Austria, the Society of German Railway Managers fixed upon a test of 2000 pounds falling 138 feet. They found that this test represented the steel which suited their necessities, and also found that with steel of otherwise average purity, this test represented about one-half per cent. of carbon, and made it a rule to take no steel containing under three-tenths of a per cent. of carbon, because it was too soft. They expressed a hope that a harder steel could soon be made tough enough to stand the same test. I n England, a test was adopted of 2240 pounds, falling 15 to 17 feet on the rail on heavy bearings. This test has been found satisfactory under heavy traffic on average road-beds, and has been invariablyPILLAR of coal, which are designed and left in our mines as the means of supporting the overlying strata of our coal-beds, and securing protection to the miner, are not only insufficient for these purposes, but the most objectionable means of providing the desired security to life and property. Though such pillars are left at a great sacrifice to the owners of collieries and mineral lands, loss to the resources of the commonwealth, and ultimate privation to the public, they defeat the very ends for which they are designed, and not only do not secure life and property, but are as fatal to the one as ruinous to the other. Propositions.-It is eminently desirable that the resources of the Common wealth-particularly our mineral wealth in anthracite coal, which is limited-should be carefully utilized and made available, not only to ourselves but our posterity. Yet, speaking as an anthracite miner, we are wasting our resources of coal with a recklessness which will bring ruin in the end, unless checked, not only to the iron manufacturing interests of the East, whose chief protection, in competition with the iron-masters of the interior, lies in a cheap and abundant supply of anthracite furnace fuel, but to the best interests of the State. It is, consequently, not only desirable, but necessary, that we should realize every ton of coal from each colliery, and every acre of coal area, in order that our resources may not be wasted at present, to breed want in the future, and to realize the greatest present benefit to the mine, the miner, and the public. It is not possible, however, to secure these ends, if we continue to leave, as we do at present, from one-third to one-half our resources in anthracite as pillars in our mines. But our sad bills of mortality, our abandoned collieries, and our yawning mountain-sides, present ample evidence that even this enormous waste of coal, though left for the purpose of protecting life and property, is as totally inade-

By O. W. Ellis

Before entering upon a general discussion of the fascinating, but at present rather controversial, subject of gases in copper, the author feels that some attention should be directed to the work which has been carried out on the solubility of gases in metals. The outstanding contributions to our knowledge of this subject have been made by Sieverts and his collaborators(1) † and by Iwasé.(2) Both the German and the Japanese investigators employed apparatus of the same general design in their investigations. Full reference need only be made, therefore, to the German scheme, which is diagrammati-cally illustrated in Fig. 1. This, to use the words of Sieverts, which he employed in describing his arrangement of apparatus at the 1928 meeting of der Deutschen Gesellschaft für Metallkunde(10): shows a diagram of the usual mode of procedure. A quartz flask contains the metal (for high temperatures a highly refractory porcelain tube is used) and is connected with the rest of the apparatus by means of a stopcock. The other parts consist of a burette divided into 1/10 c.c., a leveling tube, and an adjustable mercury reservoir. The burette is connected on one side with the gas inlet and on the other side with the absorption flask; the connection can be interrupted by means of suitable taps. Another tap permits the connection of the flask and the Töpler pump. In an experiment the quartz flask is first exhausted and then a measured quantity of gas is let in from the burette. In the first place the volume of the flask when empty is found by calibration with an inert gas. The pressure is read off on the level tube of the burette, and the temperature measured by means of a thermocouple whose welded junction lies close to the outside of the quartz flask, which is heated in an electric furnace. In several experiments the quantity of absorbed gas was transferred to a burette by means of a pump and measured. Care was taken by Sieverts to cool the ground-in joint between the reaction flask (or tube) and the rest of the system. Suitable precautions

Jan 1, 1934

By E. A. Gulbransen, K. F. Andrew

THIS paper. will present the results of our studies on the kinetics of the gas phase reactions of co-lumbium and tantalum with O2, N2 and H2. Studies on zirconium and titanium have been previously reported.9,11 Due to their high melting points and their many excellent physical and chemical properties, colurn-bium and tantalum have been finding their places as useful metals in science and industry. Chemically the metals are very inert to gas and liquid phase corrosion at room temperature. However, at moderate temperatures of 250°C and higher the metals react readily with oxygen and hydrogen. It is of interest to study the gas phase reactions of these metals for the following reasons: (1) to compare the rates of the several reactions with those of other metals and to correlate the rate data with theory and structure predictions, (2) to add basic information on the practical difficulties of the reduction, refining and working of columbium and tantalum, (3) to anticipate new uses for the metals and (4) to understand the role that the metals play as components in high temperature alloys. Literature Survey A number of papers exist in the literature on the gas phase reactions of columbium and tantalum with O2, N2 and H2; however, only fragmentary data have been reported on the kinetics of these reactions. Although the free energy of formation of the oxide Ta2O520at 25°C has been determined and some thermodynamic work has been reported on columbium and its oxides, the data are not sufficient to make equilibria calculations at elevated temperatures. Oxidation: 1. Kinetics: The rate of oxidation of columbium and tantalum has been studied by McAdam and Geil23 using the interference color method. Tantalum is reported to oxidize more slowly than columbium, zirconium and iron. The amounts of oxidation of columbium after 20 hr in air for several temperatures are reported by the Fansteel Metallurgical Corporation.36 Columbium starts to oxidize in air at about 200°C. The oxide is said to be adherent and prevents further oxidation until the temperature is raised. Columbium is reported not to become brittle on heating in air for short periods like tantalum because the oxide film prevents further reaction. The columbium oxides dissolve into the metal when heated in a vacuum at temperatures between red heat and 1200°C. Above 1200°C the oxides are reported to evaporate. Columbium, in the oxide free state, is an active getter in vacuum tubes. 2. Preparation and Structure: G. Grube, O. Kuba-schewski and K. Zwiauer15 have studied the decomposition of Cb2O5 in argon. They find that CbO2 is formed slowly at 1150° and rapidly at 1300°C. Cb2O5 can be reduced to CbO2 by moist H2. The reduction of this dioxide to the lower oxides is rapid. By varying conditions CbO2, Cb2O3, CbO and Cb2O can be formed. CbO2, CbO and Cb2O have independent lattices while Cb2O3 is a stoichiometric mixture of CbO and CbO2. In a later paper O. Kuba-schewski" has shown that CbO is cubic and that CbO2 has a rutile type of lattice. The most exhaustive work on the oxides of columbium is that of G. Brauer.8 The number and constitution of the solid phases in the binary system Cb-O were investigated by the preparative, analytical and X ray methods. He indicates that only the oxides Cb2O5, CbO2 and CbO exist with limited regions of homogeneity. The solid oxide Cb,O, exists

Jan 1, 1951

By T. C. Campbell

Copper is extracted from the Chuquicamata oxide ore by a hydro-metallurgical process. The ore is crushed to 3/8-in. sizing, and leached with a sulfuric acid electrolyte. Chlorine is precipitated. and the ferric iron reduced in the enriched electrolyte, after which the copper is recovered by electrolysis with insoluble Chilex and lead-antimony anodes, the spent electrolyte being returned to leaching. Cathodes are melted and refined in market furnaces and cast into commercial wire-bar and cake shapes. Sulfuric acid for the process is supplied by the brochantite in the ore. Water for washing the ore is advanced through the solution system and finally after cutting by electrolysis to from 6 to 16.0 grams per liter is completely stripped by the cuprous chloride method and run to waste. The cuprous chloride so obtained plus that resulting from precipitation of chlorine from strong solution is dissolved in ferrous chloride brine and the copper cemented on scrap iron. The cement so obtained is in part used to reduce ferric iron in electrolyte, and as it is of exceptional purity is also furnace refined and cast into an exceptional quality of fire-refined copper. Fig. 1 is flow sheet of the operation. The reduction plant is composed of seven divisions as follows: (1) crushing, (2) leaching, (3) tailings disposal, (4) dechloridizing, (5) sulfur dioxide treatment, (6) electrolytic tank house, (7) smelting and melting. A brief description of each of these divisions will be given. Crushing Ore that is frequently blocky and up to 5 ft. in section is crushed in four stages to 3/8-in. mesh. The first stage of crushing is called primary and the next three stages are known as secondary crushing. The first stage is accomplished with two 60-in. Superior McCully gyratory breakers by which the ore is broken to 9 in. The second coarse gyratory breakers crush the ore to 3-in. sizing. There are seven No. 10 McCully gyratories in this service, for one of which recently has been substituted a No. 7 cone crusher. The third and fourth stages of crushing are accomplished by fifty 48-in. disk crushers. In the third stage 14 disk crushers break the ore to 1 in., and in the last stage 36 disk crushers turn out the final product of 0.371-in. sizing. Reference to the flow sheets of Figs. 2 and 3

Jan 1, 1934

By Fred P. Peters

At first glance this paper may seem unique among those comprising this symposium on designing for powder metallurgy, since it is evidently concerned with everything but Powder metallurgy. This paradox mill be at once resolved, however, when the real Purpose of the symposium meeting is understood—to present factually and objectively the things that can and cannot he done with powder-metallurgy methods, and especially to set forth with brutal frankness the limitations of Powder metallurgy, so that costly over-enthusiasm and misapplication may be prevented. Since much of powder metallurgy's future development (and many of its present disappointments) has been in competition with other and long-established methods of fabricating parts, it is easy to see that its true competitive possibilities in any given case cannot be accurately appraised without rudimentary knowledge of what the competition can do. This paper, therefore, presents a few reminders of things you all know—a background to be kept in mind as you examine the data on each of their specialties given by the authors of the papers that follow. From the standpoint of the design engineer powder metallurgy is often in Competition with one or more of the following: sand casting, die casting, permanent-mold casting, precision investment casting, ma- chined bar stock, screw-machine parts, cold-heading, drop-forging, and stamping and drawing. We will here briefly discuss each of these individually, and set forth so, rough (and therefore dangerous) generalizations about its competitive position, especially with respect to powder metallurgy. Sand Casting Sand-casting of small parts is sometimes an inexpensive method of manufacture, for its raw-material costs (i,e., for gray iron) are the lowest of the various methods and it involves the use of no tools and dies of jewelry quality. its design flexibility— that is, the ability to make complicated, cored and undercut parts in one piece—is about the highest of the competing processes and in this respect it can do many things that are impossible by powder metallurgy. For small castings its production rates may be higher, by using multi-impression molds, than can be achieved with parts of the same size by powder metallurgy. But its labor charges for melting, pouring, patternmaking and molding are comparatively high and the waste it entails in gates and risers is always considerable. Tolerances of sand castings are wide (+1/6 to 1/32 in- per in.) and where precision approaching that of powder metallurgy is required sand castings must be machined to final size, thus increasing their cost. Steel sand castings are more expensive and less precise than iron, and nonferrous the most costly of all.

Jan 1, 1945

By W. W. Wirth

RESEARCH looking toward the reduction of the cost of roof support by substitution of longer-life materials for wooden timber is fully justified by the fact that roof support is an important element of all production costs. Permanent or semipermanent roof support will reduce the expense of timber-maintenance crews during steady operations, but even more markedly during idle periods. The management, therefore, is faced with the important problem of timber renewals, which in times of regular operation are frequently obscured by more urgent problems of production. Increased use of timber in the Anthracite Region of Pennsylvania is caused mainly by increased percentage of coal obtained from robbing areas, which requires more timber for safety than develop-ment work, and maintenance of long haulageways through crushed ground despite the fact that less coal is transported over them. The timber reserve in the immediate vicinity of the mines has been depleted to such an extent that it is necessary to secure 70 per cent of the timber requirements from distant sources. Originally, mine timbering in this region was done by largely utilizing round props or legs with round collars. The whole cross section of the tree from which timber was cut was utilized by forming the joints between the legs and the collars with saw and adze, using plank or round saplings for lagging behind the legs and over the collars to prevent material falling between them. The Mine Inspectors' report of 1898 records an important digression from the use of round timber, when two tunnels at the Babylon colliery, near Pittston, Pa., were timbered with 12 by 12-in. square timber sets with mud sills, through quicksand for distances of 425 and 160 ft., respectively. The timbers were lagged with 3-in. plank to control sand.

Jan 1, 1940

By Harry H. Angst

The history and efficiency of 40-ton capacity dump trucks for surface waste removal at the New Cornelia opencut copper mine, at Ajo, Ariz., are summed up in this paper. Tabulations of truck performance, together with detailed cost figures, prove the value of truck transportation for certain special operations at a large mine. Emphasis is given to the fact that truck transportation not only reduces the haulage costs but also reduces the breaking ground and shovel-loading costs per ton. The New Cornelia opencut copper mine is in the desert. The climate is ideal for seven months of the year, but the five summer months are uncomfortably hot. 'She elevation is 1820 ft. above sea level and the annual rainfall is approximately 6 inches. The ore body is a monzonite-porphyry deposit. The waste-rock capping consists of hard rhyolite, conglomerate and porphyry. All the material is hard and cannot be excavated without blasting. The rhyolite and conglomerate waste is very blocky. The weight of the rock in place is 2.16 tons per cubic yard. The average mine production per day is 22,090 tons of 012 and 18,000 tons Of waste rock, the latter consisting of surface rock and material within the ore body too low in grade to send to the concentrator. The rock is loaded by electric shovels with 4½-cu. yd. dippers. The primary transportation of ore to the concentrator and waste to the dumps is by 30-yd. dump cars and oil-burning steam locomotives; therefore truck haulage of waste is a minor factor of the New Cornelia operation. However, as New Cornelia was one of the first mines to use large-capacity dump trucks for haulage, the results obtained are of interest. Equipment On the south side of the New Cornelia is a hill rising about 500 ft. above the normal surface elevation, roughly 1000 ft. in diameter at its base, called locally Arkansas Mountain peak. Exploratory drilling in this area developed additional ore that could be made accessible for open pit operation only by moving the mountain face a maximum distance of 400 ft. to the south. The slope of the mountain adjacent to the pit averaged 35°, and the small area involved precluded the use of train haulage for the removal of the 4,000,000 tons of waste estimated above the normal pit elevation. A study of dump-truck haulage at Boulder Dam and San Gabriel Dam led to the decision to use large-capacity end-diii~lp trucks rather than smaller trucks, for the following reasons: I. The truck would have to withstand the impact of hard, heavy chunks, some of them weighing as much as 10 tons, which would be dumped from the 4½-cu. yd. dipper. 2. The coarse, rocky, abrasive material would soon damage a light truck body. 3. The body of a small truck could not be efficiently loaded with coarse rock.

Jan 1, 1943

By Harry H. Angst

The history and efficiency of 40-ton capacity dump trucks for surface waste removal at the New Cornelia opencut copper mine, at Ajo, Ariz., are summed up in this paper. Tabulations of truck performance, together with detailed cost figures, prove the value of truck transportation for certain special operations at a large mine. Emphasis is given to the fact that truck transportation not only reduces the haulage costs but also reduces the breaking ground and shovel-loading costs per ton. The New Cornelia opencut copper mine is in the desert. The climate is ideal for seven months of the year, but the five summer months are uncomfortably hot. 'She elevation is 1820 ft. above sea level and the annual rainfall is approximately 6 inches. The ore body is a monzonite-porphyry deposit. The waste-rock capping consists of hard rhyolite, conglomerate and porphyry. All the material is hard and cannot be excavated without blasting. The rhyolite and conglomerate waste is very blocky. The weight of the rock in place is 2.16 tons per cubic yard. The average mine production per day is 22,090 tons of 012 and 18,000 tons Of waste rock, the latter consisting of surface rock and material within the ore body too low in grade to send to the concentrator. The rock is loaded by electric shovels with 4½-cu. yd. dippers. The primary transportation of ore to the concentrator and waste to the dumps is by 30-yd. dump cars and oil-burning steam locomotives; therefore truck haulage of waste is a minor factor of the New Cornelia operation. However, as New Cornelia was one of the first mines to use large-capacity dump trucks for haulage, the results obtained are of interest. Equipment On the south side of the New Cornelia is a hill rising about 500 ft. above the normal surface elevation, roughly 1000 ft. in diameter at its base, called locally Arkansas Mountain peak. Exploratory drilling in this area developed additional ore that could be made accessible for open pit operation only by moving the mountain face a maximum distance of 400 ft. to the south. The slope of the mountain adjacent to the pit averaged 35°, and the small area involved precluded the use of train haulage for the removal of the 4,000,000 tons of waste estimated above the normal pit elevation. A study of dump-truck haulage at Boulder Dam and San Gabriel Dam led to the decision to use large-capacity end-diii~lp trucks rather than smaller trucks, for the following reasons: I. The truck would have to withstand the impact of hard, heavy chunks, some of them weighing as much as 10 tons, which would be dumped from the 4½-cu. yd. dipper. 2. The coarse, rocky, abrasive material would soon damage a light truck body. 3. The body of a small truck could not be efficiently loaded with coarse rock.

Jan 1, 1943

By R. W. Gurry

IN the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and 1110°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon content of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched. The mean result of concordant measurements is that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C. (1755°F.) and 1.89 per cent at 1110°C. (2030°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutectoid temperature (738°C. or 1360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature 1135°C. (2075°F.) instead of the commonly accepted value of about 1.7 per cent. EXPERIMENTAL PROCEDURE The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.1 In the empty furnace the range of temperature over a distance of 4 cm. at the heat center was 9°C. (16°F.) but at a given location the temperature remained within ±0.5°C. (0.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (1760°F.) and 1110°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with

Jan 1, 1942

By James J. Halbouty

The restraining effects of war conditions on oil-field development are well illustrated by statistics on South Texas † fields for the year 1942. Statistics also show a decline in wildcatting, but this decrease is much smaller than that shown by field activity. In comparison with results in 1941, there was a reduction of about 44 per cent in the total number of wells drilled in the area (802 in 1942 against 1452 in 1941) This 44 per cent drop represents declines of 50.4 per cent in field wells and of 17.4 per cent in rank wildcats. Field wells (exclusive of 18 wildcats that discovered fields) totaled 598; also, 204 rank wildcats were drilled (see Tables 3 and 4). The corresponding figures for 1941 are 1205 field and extension wells and 247 rank wildcats. In this connection, it should be noted that whereas in 1941 the rank wildcats made up only 17 per cent of all wells drilled in the area, their percentage of the 1942 total was 25.4 per cent. In spite of the decrease of 43 in the number of rank wildcatst drilled, there were 18 new productive areas discavered, one more than the 17 new fields found in 1941. Thus the field-discovery rate for 1942, one new field for every 11.3 wildcats drilled, compares favorably with the I to 14.5 rate of 1941. Thirteen new producing levels were opened in fields; this includes the finding of 44' gravity crude on the west flank of the Sejita field in a 5700-ft. sand, which produccs distillate higher on the structure. Interesting, and showing the general scarcity of sand discoveries, is the fact that of the 13 new sands discovered, 7 were found in Jim Wells County. The year's sand-discovery rate of one new level for every 46 field wells compares very unfavorably with the I to 19.7 rate of 1941, during which 61 new producing horizons were found. Developmeilt of new fields was on a small scale. In several instances activity in these areas was not carried beyond the drilling of the discovery well. The field discoveries of 1942, together with the type of production obtained from the discovery wells and the total number of wells (productive and dry) drilled in them, are given below: Discovery Num- Field.I County Well her Produced Wells I. Cadena....... Duval Oil 2 a. Garza......... Duval Oil I 3. Dinn......... Duval Gas I 4. Keller......... Duval Gas I 5. East.......... Tim 0 11 I 6. Green Branch . McMullen Gas 3 7. Frost. .GreenBranch........ Starr Oil I 8. Garcia........ Starr Oil 6 9. Ulrich........ Starr Oil 4 10. Lykes......... Webb Oil 2 11. McLean....... Webb Oil 9 Iz. Alamo........ Hidalgo Distillate 2 and gas 13. Penitas....... Hidalgo Distillate 2 and gas 14. Xmpnan..... Jim wells i a 15. South Premont Jim Wells Distillate i and gas 16. Quinto Creek.. Jim Wells Gas 4 17. Roth.. . :..... Jim Wells Gas I I 18. South Sinton.. San Distillate I Patricio and gas . • See Table 2 for data on discovery wells and Table I for information on these fields. • See Table 4.

Jan 1, 1943

By James J. Halbouty

The restraining effects of war conditions on oil-field development are well illustrated by statistics on South Texas † fields for the year 1942. Statistics also show a decline in wildcatting, but this decrease is much smaller than that shown by field activity. In comparison with results in 1941, there was a reduction of about 44 per cent in the total number of wells drilled in the area (802 in 1942 against 1452 in 1941) This 44 per cent drop represents declines of 50.4 per cent in field wells and of 17.4 per cent in rank wildcats. Field wells (exclusive of 18 wildcats that discovered fields) totaled 598; also, 204 rank wildcats were drilled (see Tables 3 and 4). The corresponding figures for 1941 are 1205 field and extension wells and 247 rank wildcats. In this connection, it should be noted that whereas in 1941 the rank wildcats made up only 17 per cent of all wells drilled in the area, their percentage of the 1942 total was 25.4 per cent. In spite of the decrease of 43 in the number of rank wildcatst drilled, there were 18 new productive areas discavered, one more than the 17 new fields found in 1941. Thus the field-discovery rate for 1942, one new field for every 11.3 wildcats drilled, compares favorably with the I to 14.5 rate of 1941. Thirteen new producing levels were opened in fields; this includes the finding of 44' gravity crude on the west flank of the Sejita field in a 5700-ft. sand, which produccs distillate higher on the structure. Interesting, and showing the general scarcity of sand discoveries, is the fact that of the 13 new sands discovered, 7 were found in Jim Wells County. The year's sand-discovery rate of one new level for every 46 field wells compares very unfavorably with the I to 19.7 rate of 1941, during which 61 new producing horizons were found. Developmeilt of new fields was on a small scale. In several instances activity in these areas was not carried beyond the drilling of the discovery well. The field discoveries of 1942, together with the type of production obtained from the discovery wells and the total number of wells (productive and dry) drilled in them, are given below: Discovery Num- Field.I County Well her Produced Wells I. Cadena....... Duval Oil 2 a. Garza......... Duval Oil I 3. Dinn......... Duval Gas I 4. Keller......... Duval Gas I 5. East.......... Tim 0 11 I 6. Green Branch . McMullen Gas 3 7. Frost. .GreenBranch........ Starr Oil I 8. Garcia........ Starr Oil 6 9. Ulrich........ Starr Oil 4 10. Lykes......... Webb Oil 2 11. McLean....... Webb Oil 9 Iz. Alamo........ Hidalgo Distillate 2 and gas 13. Penitas....... Hidalgo Distillate 2 and gas 14. Xmpnan..... Jim wells i a 15. South Premont Jim Wells Distillate i and gas 16. Quinto Creek.. Jim Wells Gas 4 17. Roth.. . :..... Jim Wells Gas I I 18. South Sinton.. San Distillate I Patricio and gas . • See Table 2 for data on discovery wells and Table I for information on these fields. • See Table 4.

Jan 1, 1943

By R. Floyd Farris

A method is presented for determining minimum waiting-on-cement time, which takes into account the differences that exist between types and brands of cements and such individual well conditions as depth, temperature, and pressure. The basis for the method was determined by laboratory tests. Being a laboratory development, several steps were required to prove its merit. The first step consisted of laboratory tests designed to determine the minimum cement strength required in wells. Basis was found for setting a minimum value of 8 Ib. per sq. in. tensile strength. Next, it was shown by laboratory tests that the time to 8 Ib. per sq. in. tensile strength may be expressed as a function of consistometer stirring time to loo "poises," the approximate relation being "the time to 8 Ib. per sq. in. tensile strength equals the time to I00 'poises' times three." Next, it was shown that the time of maximum temperature development in cement slurries, due to heat of hydration, is also related to consistometer stirring time to loo poises, but only by a factor of approximately two. It was shown also that the shut-in casing pressure will build up after cement is placed and register a maximum pressure at approximately the same time the slurry down the hole attains maximum temperature. From this and the relationships listed above, the general rule was established that minimum waiting-on-cement time (time to 8 Ib. per sq. in.) after casing cement jobs in any well is equal to the time when the shut-in casing pressure reaches a maximum, as measured from the initial mixing of cement, times a factor of 1.5. Cement plugs drilled in the field at the time prescribed by this formula were found to drill "firm to hard," thus confirming the laboratory tests. These tests prove that many of the present regulations for waiting on cement require a longer time than is absolutely necessary. Use of the method herein proposed offers the possibility of a saving of $1200 per well. Introduction The length of time allowed for cement to set after casing is determined either by state-wide rules, field rules, or self-imposed rules written into drilling contracts. In general, the time is dictated by experience and common practice. However, owing to differences in opinion and in experience of the various groups involved, waiting-on-cement time often varies from one area to the next. For example, an operator in an area where no rules exist may drill out of surface pipe at 24 to 36 hr., while another operator in another area may wait 48 hr. or more to comply with state or field rules, although the depth of the well, hole size, type of cement, and other data are identical. An even greater difference in practices will be found by making similar comparisons with respect to oil-string cement jobs. Differences in waiting-on-cement times of 36 to 48 hr. are common. Further complicating the picture is the rather common practice of allowing more waiting time for cement to set at the greater depths than is allowed at the shallow depths. This practice has existed for years in spite of the common knowl-edge1,2,3 that the temperature of the earth at the usual setting depths of surface

Jan 1, 1946