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John R. Lucas was born Apr. 15, 1866, at St. Louis, Mo. From his 19th year until his death on his 51st birthday anniversary, he was engaged in the mining industry, in which, however, he made his beginning and achieved his progress after a somewhat unusual fashion. Presumably without previous training as a mining engineer, he became in April, 1885, and remained until the end of 1892, bookkeeper of the Granite Metal Mining Co. at its St. Louis office. Then for a year he served as bookkeeper of the same company at its mine in Montana. This official of a mining company is at the very heart of its operations; and if he possesses ambition and intelligence, may learn the business thoroughly. It is not surprising, therefore, that Mr. Lucas, being that kind of man, assumed, in 1894, a position with the Hope Mining Co., another St. Louis corporation operating in Montana, which he retained for eight years, serving not only in the office, but becoming also successively mill foreman and assayer. This varied practical experience prepared him for important administrative functions. From April, 1902, he was for many years the business agent of the Goodhope Mining Co. In 1903, he became also the assistant superintendent of the Granite Bi-metallic Consolidated Mining Co., and subsequently its superintendent-a position which he retained until his health failed at the end of 1916. In addition to the other Granite Bi-metallic properties, he had charge of the sapphire mines in the west fork of Rock Creek, Mont., the Basin Gulch placer mines, the claims of the former Hope Co., etc., all of which he managed with skill, integrity and recognized business ability.

Jan 6, 1918

By W. S. Ayres

I AM well aware that this subject is not strictly in the line of mining engineering, yet as it is a subject with which mining engineers at times have something to do, I have thought, perhaps, it might not be wholly out of place to communicate to the Institute something of the method used in determining deflection, and of its final issue. Having felt a need of some simple formula for determining the deflection of girders of uniform cross-section under any load, I have been led through the counsel of Mr. Frederick J. Slade, to whom I am indebted for many suggestions, to undertake a labor of which I shall in this paper give only some of the most important features. The results arrived at agree with those obtainable from the received formula; of Weisbach, Rankine, and others, but the expressions given by these authors, and by all others that I have at hand, are of such a form that they are not easily applied by those whose higher mathematics have become rusty from disuse, and they require, moreover, too much labor in their application to be of practical value. The general formula that I have adopted is similar to that given by Rankine, and is deflection = [mWl3 -El], in which l is the span, W the total load, E the modulus of elasticity, I the moment of inertia of the uniform cross-section of the girder, and m a factor depending for its value on the mode of distribution of the load. The problem then is to find the value of this factor m in the formula, and for some distributions of the load it leads to a number of mathematical difficulties. In this place I shall consider first the girder to be loaded symmet-

Jan 1, 1877

By S. W. Parr

THE low-temperature carbonization of coal involves the carrying out of the coking process under conditions wherein neither the coal mass nor any of the passageways through which the volatile products pass are heated above 700° or 800° C. For convenience in this discussion, the single number 750° will be used to designate the maximum range. This temperature is not selected arbitrarily; it is the result of certain natural conditions that are inherent in the substances involved. Two of these conditions arc sufficiently pronounced to suggest a line of demarcation at this point as follows: (1) Below 750°, all the heavy hydrocarbons are expelled, which means that, at these lower temperatures, the illuminants, the gases of high calorific value, and the condensible oils are discharged; above 750°, there are given off the lean, non-illuminating gases consisting for the most part of hydrogen and marsh gas and having no condensible constituents present. (2) Below 750°, there is substantially no secondary decomposition; above 750°, the volatile products are readily decomposed, forming tars, naphthalene, free carbon, etc. It is not intended to maintain that no secondary decompositions occur below 750°. Many recent studies have demonstrated the practicability, especially in the presence of catalytic substances, of cracking certain of the hydrocarbon compounds; but at these lower temperatures the step is a moderate one, as, for example, from xylene to toluene or from toluene to an anthracene. These changes are moderate in amount. Not only do the reactions proceed slowly but they are subjected to the decomposing conditions for only a short time. This is evident when it is recalled that at these initial temperatures the decomposition, of the coal is very rapid and, if anywhere near a neutral pressure is maintained, the movement

Jan 2, 1920

By Robert E. Hochscheid

The subject article appears in the February, 1978, edition of Mining Engineering, and on page 190 carries the statement, "As most operators know, in a Water-Only cyclone, ability to adjust the vortex finder is quite important due to the sensitivity of this type of operations to variance in feed conditions and having to take them apart to make this adjustment is always a nuisance." Our experience as a manufacturer and supplier of Water Only cyclones does not agree with the above statement, and we offer the following brief discussion. Our company finds that the sensitivity of the Water-Only cyclone is a function of the fineness of the coal being washed, with the sensitivity to vortex finder length being greatest for coals coarser than 3/4-in. and virtually insensitive to raw coal finer than 1 mm. For coals finer than 1 mm, we find that we can use the shortest standard vortex finder in either our 10- or 15- in. Water Only cyclone without exception. For coals 5/8-in. x 1 mm ceramic vortex finders of fixed length have been satisfactory in all cases following an initial testing period in which the length was determined. Coals coarser that 5/8-in. are more sensitive to the vortex finder length, and two installations handling coals coarser than 1-in. continued to experiment with vortex finders of various lengths long after the initial startup period. For coals of all sizes, final adjustment of product quality has been achieved by determining the optimum size of apex orifice, which is accessible and readily changed, if necessary. Comments above assume no unusual change in the coal being washed or the required quality of clean coal.

Jan 1, 1979

By C. Max Bauer

THE most important development in the petroleum industry in Montana, during 1924, was the bringing in of the Hepp well on May 12, by the Midwest Refining Co. This followed three years of active prospecting and drilling in the state by this company. The well is located 35 miles west of Billings, in Stillwater County, and makes a high-grade paraffin oil from the Dakota sand at a. depth of 3825 ft. After flowing for several months at the rate of 150 bbl. per day, the well has been shut in for lack of storage, awaiting further development in the field. TABLE 1. Production Cat Creek and Kevin-Sunburst Fields Month Kevin-Sunburst Pipe- Estimated Cat Creek Pipeline Estimated line Runs Fuel Oil Used Runs Fuel Used January 3,214 0 133,400 140 February 3,250 0 134,248 490 March 11,989 595 163,129 910 April 116,897 1,155 125,918 980 May 179,726 2,310 152,179 770 June 126,131 2,697 122,270 910 July 132,173 2,170 118,034 770 August 202,001 2,695 142,313 1,120 September 187,201 2,135 110,012 350 October 162,589 1,995 121,683 350 November 130,392 2,520 139,900 350 December 107,465 1,780 124,169 0 Total 1,363,028 20,052 1,581,255 .7,140 There are two producing oilfields in the state-the Cat Creek and the Kevin-Sunburst, which produced a total of 2,971,475 bbl. in 1924. The only other field of proved value is Soap Creek, which contains a heavy black oil and is shut in for lack of marketing facilities. Gas is produced for local use at Havre, Chinook, and Baker, and a carbon-black plant is located at Baker.

Jan 3, 1925

By I. N. Knap

Discussion of the paper of HOWARD R. HUGHES, presented at the Now York meeting, February, 1915, and printed in Bulletin No. 99, March, 1915, pp. 629 to 635. I. N. KNAPP, Ardmore, Pa.-The claims made in the paper for the Sharp & Hughes cone bit are modest. I have used it with satisfactory results. It is a wonder for eating through hard, brittle rock; it will drill through iron pyrites that the fish-tail bit cannot enter. Soft ground is apt to contain concretions that are much harder than the containing ground. ?We have at times in drilling with a fish-tail bit thought the hole was partly in such hard material and partly in soft, thus making a difficult place to drill past in the ordinary way. By changing to a Sharp & Hughes cone bit such obstructions were easily overcome. This cone bit has greatly widened the range of work that may now be profitably accomplished by the hydraulic rotary method of drilling. It is not, however, adapted to drill soft or sticky ground, the old fish-tail bit being better for such purposes. The fish-tail bit usually scrapes its way through the material drilled, but not always. I believe at times it cuts and chips in soft friable rock. That is, the bit seems to bite into the rock, and when sufficient torsion is put on the drill stem it breaks its hold and jumps around and chips the rock from the bottom of the hole, like a stone cutter's chisel, and then takes another hold. This condition occurs only in friable material when the bit is sharp and the feed properly controlled. The Sharp & Hughes cone bit is a most valuable -tool for the rotary driller.

Jan 5, 1915

By H. R. Peterson

The control of autogenous grinding circuits is very similar to the control of conventional grinding circuits. In both circuits the classifier must be operated to achieve the required degree of mineral liberation. Crude ore grade and grinding mill density are important in both circuits. The biggest difference between the control of the two circuits is the pronounced effect that crude ore changes have on autogenous circuits. The importance of crude grade, weight recovery, size and hardness is greatly magnified because the crude ore in the autogenous circuit is also supplying the grinding media. Crude ore sizing is a particularly critical factor in achieving efficient primary autogenous milling. Because of the above, computer control is advantageous for autogenous circuits. Changes in crude ore may necessitate rapid changes in the entire circuit control strategy, and only a computer can facilitate these changes. Also, the techniques of control are very different for conventional and for autogenous grinding. Set-point control is usually satisfactory in conventional circuits, but the multiplicity of varying conditions in an autogenous circuit usually require some type of computer optimizing search routine. Autogenous mills can be separated into two general types. In primary autogenous mills the media is obtained from the mill feed, and in pebble mills the sized pebble media is added separately from the material being ground. PRIMARY AUTOGENOUS MILL CIRCUIT CONTROL - FIGURE 1 Power Draw Primary autogenous mills are usually run at full motor loading so production will be at maximum. Therefore, power draw must be closely controlled; this is done by adjusting feed rate. The versatility of computer control is extremely valuable for power control, especially during starting and stopping. To bring the mill power up as fast as possible at

Jan 1, 1982

1. Any society of undergraduates at a technical school, comprising students in any branch of engineering, metallurgy, chemistry, geology, etc., may be recognized by the Board of Directors in its discretion as an Affiliated Student Society, and its name, together with the names of its president and its secretary, may be published in Mining and Metallurgy or the Directory of the Institute. The members of such a society will not be named individually on the membership list of the Institute, but may at any time be proposed for election as Student Associates or Junior Members of the Institute, in the usual way. 2. Membership in an Affiliated Student Society, certified by the official representative thereof, and accompanied by suitable testimony as to character and qualifications, will have special weight with the Membership Committee of the Institute, in its consideration of the proposal of a candidate for membership. The Affiliated Student Societies receive the following privileges: 1. A copy of Mining and Metallurgy free; 2. The use of Employment Service; 3. The use of Mining and Metallurgy for publicity purposes, such as printing notices of their activities, etc.; 4. Assistance in obtaining speakers for their meetings, lists of moving-picture films, etc. To retain the privileges of the Affiliated Student Society, its officers must keep in touch with the Secretary of the Institute and advise him of its activities. The foregoing plan is not confined to students in mining engineering or metallurgy. Any undergraduate society of engineers may avail itself of the privileges offered. Societies desiring to be recognized as Affiliated Student Societies are requested to make application to the Secretary of the Institute, giving particulars as to the nature, membership and officers. The following list gives the names of the Affiliated Student Societies as of December 1, 1929, and their presidents and secretaries.

Jan 1, 1929

By Paul F. Gnirk

INTRODUCTION The notion of permanent disposal of high-level radioactive waste in natural salt formations was formally proposed in 1957 by the Committee on Waste Disposal of the National Academy of Sciences- National Research Council. This proposal was the result of the Committee's deliberations in response to a request from the U.S. Atomic Energy Commission (AEC) in 1955 to assess the viability of geologic formations for such disposal. The research investigations for waste disposal in salt were initiated at the Oak Ridge National Laboratory (ORNL) in the late 1950's and continued through the 1960's and early 1970's at a relatively modest level of funding. During the period of transfer of Federal Administrative authority from the AEC to the Energy Research and Development Administration (ERDA) to the Department of Energy (DOE) during the latter part of the 197O’s, financial support for the research program was enhanced substantially. The program at OWL was regrouped under the Office of Waste Isolation (OWI) and later transferred to the Office of Nuclear Waste Isolation (ONWI) at Battelle Memorial Institute. With this enhancement in the research program, consideration of basalt, granite, shale, and tuff as repository host media was initiated. In February, 1980 the President of the United States established officially a comprehensive waste management program which called for the selection of the first repository site in 1985 from a suite of four or five potentially suitable sites, and initiation of a repository operation at that site in the mid-1990's. The current disposal system concept involves the emplacement of radioactive waste packages in mined disposal rooms at depth in suitable geologic formations, with closure of the facility effected

Jan 1, 1983

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By J. Parke Channing

THE caving system of mining is that method of removing the ore from an underground body in which the top is first attacked and mined out and the capping, or roof, as the case may be, is allowed to fall in or "cave" and fill the space formerly occupied by the mined ore. It is applicable, generally, to large orebodies where the ore is relatively soft. We may say, within moderate limits, that when ore is removed the surrounding ground either caves in or remains open. For example, in mines like the Lake Superior amygda-loid copper mines, when the lode is removed there is left a void which remains open for many years; on the con-trary, mining a Lake Superior soft hematite ore, however carefully it may be done, always produces a caving of the surrounding rocks. We may say, with propriety, therefore, that the caving system is applicable to that type of ore deposit in which the surrounding rock readily caves. Another factor which governs the use of the caving system is the value of the ore itself. Admittedly, any form of caving will not extract 100 per cent. of the ore, and in the case of a very rich orebody it is often desirable to get out every pound of the ore with as little dilution as possible. Under these conditions, some other method of mining, such as that by square sets and filling, is indicated and should be followed. We may say, however, that, looking at it broadly, in the United States the caving system of mining is generally used for the removal of soft hematite and for the so-called "porphyry" coppers. In both cases this method is used where open-pit mining is not possible.

Jan 1, 1922

By P. J. Jorgensen, J. H. Westbrook

The anomalous indentation creep of nonmetallic solids is shown to be due to the presence of adsorbed water. Although a specific mechanism is not proposed, it is suggested that the water may be present in two chemically adsorbed states. DEPENDENCE of the indentation hardness of solids on the time of load application was first studied systematically by Hargreaves1 in a study of lead and certain low-melting alloys. This technique has become known as indentation creep and has been exploited particularly by underwood and by Mulhearn and Tabor3 as well as by others.4°7 For metals good correlations have been found between the time dependence of hardness over several orders of magnitude in the seconds range and creep parameters from conventional tests at the same temperature where creep strains were measured over hundreds or thousands of hours. However, Walker and Demer8 and Mitsche and Onitsch9 observed extensive indentation creep at room temperature (particularly for light loads) in certain minerals and in materials such as lithium fluoride and magnesium oxide for which conventional creep is ordinarily not observable unless the testing temperature is at a very much higher fraction of the melting point. Normal behavior was observed' for metals such as copper, i.e., negligible time-dependent flow during room-temperature indentation. Walker and Demer attempted to find some distinctive difference in the surface deformation markings about indentations in the anomalously behaving materials at short and very long indentation times. No significant differences were observed. Increased duration of loading was found, however, to result in an increase in the length of glide of edge-dislocation loops extending into the interior of the crystal. Jorgensen and Westbrook,10 in a recent study of

Jan 1, 1965

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons- per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead, mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding, and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1952

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By M. E. Wadsworth, D. R. Wadia

The rate of reaction of cuprite was measured in a series of sulphuric acid solutions, from which oxygen had been excluded, at various concentrations and temperatures. The overall reaction CuzO + H2S04 + Cu " + CuO + H20 + So4' may be explained quantitatively by two simultaneous rate reactions involving not only H+ but also the hydrolytic adsorption of H2SO4. The results indicate the importance of considering surface mass balance relationships associated with heterogeneous surface reactions. T EACHING has long been an important method for the recovery of metals from certain minerals. Many processes have assumed minor importance in modern practice but such operations as the recovery of uranium and present interest in high pressure-high temperature solution of sulphides' ' well illustrate the importance of leaching as a unit operation. Little quantitative information is available in a form sufficient to determine the mechanism of solution at the solid-aqueous solution interface for given systems. For this reason, it is important that fundamental studies of leaching processes be conducted so that parameters associated with such surface reactions may be understood and correlated. Cuprite was chosen for this study as a result of a recent investigation carried out at the University of Utah on ores of the Ray and Chino Mines Divs. of the Kennecott Copper Corp. These sulphide deposits contain sufficient oxidized minerals, such as cuprite, malachite, and azurite, so that recovery of metal values contained in the oxidized fraction is of primary interest. Of these oxides, cuprite (Cu,O) is the most interesting, since it dissolves in sulphuric acid by an oxidation reduction couple resulting in the formation of 50 pct metallic Cu and 50 pct cup-

Jan 1, 1956

By Peter I. Brooker

Elementary estimation methods in which blocks are assigned values according to a geometrical area of influence concept are investigated for a tabular deposit. This quantitative assessment of the estimation technique is made using Matheron's geostatistics which allows the variance of the error between actual block values and estimated values to be found once the semi-variogram is known. For the deposit in question it is shown that the elementary technique makes poor use of the available information. There is a 20% relative error associated with it, whereas by using the optimal technique of kriging this can be reduced to 9%. An infill drilling program, which might have been suggested on area of influence results, would reduce the estimation variance only marginally over that of the current campaign; the relative error being lowered to 6%. Thus the geostatistical analysis shows this would be an unnecessary expense. The accuracy of an earlier drilling stage is also studied and it is shown that with the information available at the earlier stage block values could have been known with a relative error of 13%. Drilling could have stopped then if this were reckoned a tolerable error level. These results are all obtained for blocks whose size is dictated by the mining engineer and is set at the current grid size. In the area of influence approach smaller blocks are considered as the grid becomes finer and a study of the accuracy of area of influence estimates and kriged estimates for differing block sizes is presented. Between drilling stages more information from which the semi- variogram is calculated becomes available. The sensitivity of the estimation to changes in the semi-variogram parameters between earlier and current drilling programs is investigated.

Jan 1, 1977

By H. E. Stauss

THE short-time tensile-strength test, while it has not attained the practical importance of the creep test at elevated temperatures, has the advantage of being rapid and is satisfactory for determining relative strengths. For platinum and its alloys, which frequently are used at high temperatures under ill-defined conditions, the short-time test provides useful data that aid in the selection of the most promising alloy for each use. Measurements have been made upon 13 alloys, using wire 0.0031 in. in diameter, the choice of alloys and of wire size being determined by the requirements of a practical problem. Later the measurements were extended to higher temperatures by the use of heavier wire, 0.010 in, in diameter. APPARATUS The use of fine wire made the necessary equipment very simple, since no effort was made to measure the elongation. A small platinum-wound horizontal furnace was used, ½ in. in internal diameter by 1 ft. long, with the wire threaded through it. The wire, clamped at one end while the other ran over a ball-bearing pulley, of a total length of 29 in., was loaded directly by running sea sand into an attached basket, making the rate of loading and not of elongation constant. The sand flowed from a funnel, under a head of 3 in., into the basket. 3 in. below the orifice, at the rate of 170 grants per minute. Pulley friction and bending resist-ance of the wire as well as the force of impact of the sand could be neg-lected, being 3 grams and 0.5 grams, respectively. Furnace temperature .was uniform within ±5° C. over a length of 3 in., although a drift with time occurred, the largest amount being 20° C. No provision was necessary for a controlled atmosphere.

Jan 1, 1939

Joseph K. Stone, Jr. (Kaiser Engineers, Oakland, Calif.)—I would like to compliment the authors on their clear presentation of the results of their study of the pneumatic processes. We hope that the Inland Management decision to make this information available to all will induce others to release additional valuable fundamental data which they have accumulated in the field. There are two points upon which we would like to comment. In the discussion of Table VII it was indicated that the "oxygen consumptions were about the same as has been reported for other oxygen steelmaking processes". This conclusion is apparently based upon the consumption of 2100 cu ft per ton shown in Table VII. The data in Table IV also permits calculation of 2520, 2280, 2250 and 2700 cu ft per ton in heats 1 to 4, respectively. These figures are in the same general range as those reported for the Kal-Do and Rotor Processes using Thomas iron, but considerably above the following figures for four of the operating L-D Process Plants while using iron which is similar to American basic iron: Donawitz 1760 cu ft per ton Adapted from Cuscoleca' LinZ 1900 cu ft per ton Trenkler & Hauttmann2 Dofasco 1850 cu ft per ton D. O. Davis:3 In addition to the differences between the consumption of gaseous oxygen in the Inland operation and in the L-D Process, there is another closely related difference. The L-D plants for which data are cited are using approximately 1 pct ore while the authors' process used about 7 pct of sinter. Assuming 65 pct Fe in the sinter, the difference amounts to 400 to 500 cu ft

Jan 1, 1959

By Z. T. Bieniawski

The importance of understanding the phenomena associated with rock fracture has long been fully appreciated in rock mechanics. This is clearly apparent from the special attention paid to rock fracture in the comprehensive papers published within the past four years.l-5 These papers present a picture of how the knowledge of brittle fracture of rock was gradually enriched and also show that what was often taken for granted three or four years ago is no longer considered as valid today. The Griffith hypothesis,6 thought in the past to predict the strength of rock, is now known to be applicable to fracture initiation only. The knowledge of a failure criterion for fracture initiation is not sufficient, however, as after fracture initiation the rock is still a useful material even possessing increasing load-bearing ability. The importance of the fracture propagation process, subsequent to fracture initiation, is thus obvious. Yet, relatively little attention has so far been paid to brittle fracture propagation in rock. As it is now realized that understanding of this process is a prerequisite for practical rock mechanics applications, studies of propagation of brittle fracture in rock constitute an important aspect of rock mechanics research. It has been pointed out 4 that it is essential that the failure mechanism is fully understood before a failure criterion is postulated. Many failure criteria have been propounded in the past on the basis of theoretical reasoning alone and could not be verified by experimental evidence. Recently, a hypothesis for the mechanism of brittle fracture of rock was propounded, 4 describing the events taking place in the rock from the initial application of gradually increased load to complete disintegration of the material.

Jan 1, 1972