Search Documents

By Waldemar Dyrssen

I CANNOT agree with the results obtained by the committee. The gas obtained per pound of coal is too high in calorific value and the amount of gas is too large. It should require about 190 cu. ft. of oxygen per 1000 cu. ft. of gas and not 75.14 cu. ft. as is given on page 35 of the report, or about 2.5 as much. Calculating the calorific value per pound of coal from the coal analysis given on page 34, according to Dulong's formula, the gross and net calorific heat are about 12,890 and 12,430 B.t.u., respectively, which is about 18,600 and 17,900 B.t.u., respectively, per pound of carbon in the coal. This is an unusually high figure, as ordinary bituminous coal has less than 18,000 gross B.t.u. per pound of carbon. Assuming, however, that the calorific values calculated are correct, the following results are obtained. On page 35, the net calorific heat in the gas per pound of coal is given as 12,856 B.t.u. and calculation for the gross value shows it to be about 13,950; in other words, the gas contains (13,950 - 12,890) ÷ 12,890 = 8.2 per cent. more heat than the coal on a gross basis and (12,856 - 12,430) ÷ 12,430 = 3.4 per cent. more on a net basis. The calculated amount of carbon in this gas per pound of coal gasified is about 0.688 lb., or nearly the same as in the coal (0.6972) and the gross and net calorific heat per pound of carbon in the gas are 20,300 and 18,710 B.t.u. which are 9.1 and 4.5 per cent. more than for the carbon in the coal. These calculations show that the results given in the report cannot be correct. There must always be some heat losses in the gasification of coal, in radiation from the producer, and in sensible heat in the gas; there must be some loss of carbon in ashes and in soot, and there are always tarry vapors in the gases that are condensed and do not appear in the fixed gases. From the excessive heat in gas per pound of carbon contained therein, it can also be concluded that the percentage of hydrogen in the gas is too high, from which it follows that the moisture decomposed is too high and that more free oxygen must be used.

Jan 11, 1924

By R. D. Reiswig

An increasing amount of high-temperature metall~?~gical research is carried out in resistively heated tube furnaces in which a bare specimen is suspended by a fine wire at the midpoint of the tube. It is shown that the assumption of near-blackbody conditions on the surface of the sample may give rise to errors in optical pyrometry of 50°C or more. A number of papers published in the metallurgical literature in recent years have described the use of high-temperature tube furnaces for heat treatments, melting-point determinations, and other work. Typically, the tube is of a refractory metal with open ends, is heated by the passage of an electric current, and has concentric cylindrical radiation shields outside it. In at least one of these,' the specimen is contained within a crucible having a small hole in its lid through which optical-pyrometer measurements are made under essentially blackbody conditions. In another design,* the specimen is inside a muffle which is concentric with the heater tube and which contains a number of washerlike longitudinal radi- ation shields to reduce axial temperature gradients and to give conditions near the specimen closer to those of a black cavity. Yet another design3 employs a wire specimen coaxial with the heater tube. Blackbody conditions are simulated by passing current from a separate supply through the wire until it "disappears" against the background (the inside of the heater tube) as seen through a radial spy hole in the tube. Good agreement with known melting points is achieved when an optical pyrometer is used under such disappearance conditions as the melting point is approached. That this latter method works so well is due, evidently, to considerations such as those mentioned by Dike4 and others.5-8 A more common method than the above, however, has been the suspension of an uncovered specimen by a fine wire in a resistively heated tube with open ends. A radial spy hole about midway between the ends of the tube permits optical-pyrometer observations on the specimen, the inside of the tube, or both. So observed, the specimen typically appears to differ but little in brightness temperature from the inside of the tube. According to De vos9 the radiation from such a spy hole in an empty tube of reasonable 1ength:diameter ratio (e.g., 10 or greater) should be essentially blackbody radiation. For these reasons, apparently, descriptions of work carried out in such furnaces usually include a statement to the effect that conditions surrounding

Jan 1, 1964

By Maxwell Adams

Considerable interest has recently been developed in sage-brush oil because of its possible utilization as a flotation agent in the mining industry. A list of some of its physical properties, together with the method used in its extraction, may prove of interest at this time. Something over a year ago, a study of the essential oils in desert plants was begun in the Chemical Laboratory of the University of Nevada. None of the oils so far studied possess properties of special interest to engineers, except the oil of sage, Artemisia tridentata, which has exceptional power as a flotation agent. This plant, known as common sage brush, also called black sage, is widely distributed over the semi-arid West, being found quite generally on most of the dry plains and mountains west of Missouri. The method of extracting the oil followed in these experiments is very simple. The leaves, twigs and small branches are placed in an airtight drum, having a capacity of about 27 cu. ft. Steam is admitted through a number of ,small openings at the bottom of the retort, and the pressure maintained at 20 to 25 lb. per sq. in. for 3 hr. The escape of the steam from the retort is regulated by allowing it to pass through a stop-cock into a condenser. The water in the receiver is drawn off from time to time and the oil, which is insoluble and floats upon the water, is thus collected. At the end of 2 hr. most of the oil has been driven out, though traces continue to come over for a much longer time. By raising the pressure, the time required could probably be shortened and the yield increased, but the lack of laboratory equipment has prevented the carrying out of this experiment. The stock wood, bark and branches contain no oil, the distribution of the oil being limited to the leaves and young shoots. There is a seasonal variation in the amount of oil contained. Samples collected on different dates gave the following amount of oil: May 1, 0.42 per cent.; May 27, 0.6 per cent.; June 30, 0.72 per cent.; Aug. 1, 0.9 per cent.; Sept. 10, 1.0 per cent. The increase appears fairly constant from early spring, when the leaves first appear, until light frosts occur in the autumn. When the plant is air-dried there is some loss of oil, as the following data will show: Two 100-lb. samples were collected at the same time. One was distilled when green; the other was air dried for

Jan 1, 1917

By John Griffen

This paper records over twenty years' experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years' experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By John Griffen

This paper records over twenty years' experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years' experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By Whitman E. Brown

The installation in 1948 of a Humphrey spiral concentrator section at the Hill-Trumbull plant of The Cleveland-Cliffs Iron Co. is the latest commercial method on the Mesabi Range being used for the recovery of fine iron ore. In two stages of concentration 84 spirals treat approximately 120 long tons per hour of minus 1/8 in. ore. These spirals augment the production of the heavy-density plant which recovers the plus 44 in. iron from the plant crude ore. Structure of the ore is such that, when crushed, about 50 pct of the plant crude is minus 1/8 in. size. After desliming in a 66 in. Akins classifier, grinding the classifier product in a ball mill and again desliming in a 78 in. Akins classifier about 15 pct of the crude ore remains to be treated in the spiral plant.

Jan 1, 1949

By W. F. Wheeler

IN the study of the coals of Illinois now being carried on by the State Geological Survey, an attempt is being made to determine the most satisfactory basis of comparison between different coals. The following discussion is based upon work done for the survey in the laboratories of the State University and under the general direction of Prof. S. W. Parr. " Pure coal " has been defined by Mr. A. Bement1 as ash- and moisture-free coal, and the use of this " pure coal " as a basis for comparing different bituminous coals has lately been much discussed among engineers in the middle West. By some the B.t.u. of the pure coal is being used to check calorimetric results, and to calculate the calorific value of different coal-samples from the same seam, for which purposes it is necessary to assume-first, that the composition and calorific value of pure coal from different parts of the same coal-seam are uniform, and, second, that these values remain constant after the coal is mined, both of which assumptions there are reasons to believe are inconsistent with the facts. Certain variables are included in pure coal, as ordinarily defined, and disregard of them may lead to serious error. In any event, where delicate shades of difference are involved, no sure interpretation of results can be made, unless these variables be eliminated. As a basis of comparison between coals, ash- and moisture-free coal is much to be preferred to " dry coal," and still more is it to be preferred to " moist coal," as ordinarily reported in analyses. There are, however, two constituents other than ash and moisture which are similar in effect to them, and which

Jan 1, 1908

By C. F. Floe, N. J. Grant, A. Joukainen

A CCORDING to Guertler,¹ Smith and Hamilton were the first to study the Cu-Ti alloy system, but because of the presence of large amounts of impurities their data are inconclusive. Hensel and Larsen² in 1930 and Kroll³ in 1931 determined the high copper end of the diagram. In 1939 Laves and Wallbaum,4 in a short summary of compound determinations in titanium binary alloys, mentioned the existence of a face-centered cubic compound, Ti2Cu, with 96 atoms per unit cell, and also a compound, TiCu3, which was said to have a deformed hexagonal close-packed structure in which ordering had been observed. The existence of a compound, TiCu, was also found but the crystal structure was not determined. Craighead, Simmons, and Eastwood determined the effect of copper on the allotropic transformation in titanium. The investigation was limited to alloys containing from 0 to 5 pct Cu, using titanium of fairly high impurity content. Experimental Procedure The methods used for determination of the diagram included metallographic examination and X-ray and thermal analyses. The system was first worked out using the higher purity "Process A" sponge titanium and oxygen-free copper. The high titanium region was next determined more accurately using iodide titanium and vacuum-melted, oxygen-free copper. Preparation of Alloys: The alloys were prepared by arc melting under prepurified helium atmos- pheres in water-cooled copper crucibles. The high purity iodide titanium alloys were prepared by melting three compositions along with a specimen of unalloyed titanium in a single charge. A multiple crucible was used which consisted of a copper plate with four cup-like depressions. The unalloyed titanium was melted first and acted as a getter for any oxygen and nitrogen in the melting chamber. Since these elements harden titanium, the hardness of the standard specimen was an index of the highest probable contamination of the alloy melts. The sponge titanium alloy melts weighed from 100 to 200 g each and the iodide titanium melts from 20 to 25 g. In order to insure greater homogeneity most of the alloys were melted, turned, and remelted twice. The larger buttons made from sponge titanium were sectioned or crushed before remelting. These alloys up to 30 pct Cu were further homogenized by forging at 925°C. They forged readily up to 20 pct Cu but numerous cracks were formed in the alloys containing from 20 to 30 pct. Above 30 pct Cu the sponge titanium alloys were homogenized by heating for 200 hr at 850°C under prepurified helium. Similarly, the iodide-grade titanium alloys containing up to 3 pct Cu were homogenized at 1000°C for 24 hr and all others at 880°C for 200 hr.

Jan 1, 1953

By R. E. Boehler, N. C. Ludwig

At Buffington Station, Gary, Ind., Universal Atlas Cement operates fourteen 8 x 10Y2 x 155-ft cement kilns in mill 6 and two 11 x 360-ft kilns in the Harbor plant. The No. 11 and 12 kilns in mill 6 are equipped with Manitowac recuperator sections. This report describes studies in measuring exterior shell temperatures on several of these kilns and the development of a traveling radiation pyrometer with certain novel features.

Jan 2, 1960

By Heihachi Kamura

In this research, the proper temperature for the reduction of iron ores by carbon monoxide was determined. Also, the difference of the rate of reduction on the sizes of ore under four mesh per linear inch was estimated, but no evidence could be found. Rates of reduction on the seven kinds of ore were compared with the results of the experiment, showing a close relation between reducibility and density of ore. Finally, the maximum size of ore that can be reduced in applicable length of time was determined. The author thinks that the process of production of steel may be replaced by low-temperature reduction of iron ores some time in the future. THE reduction of iron oxide, such as Fe203, to iron in the blast furnace is performed principally by carbon monoxide, but partly by solid carbon by the two following reactions: Fe203 + 3CO = 2Fe + 3CO2 2Fe203 + 3C = 4Fe + 3CO2 It is known that reduction by carbon monoxide takes place at a much lower temperature than reduction by carbon and that the reducing action of carbon monoxide progresses to the inside of the ore in lump form by gaseous diffusion. In the reduction by carbon, however, the reducing action takes place only at the contact surface of the ore and carbon, so that the more complete the reduction by carbon monoxide, the more efficient is the operation. However, the reduction does not progress as shown in the foregoing formula, because carbon dioxide acts oxidizingly on the metallic iron to such a degree that when the gaseous composition reaches a certain proportion of carbon dioxide to carbon monoxide, the reduction practically stops, even though the gases are moving slowly through the heated ore. The proportions of carbon monoxide and carbon dioxide in the gas depend on the kind of iron oxide and the temperature of reduction, also the duration of time in the reaction. The longer the time of the reaction, the higher the proportion of carbon dioxide obtained in the gas, but finally it will attain to a condition of equilibrium.

Jan 1, 1925

By I. N. Knapp

Introduction.—The spccial object of these notes is to describe the mixing, testing, and use of mud-ladened water for rotary drilling in such a way as to make them helpful to the driller, the operator, or the engineer in solving his own special drilling problems. The structures, apparatus and tools used are indicated in a general way. No attempt is made to describe the art of rotary drilling; only such descriptions as are necessary to make plain the use of mud are given. The information is the result of actual experience in drilling in Coastal Plain formations. The materials encountered in the wells drilled were unconsolidated sands, gravels, and clays, in which thin layers of sandstones, shell conglomerates, and shales began to appear at about 1,200 ft, in depth, although one well was drilled to 3,018 ft. without encountering any cemented or indurated materials. Unusual Conditions.—-The general surface of the ground where the drilling was done being hardly a foot above mean tide, and the daily tidal variation being about 15 in., it was found necessary to plank over the sod surface with 3-in. plank to work from. A 30 by 30 ft. planked area was sufficient to carry the derrick with equipment and 2,500 ft. of 4-in. drill pipe stacked in it. The derrick was set up 3 ft. on cribbed blocking arranged to spread the load over the planking. The engine, bailers, and pipe yard were also carried on 3-in. planking. Drilling 0utfit.—This consisted of a derrick, 20 by 20 ft. and 84 ft. high, two 35-h.p., "oil country" boilers, one double 8 1/4 by 10 in. reversible engine, two 10 by 6 by 12 in. duplex mud pumps, one 15-in. rotary, "hoisting works" with chain drive, hoisting block with line, crown block with pulleys, rotary jetting swivels, hose, complement of tools, pipes and connections, and a mud mixer, with engine. Hydraulic Rotary Drilling Method.—This method of drilling is an adaptation of Fauvelle'sl invention made in France in 1833 for drilling

Jan 1, 1915

By William H. Waggaman

CURTAILMENT of the mineral industry as a whole undoubtedly will follow world peace, but the output of certain minerals should pursue a course well above the average on any curve of probable output projected beyond the trends exhibited through 1939. Without speculating on industrial conditions in the pending recomversion period, demand for fluorspar should be greater than before the war because of these factors, introduced by or resulting directly from military necessity: (1) Discovery and development of new deposits. (2) Improved technology in benefication. (3) New and expanded uses for fluorspar and its derivatives. The steel industry has been and continues to be the greatest consumer of fluorspar, and it wag largely to meet the wartime needs for steel that the domestic fluorspar industry has been expanded so greatly. Not only were the search for and development of new deposits of this mineral considered esstinal but the marginal nature of some of the,c- domestic deposits made it necessary to improve mineral dressinc and processing techniques to utilize them

Jan 1, 1945

By Clyde E. Williams, JAMES L. GREGG

THIS review of the past year's progress in iron and steel metallurgy presents examples of only a few of the interesting or important accomplishments made in the United States. In the field of iron ore concentration increased attention was given to the study of the more difficultly separated wash ores and some progress was made in tabling. One difficulty with the table, in concentrating iron ore where the percentage of concentrate in the feed is large, has been the removal of fine silica from the thick layer .of concentrates in the cleaning zone. One company has -Made an important contribution toward the solution of this problem by an extension of the well-known plateau principle. By a special riffling of the plateau section of the table, the concentrates are crowded together. This action forces the fine silica to the top of the bed in the cleaning zone, whence the cross-flowing currents of water -can wash it into the middling product. Some work has been done on the application of froth flotation to the concentration of iron ores. Laboratory experiments were made at the Colorado School of Mines in which hematite was, successfully floated. Using oleic acid as a frothing and collecting agent and sodium carbonate as the conditioning agent, a high recovery of rich ,concentrate was obtained. A laboratory flotation unit is in use at Taconite, on the Mesabi range, and a commer-

Jan 1, 1932

By N. Brown

IT has been observed that a hexagonal-close-packed crystal will undergo the same macroscopic displacements as an isotropic material if the basal plane is perpendicular to the axis of twist.&apos; Other orientations may produce both bending and twisting, but this analysis will be confined to pure torsion of a right circular cylinder. The z axis, the axis of twist, and the specimen axis are coincident. The displacements are given by Uq = rz, Ur = Uz = 0 [1] where 4 is the angle of twist per unit length. This requires rotational slip, as discussed by Frank&apos; and Wilman, which may be observed readily. The relative rotation of adjoining slip planes is simply produced by a movement of a planar grid of screw dislocations."." The screw dislocations are of the same sign and they move radially inward. For hexagonal-close-packed metals, the grid is hexagonal and the screw dislocations lie in the < 2110 > directions. The displacement of a point is the resultant of two displacements produced by intersecting screws sweeping over the point. The dislocations must pass over the point (r, o) in such a manner that its resultant displacement corresponds to the observed macroscopic displacement. If 6&apos; is measured from the [1010] direction, then for 0 < ? < ~/3, only the [1210] and 121101 type screw dislocations are needed to produce the required displacement. If n1 screws of type [1210] and n2 of type [2110] pass over a point, then, in order that Ur = 0 n, sin ? - n2 sin (/3 — 6&apos;) = 0. [2] If the tangential displacement is to be independent of 0, then n1 cos ? + n2cos (/3 —?) = N [3] where N is independent of ? and thus depends only on r. Nb is the tangential displacement where b is the Burgers vector. Since U? is a linear function of r, then N must be a linear function of r. If a2 is the outer radius and a, is the inner radius to which the plastic deformation has extended, then N = Nm2 r-a1 [4] This function meets the conditions that at

Jan 1, 1956

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 different 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, 1952

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 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 J. G. Faller, L. P. Skolnick

The distribution of cobalt and vanadium over non-equivalent crystallographic sites in C14-type Zr(VCo), alloys has been investigated. An anomalous X-ray scattering technique developed by Skolnick, Kondo. and lavine7 by which the separation in the scattering factors of two similar atoms can be enhanced was employed. Six alloys spanning the pseudobinary section ZrV1.6Co0.4-ZrVO.6CO1.4 at 10pct steps showed a nonrandom compositionally dependent distribution. Specifically, at high vanadium content cobalt preferentially occupied sites of type (6h) and vanadium, sites of type (2a; at low vanadium content the reverse was observed. In addition to the distribution fraction the structural parameters x and z were obtained. There was no significant deviation of these parameters from those obtained in the ideal C14 structure. Certain suggestions are made to account for the observed nonrandomness in the distribution of atoms on the two types of sites. INTERMETALLIC compounds of formula AB2 iso-morphous with MgCu2, MgZn2, and MgNi2 are known as Laves phases. Because Laves phases exhibit high symmetry and coordination numbers, the highest possible for an AB2-type compound,1 they are among the most frequently observed compounds in nature. In recent years interest has centered about the purely transition metal Laves phases2-&apos; in efforts to understand the function of atomic size and electronic structure in the formation of these compounds. It has been observed that pseudobinary Laves phase systems often show a variation of structure across the phase diagram. Such a system is the ZrV2-ZrCo2 in which the structure varies from cubic MgCu2 to hexagonal MgZn2 to cubic MgCu2.4 Some understanding about the conditions under which the second modification is stable can perhaps be gained by studying the distribution of cobalt and vanadium atoms on lattice sites in the MgZn2 modification of the system ZrV2-ZrCo2. In both the MgZn2 and MgNi2-types there exist nonequivalent positions open to occupancy by the B element, whereas in the MgCu2 prototype all sites are equivalent. Skolnick, Kondo, and La-vine7 have developed an anomalous scattering technique suitable for this type of investigation. Whereas the influence of size on the formation of a Laves phase is well recognized, no substantial evidence has been put forth in support of the size ratio dependence of a particular prototype. Berry and Raynor8 suggested that RA /RB ratio does indeed affect the type of structure that is chosen, MgZn2 compounds tending to cluster about 1.225 while MgCu2 compounds were found at larger deviations from this ratio. Dwight,3 however, from a study of 164 Laves phases does not believe this generalization to be justified. Electronic effects are certain to play a part in the stability of Laves phases in general and in the choice of a structure type in particular. For example, size along would favor the formation of Laves compounds of Ti, Zr, Hf, Ta, or Nb as the A element with nickel or copper as the B element. The absence of such is attributed to an unfavorable electron : atom ratio by Elliott and rostoker.4 Early experiments of Laves and witte9 with pseudobinary and pseudoternary systems of the three prototypes established the dependence of crystal structure upon electron: atom ratios. They observed that the MgCu2 structure dissolved elements of higher valency until the electron: atom ratio of =1.8 was reached; the MgZn2 likewise dissolved elements of lower valency replacing zinc. witte,6 from calculations of the electron : atom volumes of Brillouin Zones, obtained limits of stability for the prototypes MgCu2 and MgZn2. Elliott and Rostoker4 used these limits with considerable success in the all-transition element Laves phases they investigated. According to witte,6 compounds between the electron :atom ratios of 1.80 and 2.32 were of the MgZn2 type; those above and below exhibited the MgCu2-type structure. On the basis of these limits and an assumed valency of zirconium based upon the near tetra-valence of titanium, Elliott and Rostoker obtained valencies for the first transition series elements. For the Laves phases with which this investigation is concerned, ZrV2 and ZrCo2, the authors calculated electron :atom ratios of 2.54 and 1.56, respectively. These ratios are for the MgCu2-type structure and straddle the stability band of the MgZn2 modification. One could, therefore, predict that a pseudobinary system ZrV2-ZrCo2 should pass through the MgZn2 modification in traversing the composition diagram from one end to the other. Implicit in this assumption is a smooth change of the electron: atom ratio from 2&apos;54 to 1.56. MOSS10 states that his finding the low temperature structure of ZrCr2 to be C15 instead of C14 alters greatly Elliott&apos;s valency of zirconium and hence the assumed valencies of the other metals. Such a quantitative correlation of structure with electron : atom

Jan 1, 1963

By Hsun Hu, R. S. Cline

The annealing behavior and the mechanism of re-orientation during recrystallization of iodide titanium cold-rolled 94 pct have been studied in detail. Results indicate that recrystallization occurs by the nucleation and growth of new grains, as in other common metals. Recrystallization nuclei form by the coalescence of subgraim, and the change in texture as a result of recrystallization is largely due to selective growth among the nuclei formed. The annealing of titanium is characterized by a wide range of overlap of the various stages of the annealing process, which may be responsible for a range of activation energies observed, and for the apparently gradual change in the annealing texture as a function of time or temperature. The deformation and recrystallization characteristics of titanium and zirconium are very similar. In cold-rolled strip, the deformation texture consists of two symmetrically oriented components, each having the basal plane laterally tilted at about 30 deg from the rolling plane and the [1010] direction parallel to the rolling direction. Upon annealing for recrystallization, the change in texture can be described, for simplicity,* as rotations around [0001].2&apos;6&apos;8 According to McGeary and Lustman,&apos; recrystallization occurs in zirconium through normal growth of the subgrains, which they called "domains", without the nucleation of new grains; and the magnitude of rotation around the [0001] axis increases gradually during the progress of recrystallization. If these conclusions were true, the mechanism of recrystallization in zirconium would be basically different from that in most metals, since it is commonly known that recrystallization with reori-entation always involves the migration of high-angle boundaries. In an attempt to clarify the situation, the mechanism of reorientation during recrystallization in iodide titanium cold-rolled 94 pct was studied in detail. The structural and textural changes upon annealing at various temperatures were examined by optical and transmission-electron microscopy, X-ray pole figures, pole density distribution measurements, and micro-beam techniques. EXPERIMENTAL PROCEDURE Material and Specimen Preparation. An iodide titanium crystal bar was are-melted and solidified in a cold-hearth crucible under a purified argon atmosphere. The solidified ingot had dimensions of approximately 3 by 1/2 by 3 in. One face of the ingot was somewhat uneven, but was as clean and shiny as the remaining parts of the ingot. Large grains with a Widmanstatten internal structure were clearly shown on the shiny surfaces, indicating the occurrence of P — a transformation upon rapid cooling from the melt. Analysis of the are-melted ingot indicated C 0.033, N 0.010, H 0.013, 0 0.002 in weight percent, and traces of iron, copper, and silicon as detectable impurities. The ingot was cold-rolled -40 pct to 0.300 in. thick with a reduction of 0.005 in. per pass. The defects on the uneven side of the ingot were then removed by machining. This reduced the thickness to 0.285 in. The piece was then recrystallized by annealing at 800°C for 1 hr in a fused silica boat charged into a fused silica tube furnace under a vacuum of 10~5 mm Hg. To refine the grain size, the recrystallized metal was again cold-rolled 40 pct to 0.170 in., then annealed at 700°C for 1 hr. These treatments yielded a strip with a uniform equiaxed grain structure, having a penultimate average grain diameter of 0.04 mm and a hardness of approximately 90 Dph. Final rolling reduced the thickness from 0.170 to 0.010 in., corresponding to a reduction of 94 pct. The strip was rolled in both directions by reversing end for end between passes. Surface lubrication was provided by oil-soaked pads attached to both rolls. Specimens of 1 in. length (for X-ray examinations) and +in. length (for hardness and microstructure examinations) were cut from the rolled strip, and a width of & in. was cut from the edges of each specimen by a jeweler&apos;s saw. These specimens were then etched in a solution of 10 cu cm HN03, 5 cu cm HF, and 50 cu cm H,O to 0.008 in. thick to remove the surface metal, as well as the distorted metal at the saw cuts, prior to annealing or measurements. To minimize any surface reaction with the atmosphere, all specimens were kept in an evacuated desiccator. Isothermal Anneals. All annealing treatments were conducted in vacuum in a fused silica tube furnace as described earlier. The temperature of the furnace was controlled to within *2"C. The specimen was placed in a fused silica boat, then pushed into the hot zone of the furnace. It took about 5 to 6 min for the specimen to reach the furnace temperature. After the specimen was held at temperature for a desired length of time the boat was pulled to the cold zone of the furnace; the heating-up period was excluded from the isothermal annealing time. Thus, the uncertainty in annealing time is higher for very short anneals, but negligible for long anneals.

Jan 1, 1969

By Oscar A. Glaeser

MINE VENTILATION is an important factor in mine maintenance as well as having direct bearing on labor efficiency. Effective ventilation systems are costly, especially those for the deeper mines, but with intelligent direction the worth of a system is usually greater than its cost. In most mines using mechanical ventilation, and practically all deep mines, it has been found beneficial to equipment, timber, and the ground, to run the fan or fans continuously even though production has been curtailed or entirely suspended. Therefore, economies in ventilation may be overlooked when general cost reductions are necessary. The operating costs of a well-designed ventilation system should be capable of reduction during periods of curtailed or suspended operations. A mine that is ventilated by splits, or that can be sectionized, can reduce its fan-power costs considerably, and provisional divisions can sometimes be made in other mines, in which the air is ordinarily coursed through all the workings, without harmful effects.

Jan 1, 1932