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By Harlan A. Walker

Rock-drill steel has an important bearing on costs in many mining operations, both directly and indirectly. Direct factors include such items as shop expense, steel consumed per ton of ore produced, cost of distribution and the dust hazard. The latter, although no specific amount may be assigned to it, is an ever-present problem, and is receiving increased attention as the science of mining progresses. Indirectly drill-steel performance affects consumption of compressed air, maintenance of drills and drilling speed. The latter, in turn, increases or retards the tempo of mining and development. These factors, while not all inclusive, are preponderant in their importance. Thus it is evident that in an extensive operation drill-steel expense may run to many thousands of dollars and that thoughtful research leading toward more efficient practice is warranted. Experimental Work During the late spring of 1939, experiments were undertaken at Homestake in reducing changes in drill-steel gauge from ? in. to 1/16 in. In drifts and crosscut headings mounted machines were set up and holes were drilled with ?-in. bit-gauge changes and other holes a few inches

Jan 1, 1940

By Harlan A. Walker

Rock-drill steel has an important bearing on costs in many mining operations, both directly and indirectly. Direct factors include such items as shop expense, steel consumed per ton of ore produced, cost of distribution and the dust hazard. The latter, although no specific amount may be assigned to it, is an ever-present problem, and is receiving increased attention as the science of mining progresses. Indirectly drill-steel performance affects consumption of compressed air, maintenance of drills and drilling speed. The latter, in turn, increases or retards the tempo of mining and development. These factors, while not all inclusive, are preponderant in their importance. Thus it is evident that in an extensive operation drill-steel expense may run to many thousands of dollars and that thoughtful research leading toward more efficient practice is warranted. Experimental Work During the late spring of 1939, experiments were undertaken at Homestake in reducing changes in drill-steel gauge from ? in. to 1/16 in. In drifts and crosscut headings mounted machines were set up and holes were drilled with ?-in. bit-gauge changes and other holes a few inches

Jan 1, 1940

By H. V. Flagg

OPERATION of a modern steel plant presents a curious anomaly. Large-scale operations, in which large volumes or heavy weights of materials are involved, are not usually subject to close control or narrow specifications. On the other hand, operations that call for a high degree of precision are usually carried out under carefully controlled conditions with delicate, precise machinery and tools, and production is made secondary to quality. In a modern steel mill both extremes are taken as a matter of course. A strip mill making auto-body sheets must supply its customers with sheets that all come within a tolerance of plus or minus 0.003 in. in gauge at any point from side to side or end to end, and without blemishes either of color or surface. Sheets that were considered par excellence as recently as ten years ago could not even find a market today. Such exacting requirements have made necessary a nicety of control that approaches the watchmaker's art. Notwithstanding this necessity for precision, the scale of operation is stupendous.

Jan 1, 1940

By L. R., McLeod

NO important changes in metallurgical practice are exhibited in the smelter that has been erected for treating the Morenci concentrate: the material is charged to reverberatory furnaces without roasting, and the matte subsequently blown in Peirce-Smith converters to separate the iron and sulphur from the copper. In fact the process is essentially the same, except for preliminary roasting of the concentrate, that was adopted in the new smelter that was built for the Arizona Copper Co. at Clifton, a few miles away, to treat Morenci concentrate just before the start of the first World War, and which went into service in October, 1913.

Jan 1, 1942

By O. W. Shimer, D. C. Helms, C. E. Brown

The early history and progress of anthracite stripping, from the first known operation at Summit Hill in 1821 through 1917, was covered in 1917 in a paper by J. B. Warriner,1 then chief engineer, now president of the Lehigh Navigation Coal Co., and stripping methods between 1917 and 1931 were the subject of a paper in 1931 by H. H. Otto,2 of the Hudson Coal Company. The present paper is confined to methods used today on pitching beds in the western middle and southern anthracite fields. The anthracite industry, not unlike other large industries, had been affected for more than a decade by the adverse economic conditions of the country. High development and operating costs and the competition of other fuels caused some collieries to be closed temporarily, while others were permanently abandoned. In order to obtain a larger tonnage at a lower development cost, a larger stripping program was put into effect by many companies. In 1932, stripping operations accounted for 3½ million tons, or 7 per cent of the total anthracite production of the year, and in the next 10 years this increased until more than double that amount, or over 14 per cent of the total production, was from stripping operations. These figures represent the total production of the entire anthracite region; unfortunately, the figures have not been separated from the different fields. There are some operations in the southern and middle western fields where the tonnage obtained from stripping amounts to more than 30 per cent of the total colliery production. Stripping operations on pitching beds present difficult problems. Owing to the various geological formations, a different situation is encountered in nearly every stripping pit; and because of the proximity of an overlying bed, it is very often necessary to load the entire overburden into trucks and haul it out of the pit to the dump area, although back filling and casting is done whenever possible. Beds in the areas described in this paper pitch from 20" to 90° and some are on the invert in many sharp folds and faults. Planning The more successful strippings have had adequate engineering analysis. Some work is being done without much preliminary planning, but many strippings would have failed completely if close supervision by the engineering department had not been maintained. Locations of stripping areas are determined as follows: I. By constructing cross sections over the areas and drilling sufficient diamond-drill or well drill holes to study the strata and to determine the location, thickness and pitch of the beds.

Jan 1, 1946

By O. W. Shimer, D. C. Helms, C. E. Brown

The early history and progress of anthracite stripping, from the first known operation at Summit Hill in 1821 through 1917, was covered in 1917 in a paper by J. B. Warriner,1 then chief engineer, now president of the Lehigh Navigation Coal Co., and stripping methods between 1917 and 1931 were the subject of a paper in 1931 by H. H. Otto,2 of the Hudson Coal Company. . The present paper is confined to methods used today on pitching beds in the western middle and southern anthracite fields. The anthracite industry, not unlike other large industries, had been affected for more than a decade by the adverse economic conditions of the country. High development and operating costs and the competition of other fuels caused some collieries to be closed temporarily, while others were permanently abandoned. In order to obtain a larger tonnage at a lower development cost, a larger stripping program was put into effect by many companies. In 1932, stripping operations accounted for 3½ million tons, or 7 per cent of the total anthracite production of the year, and in the next 10 years this increased until more than double that amount, or over 14 per cent of the total production, was from stripping operations. These figures represent the total production of the entire anthracite region; unfortunately, the figures have not been separated from the different fields. There are some operations in the southern and middle western fields where the tonnage obtained from stripping amounts to more than 30 per cent of the total colliery production. Stripping operations on pitching beds present difficult problems. Owing to the various geological formations, a different situation is encountered in nearly every stripping pit; and because of the proximity of an overlying bed, it is very often necessary to load the entire overburden into trucks and haul it out of the pit to the dump area, although back filling and casting is done whenever possible. Beds in the areas described in this paper pitch from 20" to 90 and some are on the invert in many sharp folds and faults. Planning The more successful strippings have had adequate engineering analysis. Some work is being done without much preliminary planning, but many strippings would have failed completely if close supervision by the engineering department had not been . maintained. Locations of stripping areas are determined as follows: I. By constructing cross sections over the areas and drilling sufficient diamond-drill or well drill holes to study the strata and to determine the location, thickness

Jan 1, 1944

By J. F. Peters

The term solution loss is discussed and defined. Examples are given showing that solution loss may either have a favorable or unfavorable effect on blast furnace performance. A theory is advanced explaining the contradictions encountered during earlier studies of the problem. MANY papers have been written and numerous theories advanced concerning the chemical reactions in the iron blast furnace. Richards, discussing the utilization of fuel in the blast furnace, said: "All the carbon burnt in the furnace should be first oxidized at the tuyeres to CO and all the reduction of oxides above the tuyeres should be caused by CO, which thus becomes CO,. This dictum is not Gruner's own words, but expresses their sense, and from the point of view of the present discussion, it is the correct principle upon which to obtain the maximum generation of heat in the furnace from a given weight of fuel." Richards also said the cubic feet of wind per pound of coke does not express how efficiently a furnace is running, but it will be shown later where Howland paid much attention to this figure. Richards pointed out a shortcoming of Grun-er's discussion in that Richards claimed there is not enough CO generated in a good working blast furnace to possibly combine with the oxygen of the ore, so some of it must be reduced by solid carbon. Mathesius2 made two important contributions. First, he wrote the direct and indirect reduction equations showing the heats of formation in each case and stressed that the indirect reduction reaction produces heat while direct reduction absorbs heat. Secondly, he substantiated Richards' conclusion that there may be a deficiency of CO gas and that some direct reduction is required, but he pointed out that "direct reduction in the hearth has often been confused with direct reduction in the stack, which more correctly should be termed 'premature combustion'." Then he said that "as this carbon is consumed partly by economical (direct reduction in the hearth) and partly by wasteful reactions in ever varying proportions, it is evident that the relation of this total carbon gasified above the tuyeres to the fuel consumption of a furnace cannot possibly have a direct bearing on the economy of furnace operation. The premature combustion of carbon (CO2+C?2CO) must therefore in all cases be considered a detrimental reaction." Johnson3 discussed this whole problem. Through experience Johnson found that when departure from Gruner's ideal working is considerable, the fuel economy is poor, and when the quantity of blast goes up, the fuel economy goes up in the same proportion. Johnson said that the wind per pound of coke is a measure of fuel economy, which will be shown to be wrong. Howland4 established the fact from operating data that there was no relationship between the coke consumption of a furnace and the percentage of coke burned at the tuyeres. But some of his calculations have led to questionable conclusions, such as "it is practically impossible to obtain low coke consumption unless we keep our wind low." Howland made an important contribution when, in his concluding paragraphs, he reached certain negative decisions as to why one coke works better than another. He said the reason one coke works better than another in a blast furnace is not because: 1—there is any difference in the percentage gasified at the tuyeres, or 2—there is any difference of wind required per pound of coke. Korevaar5 propounded a new theory on combustion which disagrees with Gruner because "as far as we can see, this is sufficient proof for the invalidity of Gruner's ideal, for if it was valid, the Low coke consumption should always be accompanied by a higher percentage of carbon burned at the tuyeres. This not being the case, we must deliberately give up our belief in Gruner's ideal." The part in italics will be proved to be wrong. Martin" contended that there was not enough reducing capacity in the blast furnace unless solution loss occurs. He came to a series of conclusions, such as: 1—The greatest efficiency of the blast furnace may not be attained when the reduction is performed entirely by carbon monoxide as demanded by Gruner's definition of the ideally perfect working of the blast furnace. 2—The so-called solution loss reactions, which should more properly be termed direct reduction reactions, promote furnace efficiency. 3—In modern blast furnace practice, the carbon consumption of the process is determined primarily by the carbon needed for reduction purposes, any thermal deficiency created by the reduction process being balanced in practice by the use of blast heat. From a practical standpoint further discussion of the theory of chemical reactions is of little moment. There is however one phase of the general theory of chemical reactions which is very important and that is the combustion of coke in the blast furnace. The purpose of this paper is to show that the reaction

Jan 1, 1955

By E. P. Sadowski, R. J. Raudebaugh

A study of micro structural characteristics of a 30 pct Cr, 42 pct Ni, 26 pct Fe alloy has been correlated with its behavior in creep and rupture tests at 1400°, 1600°, and 1800°F. Nitrogen pickup observed at 1600" and 1800°F was associated withfissur-ing of specimens prior to fracture. IN studying the high-temperature characteristics of a nominally 30 pct Cr, 42 pct Ni, 26 pct Fe alloy certain anomalies were observed, namely: The alloy had greater stress rupture and creep strength at 1800" than at 1600° F for time periods beyond 550 hr. As placed in test, this alloy was nonmagnetic. After test, some specimens were found to be fairly magnetic. The gain observed in magnetic response was not solely a function of time and temperature of testing. This is a report of the observations and a recounting of the results of studies made in an attempt to explain the behavior of the material under investigation. MATERIAL AND PROCEDURE All test material was taken from a production size heat of an experimental alloy melted and processed using commercial practice and having the following percentage composition: C Fe Cr Ni Ti 0.05 25.83 30.13 41.59 0.41 Al Mn Si S Cu 0.42 0.75 0.45 0.011 0.34 The material was tested in two conditions, namely mill-annealed and grain-coarsened. Mill annealing, which consisted of heating to about 1900° F in a continuous furnace for about 30 min and air cooling, gave a grain size of ASTM 8 and finer. To produce the grain-coarsened condition, mill-annealed material was soaked for 2 hr at 2050°F in a laboratory furnace and air cooled. This resulted in an ASTM grain size of 4 to 5. Rupture tests were carried out at 1400, 1600, and 1800° F in accordance with ASTM procedures. In tests where only rupture life and elongation at fracture were determined, specimens of 0.252 in. diam and 1 in. gage length were used. In tests where creep curves were obtained, specimens were of 0.375 in. diam and 4 in. gage length. Several of the tests under relatively low stress were discontinued short of rupture. Three low-stress tests are still in progress at this writing. Structural changes which occurred during testing were determined from examination of the tested specimens by light microscopy and X-ray diffraction. Diffraction pattern recordings were made with a diffractometer using copper radiation on samples heavily etched in a picric-hydrochloric

Jan 1, 1960

By H. M. Chance

In recent years an important function of the mining engineer has been the appraisal of the value of mining-properties required by those planning consolidations of a number of' individual operations. In no field of productive industry have the economic results attainable by consolidation been more clearly shown than in the mining-regions of the United States. In many important mining-centers unification of ownership of the most valuable properties has already partially or completely been attained, and, in other districts, projected absorptions and consolidations await only favorable financial conditions for their consummation. The iron-ore districts of the Lake Superior region, the coal-mines of Colorado, of Wyoming and of Utah, the anthracite mines of Pennsylvania, the Connellsville coking-coal mines, and large areas of bituminous coal-land in Pennsylvania, in Ohio, in the Indian Territory and in other portions of the United States, are now largely controlled by a few corporations operating on a large scale, or by a number of corporations working in harmony because of an affiliation and interleaving of interests. The tact of keen financiers and the diplomacy of men of great administrative ability are the chief operative forces in conceiving and accomplishing such consolidations, but in formulating and maturing them, the services of the mining engineer are often required to supply the facts upon which they must be based. For this purpose the expert is expected to furnish full and complete reports covering the quantity of ore, coal or other mineral available, its value per ton, cost of mining and treatment, etc. In reporting upon gold-, silver-, lead-, zinc-, copper- or iron-

Jan 1, 1905

By R. E. Salvoti

IN underground coal mining, the increasing trend towards mechanical methods is ever apparent. Figures for 1939 showed that 28 per cent of the total bituminous coal production was mined mechanically 1940 statistic will reveal an even higher figure. Factors causing this trend are varied, but the important are the following;

Jan 1, 1941

Bernard MacDonald, Guanajuato, Mexico (communication to the Secretary*):—Mr. Iceyes's paper is very interesting to me because of my personal experience with the development of the ore-deposits of Lake Valley.

Jan 1, 1909

By E. C. Potter

FUEL-OIL was first substituted for coal at these works in September, 1888. It was first applied in the converting-department to the battery of boilers, consisting of 14 tubular boilers, 16 feet in length by 5 feet in diameter. To operate these boilers with coal for 24 hours required 14 strikers, 3 ash-wheelers, 6 men unloading coal, and 2 water-tenders, making a total of 25 men. With fuel-oil the same boilers require for the same period 2 water-tenders and 4 men to attend to the burners, making a total of 6 men. This gives a saving of 19 men, which, at an average of $2 per day each, amounts to $38 per day. Comparing the consumption of oil with that of coal, we find that for the week ending January 5, 1889, with an ingot output of 6403 tons, 2731 barrels of fuel-oil were used as against a necessary consumption of coal for the same work of 848 tons, showing 3.22 barrels of fuel-oil to be equivalent to one ton of Indiana block-coal. Figuring oil at 60 cents per barrel, and coal at $2.15 per ton, we have as the cost of oil $1.93, as against $2.15 for its equivalent in coal. In December, 1888, the rail-mill battery of 26 boilers, of the same dimensions as those in the converting-department, was equipped for fuel-oil. For the week ending January 5, 1889, with a rail out-put of 5208 tons, 5987 barrels of oil were consumed, as against a necessary consumption of coal, to do the same work, of 1805 tons, showing 3.31 barrels of fuel-oil to be equivalent to one ton of Indiana block-coal, which at the costs named above would show $1.98 worth of fuel-oil to he equal to $2.15 worth of coal. The labor required to operate the rail-mill boilers for 24 hours, with coal as fuel,is: 26 stokers, 4 water-tenders, 6 ash-wheelers, 12 men unloading coal; total, 48 men. With oil as fuel, the same work requires 4 water-tenders and 6 men to attend to the burners; total, 10 men; a saving over coal of 38 men in 24 hours, which, at an average price of $2 per day, would be $76 per day. In the brief experience we have had with oil as fuel, we find that the efficiency of the boilers is somewhat increased and the repairs are

Jan 1, 1889

By K. L. Sutherland, J. Rogers

The relationships between collector and mineral, activator and mineral, and activator, collector and mineral will be considered herein. We propose to criticize current theories of flotation but we will not repeat the many pertinent arguments presented by Wark and Cox28 in 1937 in which the "chemical" or more strictly the "solubility" theory proposed by Taggart and co-workers22 is shown to be untenable. Recently Taggart, Arbiter and Kellogg24 have studied the long-chain aliphatic amines as collectors and have attempted to resuscitate the solubility theory for these particular collectors. Dean and Ambrose3 have felt that "we stumble over words" when we distinguish different mechanisms of adsorption. Our view is that nothing is more stultifying to the development of a science than attempts to force widely different phenomena into one classification. This oversimplification leads immediately to difficulties in the interpretation of results. Experimental Procedure The response of minerals to activator and collector was determined by "captive-bubble" tests. The methods of preparation of specimens26 and of testingin have been previously described. Results from captive-bubble tests were confirmed by flotation tests in glass-stoppered cylinders.26 The full lines shown in the figures relating solution composition to contact (flotation) distinguish solutions in which contact (flotation) is possible from those in which it is not. The curve is derived by successive approximation and the accuracy is ±o.2 pH units and ±5 mg. per liter of collector or depressant. Solution compositions for which it was shown experimentally that contact or flotation is possible are marked by a circle, those for which it is not possible by a cross and those for which it is doubtful by a cross within a circle. It is assumed that at all other solution compositions within (without) the curve, contact or flotation is possible (not possible). After testing in the collector solution, the mineral was removed, washed, and tested by the captive-bubble method in water, or floated in a cylinder with frother only. In this way the composition of solutions in which a mineral was conditioned but in which properties of the collector solution prevented contact or flotation can be defined. The composition of solutions in which conditioning was possible lie between the full and broken curve, the experimental results being indicated on the figures by a square. The curves show the limits of composition of solution in which flotation in a cell would be possible. Many factors other than those considered in these tests determine recovery or grade of a given mineral in practice.

Jan 1, 1947

By J. Rogers, K. L. Sutherland

The relationships between collector and mineral, activator and mineral, and activator, collector and mineral will be considered herein. We propose to criticize current theories of flotation but we will not repeat the many pertinent arguments presented by Wark and Cox28 in 1937 in which the "chemical" or more strictly the "solubility" theory proposed by Taggart and co-workers22 is shown to be untenable. Recently Taggart, Arbiter and Kellogg24 have studied the long-chain aliphatic amines as collectors and have attempted to resuscitate the solubility theory for these particular collectors. Dean and Ambrose3 have felt that "we stumble over words" when we distinguish different mechanisms of adsorption. Our view is that nothing is more stultifying to the development of a science than attempts to force widely different phenomena into one classification. This oversimplification leads immediately to difficulties in the interpretation of results. Experimental Procedure The response of minerals to activator and collector was determined by "captive-bubble" tests. The methods of preparation of specimens26 and of testingin have been previously described. Results from captive-bubble tests were confirmed by flotation tests in glass-stoppered cylinders.26 The full lines shown in the figures relating solution composition to contact (flotation) distinguish solutions in which contact (flotation) is possible from those in which it is not. The curve is derived by successive approximation and the accuracy is ±o.2 pH units and ±5 mg. per liter of collector or depressant. Solution compositions for which it was shown experimentally that contact or flotation is possible are marked by a circle, those for which it is not possible by a cross and those for which it is doubtful by a cross within a circle. It is assumed that at all other solution compositions within (without) the curve, contact or flotation is possible (not possible). After testing in the collector solution, the mineral was removed, washed, and tested by the captive-bubble method in water, or floated in a cylinder with frother only. In this way the composition of solutions in which a mineral was conditioned but in which properties of the collector solution prevented contact or flotation can be defined. The composition of solutions in which conditioning was possible lie between the full and broken curve, the experimental results being indicated on the figures by a square. The curves show the limits of composition of solution in which flotation in a cell would be possible. Many factors other than those considered in these tests determine recovery or grade of a given mineral in practice.

Jan 1, 1947

By J. A. Rowland, C. W. Funk

THIS investigation is a part of the United States Bureau of Mines work in conserving the Nation's resources. The isothermal sections presented were developed as a guide to a comprehensive investigation of the properties and fabricating characteristics of copper-base alloys containing manganese and tin. These alloys are being investigated as possible replacements for the commercial bronzes containing substantial quantities of tin. Development of the isothermal sections has, therefore, been limited to the area between 0 to 20 pct Mn and 0 to 25 pct Sn. Since Heusler1 reported the ferromagnetic properties of the Cu-Mn-Sn alloys in 1903, the system has been the subject of several investigations. These studies were largely concerned with correlation of magnetic properties and crystal structure.2-6 Vero,' however, established the solidus surface of the system, between 20 pct Mn and 40 pct Sn, by thermal and microscopic methods. The a, ß, and y solid solutions were also observed by Vero and reported to extend into the interior of the ternary system. For the present study, the annotated diagram of the Cu-Sn system prepared by Raynor' and the Cu-Mn diagram by Dean and coworkers9 were used as the binary borders of the ternary system. Procedure The metals used in this investigation were electrolytic manganese, washed cathode copper, and three-star tin. The manganese, produced at the Boulder City pilot plant of the Bureau of Mines, had a purity of 99.9 pct; the copper contained less than 0.004 pct Bi and 0.001 pct S; the tin contained 0.070 pct Pb, 0.070 pct Sb, 0.050 pct Fe, and 0.065 pct Si. Heats of 1 lb were melted in alundum crucibles with a 3-kw induction furnace. The heats were chill-cast in iron or copper molds to form 6-in. ingots 3/4 in. in diameter. Molds were washed with graphite or zirconia, depending on the manganese content, and preheated to 150°C. The maximum impurity content of any alloy was 0.03 pct Fe, 0.02 pct Si, and 0.02 pct Al; in most cases, they contained less than half this quantity of any of these impurities. Composition of the alloys and their treatment prior to homogenizing are shown in Fig. 1. Designations adjacent to each composition refer to the heat numbers. With a few exceptions, ingots containing 20 pct Sn or less were hot-swaged. These ingots were swaged, after a 24-hr preheat at 650°C, to obtain a total reduction of 70 pct in cross section. Intermittent reheating was necessary, and reductions were limited to either 0.025 or 0.050 in. in diameter per pass. Selection of a 200-hr homogenizing treatment was based on extensive experiments which indicated that virtual equilibrium was reached after 150 hr. Structures were retained by quenching in water. Grain sizes in the specimens homogenized at 350" and 450°C were generally small. These homogenizing treatments were preceded by a 50-hr anneal at 650°C to facilitate phase identification by increasing the grain size in these specimens. The specimens were furnace-cooled from this temperature to the homogenizing temperature, and the heat treatment was continued for the standard 200-hr period. The homogenized samples were bisected to provide an internal section for metallographic examination. Optimum definition of the microstructure was obtained by successive polishing and etching. Quarter-strength ASTM copper reagent No. 13 gave the most satisfactory results.10 Filings from interior sections of the samples were used for diffraction studies. These filings were annealed in evacuated glass tubes at homogenizing temperatures for periods exceeding 50 hr and quenched in water. During this treatment the manganese content was reduced by significant amounts, probably by vaporization. This change in composition was considered in applying the diffraction data, which were used only as a means of identifying phases where the data were consistent with the metallographic evidence. Phase Identification The a solid solution is readily distinguished by the copper-colored, twinned, polyhedral grains, as shown in Fig. 2, and a face-centered cubic X-ray pattern. Tukon indentations, using a 25-g load, indicate a Knoop hardness of 155 for a quenched from 650°C. Using white light developed by a Wratten 78 A filter, the ß grains in etched specimens have a dark brown tint in contrast to the lighter copper-colored grains of the a phase. Knoop hardness values of 284 were obtained for the ß structure in specimens quenched from 650°C. The ß structure was also distinguished from the a phase by its body-centered cubic lattice. The acicular structure appearing in Fig. 3 was evident in several alloys of higher tin content after quenching from either 750' or 650°C. Since a trans-

Jan 1, 1954

By William P. Shinn

Jan 1, 1880

By B. W. Frazier

Jan 1, 1882

By Herbert J. French

ALLOY steels have become essential to industry in meeting the rigid requirements on materials imposed by our, advanced technology. In comparison with the total ingot capacity of the steel industry, the volume of alloy steels is relatively small although for many years annual alloy-ingot tonnage has, been in seven figures. However, the monetary value of the alloy-steel production is exceedingly high and the benefits derived from the products are relatively great. The trend in total ingot production since 1932 has been generally upward with only two major interruptions. That of the year 1938 was due to a decrease in general business activity whereas that of 1946 resulted largely from work stoppages. Traditionally, the tonnage of alloy-steel ingots produced was in the neighborhood of 5 to

Jan 1, 1948

By Messervey, J. P.

NOVA SCOTIA is sharing in the rapid advance of the mining industry that is one of the remark- able features of Canada's recent progress. The production of coal and gypsum has increased rapidly, and there has also been a marked development in the mining of gold, salt, zinc, copper and other minerals. Coal, the basis of the principal mining industry in the Province, was discovered by the early explorers on the shore of Cape Breton Island, and used by the garrison at Halifax as early as 1749. The first government record shows total annual sales amounting to 1668 long tons in the year 1785, while last year production was 6,643,000 long tons. The principal mines are discussed in a separate paper by Alexander L. Hay1.

Jan 1, 1928

By William P. Blake

The Rainbow Lode is situated at Walkerville, in the Summit Valley mining district, Silver Bow County, Montana Territory, about one mile from Butte City. It was so named ill 1876 by Mr. J. E. Clayton, in recognition of the broadly-sweeping curve of the outcrop across the rounded surface of the eastern slope of a granitic mountain, at an elevation of about 6000 feet above the sea-level. It is on the western side of the range which divides the waters of the Missouri from those of the Columbia. History.—The history of the district is an instructive one. The water-courses and flanks of the hills were worked for years for placer gold. Millions of dollars' worth of low-grade gold have been washed out of the surface gravels below the veins of the district. Missoula Gulch and Silver Bow creek, just below Butte, are noted for their rich placers, and they continue to yield largely, producing within narrow limits from $50,000 to $100,000 dollars' worth of golddust yearly. It is ragged vein-gold, and evidently is near its parent source. It is no doubt derived from the disintegration of the veins of the district, and this conclusion was soon forced upon the miners. The veins which were seen outcropping boldly upon the slopes above them were located, and prospecting commenced. These outcrops were found heavily stained and blackened with oxide of manganese, and were porous and honey-combed in places, showing that the quartz was formerly filled with pyritous ores, and that these ores had decomposed and dissolved away, leaving, however, portions of the precious metal behind, especially at and near the surface of the ground. Assays of the croppings gave variable and often contradictory results; but the fact. that these were the outcrops of lodes bearing gold and silver in depth was patent to any experienced observer of mineral veins. The lodes of the district were first brought prominently to notice by Mr. Clayton in 1876, and in October of that year the Alice mine, upon the central portion of the lode, was opened by Mr. Marcus Daly, now so well known as the superintendent of the Anaconda Copper mine, upon another extremely interesting lode, in the same region.

Jan 1, 1888