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By James R. Finlay

SOMETIMES the American Institute of Mining and Metallurgical Engineers appears to be a strictly technical society, and if so my paper should deal with the technical operations of finding and producing gold. But I have not been engaged in practical gold mining since 1910. What I have been doing mainly for the last 25 years is to try to figure out what the income of a mine-gold, or any other-would be worth to an investor. So I have been in contact with gold mining, in one way or another, for some 42 years. I say that with a purpose. In that time the industry has been through one of its cycles. The value of gold mines is very dif¬ferent at different points in one of these cycles. For this reason the cycle of gold mining is a highly practical study. The cycle of gold is a more or less rhythmical change in its value or purchasing power measured in the grand average of other commodities, which is the same thing as its price. Taking the price of gold in 1897 as a stand¬ard, a dollar at that time declined to 55c. in 1913, to 26.8c. in 1920, and since then it has risen to 47.4c. in 1930 and to about 64c. in 1932. It is worth about 21/ times as much now as it was in 1920, but less than two¬thirds as much as in 1897. When this change is applied to the value of a gold mine still greater differences are evident. The cost of producing gold is of course measured in labor and other commodities, and as that cost goes down the margin of profit measured in the gold itself increases. Thus a mine which might make only 10 per cent profit in 1920 on its gross output may easily make 25 per cent now. The stockholder gets 2 ½ times more gold that is worth 2 1/2 times as much in purchasing power, therefore the real value of his holding has risen by the one factor multiplied by the other, giving him 61/4 times as much goods or services as he had in 1920.

Jan 1, 1933

Discussion of the paper of FRANCIS S. PEABODY, presented at the St. Louis meeting, October, 1917, and printed in Bulletin No. 125, May, 1917, pp. 775 to 781. THE CHAIRMAN (CARL SCHOLZ, Chicago, I11.).-Coal wastage is undoubtedly one of the most prominent topics before the coal operators of the United States today. Even those who have been in the business only 25 years, like myself, know that the exhaustion of coal is something that should and must, be curbed. If it is not, before long we shall discover that the great coal resources of the country are not what we-thought they were; we have been so accustomed to work at the best and thickest coal alone, that it will be difficult for us to realize one of these clays-and that (lay is not far distant-that we are no longer the rich nation in coal that we thought we. were. In this connection I feel obliged to attach a little blame to the Geological Survey for misleading the people. There are several members ,here who assisted in that work, and I am going to ask them to defend themselves. The Geological Survey deals with coal seams of much less thickness than we could now expect to, mine, particularly with the present high wages, excepting at a prohibitive cost. The Geological Survey, in its estimates, does not discriminate between thick and thin seams, but simply estimates our resources at so many hundred thousand billion tons of coal. I believe they include seams from 15 to 18 in thick, but you all know that a vein below 3.5 or 4 ft. is not considered economical to mine at this time. R. V. NORRIS, Wilkes-Barre, Pa.-Mr. Peabody has evidently been thinking almost entirely of Illinois conditions, with which I regret to say I am absolutely unfamiliar. In some West Virginia mines the total recovery at present is considerably over 90 per cent. and extraction-, as high as 97 or 98 per cent. are claimed.

Jan 1, 1918

By H. Rush Spedden, A. M. Gaudin

IN the belief that a critical study of its operating problems might be a sound investment, the Cuban American Manganese Corporation initiated an ore-treatment research in cooperation with the Massachusetts Institute of Technology. One of the early objectives of this research naturally became the relationship of the microscopy of the Cuban ores of the Corporation to their flotation treatment. The operations in Cuba have already been described by F. S. Norcross, Jr.7 and the laboratory study of manganese oxide flotation goes back a number of years .4 Broadly speaking, the difficulties encountered in the flotation of manganese ores are that the consumption of reagents is extremely high, the concentrate contains more silica than is desirable and the tailing losses are relatively high. An appreciable betterment on any of these scores would achieve a definite betterment in costs, quantity of minable ores, and quality of product. The microscope study discussed in this paper was made on samples from the Quinto mine, from the Ponupo mine, and from mill products at various stages of passage of the ore through the mill. The writers have had access to Company records and in particular to the petro¬graphic studies made on thin sections of ore by the late Professors R. J. Colony and Philip Krieger, of Columbia University. They acknowledge their debt to both of these men for the help derived from the reports. Manganese oxides are black, friable, smudging. It is, therefore, very difficult to judge operations by inspection. The mineralogy of manganese compounds seems to be confused, and the authors believe that many more minerals have been described than actually exist. Much of this difficulty is due to the prevalence of the structural conditions that will be discussed further in this paper; namely, the porosity and the friability of the ores, and the extremely intimate association of minute manganiferous mineral crystals with impurities of various sorts. By all odds the most important manganese mineral in Cuba is the one that in this paper is called "psilomelane." The name "psilomelane" has been used in so many different ways that the choice of it for this ore may not be approved by all. However, it seems to be the best under the circumstances. This mineral accounts for about 97 per cent of the manganese in the ores of the Quinto and Ponupo mines and in the mill feed. The remaining odd 3 per cent consists mostly of pyrolusite together with minor amounts of other manganese oxide minerals. DESCRIPTION OF PSILOMELANE Aggregate Structure In grains or lumps, psilomelane has a characteristic fibrous appearance and seems

Jan 1, 1942

By W. H. Blauvelt

The only way in which the difficult problems of classification of coal for the manufacture of by-product coke can be solved is to analyze them by the use of scientific data. It is very easy to adopt classifications for coal for the manufacture of by-product coke along certain lines; we can easily say that for a certain class of metallurgical operation coke made from bituminous coal must not contain more than so much sulfur or more than so much ash and should yield certain returns in the by-products, tar, ammonia, benzol, gas, etc. Those points we are fairly clear on and I think it is hardly worth while to discuss them because I presume that they are familiar to most of us who have any occasion either to make coke or to use it. But there is a field in the manufacture of coke which we know very little about. For example, we are operating a plant on a certain mixture of coal. Suddenly our blast-furnace man or the user of the coke tells us the coke is not as good as it was before. The coke is analyzed and the coal is analyzed. Everything seems to be about the same as it was— the ash is the same, there is no more sulfur, and so on—but the coke, for the use to which it is put, particularly in the blast furnace, is wrong. What happens? The management decide to try a slightly different coal mixture. They do not necessarily raise the total hydrocarbons, the volatile matter in the mixture, nor change the sulfur or the ash, but they do make a change which improves the value of the coke for metallurgical use. The reason for this difficulty, as I see it, is the " combustibility " of the coke—I hesitate to use the word, but it is the best one I have available. Combustibility of Coke We know now that the combustibility of coke is a very vital point in many metallurgical processes and I think one of the most notable instances of the proof of this is the experience of the Carnegie Steel Co. in the production of coke at its Clairton plant. The history of that

Jan 1, 1930

By Milton T. Pana

Computer techniques now have been applied to the solution of a wide variety of scientific and engineering problems in the mining industry, but generally not in great depth in any one area. In mining, the area covered in greatest depth and with the greatest success has probably been mine planning and to a lesser degree production scheduling. A rather comprehensive review of the literature related to mine planning and production scheduling indicates that contributions in these areas have been geographically, as well as organizationally, widespread. Of twenty-six papers reviewed, eleven were from industry, eight from educational institutions, four from computer vendors, two from the Bureau of Mines, and one from a private consultant. The conclusions that can be derived from the literature are: First, the-state-of-the art has advanced to the point that computing ore reserves and developing mining plans with the aid of the computer is now a routine practice; second, the application of the computer to mine planning problems has markedly improved our ability to estimate ore reserves, more accurately define pit limits, and determine an optimum mining sequence. MINE PLANNING The papers in this section deal essentially with the problem of determining the final pit limits of an orebody and the ore reserves associated with a given pit limit. The problem is probably best described by Soderberg25 and more recently by Reibell. 23. There are two techniques currently being applied to this problem by several mining companies: 1) the multiple-cone technique and 2) the Lerch-Grossman algorithm. The multiple-cone technique is the older of the two techniques and has the advantage of being more flexible, allowing an engineer to easily vary pit slopes in different sectors of the design. The technique can also be used to define intermediate pit contours. The Lerch-Grossman algorithm is appealing in that the final solution is given in the form of a true mathematical optimum. The technique is limited, however, in that the size of the blocks in the orebody model are a function of the desired

Jan 1, 1969

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core.

Jan 7, 1951

By J. Armanet

THREE MINING colleges are maintained in France; the École Nationale Supérieure des Mines, of Paris; École Nationale Supérieure des Mines, of Saint Etienne; and the École Supérieure de la Metallurgie et des Mines at Nancy. Besides graduates of these schools, a few graduates of the École Centrale des Arts et MIanufactures and of the École Polytechnique are employed as ]nine managers or engineers; however, most of the latter have taken a graduate course at one of the mining schools.

Jan 1, 1939

By Bjorge, Guy N.

MINING method improve through the gradual process of evolution and in 1340 there were no marked outstanding innovations. On the other hand refinements of detail and betterment: in equipment design continued to progress. Among the most important changes have been the increased application of trucks and highway equipment, including tractors and bull- dozers, to open-pit mining as exemplified by Phelps Dodge's Morenci operations; the Mesabi iron range; and the Tri-State district. Mass or bulk handling contributed its part underground with slurhers and conveyors at Consolidated Coppermines and slushers at Climax; and in open-pit mining. conveyors found increased use on the Mesabi range.

Jan 1, 1941

By CHARLES R. KETES

I. INTRODUCTORY. INDUSTRIALLY, the most important service that geological science can now render to mining in the Upper Mississippi leadand zinc-fields is to devise some practical scheme whereby the productive ore-belts may be located with greater certainty than they have been in the past. Up to the present time there has been little service of this kind, and in spite of all that has been worked out concerning the geologic features of the region, and notwithstanding all that has been written about these ore-deposits, this particular phase of the subject has been, for some reason or other, the very one that has received the least consideration. The neglect by geologists in this respect has been all but complete. The following notes are suggestive of some practical solutions to the problems presented. As here suggested, the geologic relationships of the ore-bodies were surmised some years ago, though somewhat vaguely perhaps. No opportunity for securing critical data appeared until recent investigations of a private character were undertaken, which bad in view the extensive prospecting of undeveloped areas. Fresh from rather wide experiences of a decade in the principal mining-regions of four continents, and after a lenstrum's residence among the chief mining-camps of the West and of Mexico, under conditions which measured geologic hypotheses in meters and rigidly tested their values in, terms of dollars, the recent return to the Joplin field and a new survey of its ore-conditions were fraught with somewhat unexpected results. Aspects of the ore-deposition were presented of which previously there was no hint. Ten years' absence from the district had tended greatly to clarify former hazy impressions. It enabled many features, then only dimly outlined, to be considered more nearly in their true perspective and in their real relationships to their surroundings. It permitted a great mass of details collected a dozen years before in connection with the work of the Missouri Geological Survey to be viewed under new light and to

Feb 1, 1909

By J. J. Forbes

Probably the newest thing in mining safety, or safety for mines, is the apparent dissatisfaction on the part of the mineral industries, as represented by both management and labor, and the general public of the Nation, with the safety record of the mining industry. This is reflected by published comments and papers in the technical press;1-8 in labor periodicals, 10 11 in popular journals and periodicals,12 13 and in the daily press; also by discussions of the Safety Committees sponsored by the National Safety Council, 9, 14-15 the American Mining Congress, 1, 16 the American Institute of Mining and Metallurgical Engineers, 2, 4 the Mine Inspectors' Institute of America,6,8,17,18 the Coal Mining Institute of America, 5, 19, the American Standards Association, and others.20,21

Jan 1, 1949

By G. E. Warner

This paper presents a review and analysis of a highly stratified Burbank sand waterflooding project in Osage County, Okla. Permeability values in this reservoir range from less than 0.1 md to nearly 3 darcys, and 90 percent of the permeability capacity is contained in approximately 26 percent of the reservoir volume. To date, the project has recovered 103 bbl/flood-acre-ft since waterflood initiation, which represents 50 percent of the total preflood cumulative and 12 percent of the original stock-tank oil in place (OSTOIP). The outstanding feature of this project has been the economical recovery of a sizable quantity of oil at water-oil ratios higher than those normally considered prohibitive. Nearly 70 percent of the recovery to date has been achieved at water-cut values greater than 95 percent. An ultimate waterflood recovery of 138 bbl/flood-acre-ft, or 67 percent of the preflood recovery and 16 percent of the OSTOIP, is predicted. The ultimate primary and secondary recovery will be 302 bbl/acre-ft and approximutely 40 percent of the OSTOIP. The actual performance is compared with the predicted performance by both the Stiles and the Dykstra-Parsons methods, and applicabiliry of these methods to this particular reservoir is evaluated. The performance of this project indicafes highly stratified reservoirs can be successfully and economically waterflooded when either reservoir or economic conditions preclude the use of mechanical and chemical means of mdbility control. Introduction The concept of the stratified reservoir has been used by reservoir engineers for many years to describe the vertical variations in permeability within a reservoir. Numerous techniques have been advanced to predict the expected waterflood performance of the stratified reservoir, and in most cases they prove reasonably accurate. The advisability of undertaking a prospective waterflooding project is determined, based on these predictions. This paper presents a review of a waterflood project in a highly stratified reservoir that normally would be considered a marginal prospect, at best, when evaluated with any prediction technique properly adjusted for reservoir conditions. This kind of project can be successful if the predictions are used to design a waterflooding system capable of economically recovering an oil cut of 1 to 2 percent. Experience gained in the early stages of the project's development was profitably applied in the later development of the entire reservoir and could prove beneficial to other projects with similar reservoir stratification or fracture characteristics. The East Burbank pool waterflooding operations have been developed in three separate stages, each of which has been a separate project: Mid-Burbank Unit, Flood A, and Stanley Waterflood. When referred to collectively, these projects will be called the Composite Project. The Skelly Oil Co.'s Barber lease — the only portion of the pool not included in the three projects — is operated as a separate facility with a cooperative injection-well agreement. It contains less than 80 productive acres and would not significantly affect the over-all performance picture.

Jan 1, 1969

By D. J. Jr. McAdam

PART I.-OUTLINE OF INVESTIGATION, DESCRIPTION OF METHODS AND MATERIAL Previous Investigation of the Influence of Stress on Corrosion IN 1917 Haigh1 presented evidence that under simultaneous corrosion and cyclic stress metals may fail at lower stresses than if the corrosion is prior to the cyclic stress. In 1926 the author, while investigating the cooling effect of water on the fatigue resistance of metals, found that under these conditions steel specimens fail at stresses far below the endurance limit.2 (By "endurance limit" is here meant the fatigue limit obtained with specimens as free as possible from stress concentration and corrosive influences.) This led to an investigation (at the Naval Engineering Experiment Station, Annapolis, Md.) of the influence of cyclic stress on corrosion. This investigation maybe divided into three parts. The first part was a general survey of the conditions necessary to cause complete failure under corrosion and cyclic stress of relatively high frequency. The entire process by which metals fail under these conditions was-called "corrosion-fatigue." "Corrosion-fatigue limits" for a large number of metals were determined and the relationships of these limits to chemical composition, heat treatment and physical properties were established. The conclusion was reached that corrosion-fatigue is due to the accelerating influence of cyclic stress on corrosion. Results of this part of the investigation were presented in five papers.2-6 The second part of the investigation was a preliminary study of the influence of cyclic stress on corrosion. Each experiment was conducted in two stages. The first was a corrosion stage in which the specimen was corroded under cyclic stress; the second was a fatigue stage, in which the previously corroded specimen was tested (without further corrosion) to determine its fatigue limit. The difference between this fatigue limit and the endurance limit was used as a criterion of the "damage" due to

Jan 1, 1931

By Paul M. Tyler

Higher freight rates and better methods of beneficiation now may make it more economical to open inferior deposits closer to a glass factory than to work higher-grade deposits farther away. Nature of impurities and special treatments as well as common practice of sulphuric acid leaching are described. HISTORICAL concepts of the economics of the glass-sand industry are changing rapidly. The greatly expanded demand for glass containers combined with higher freight rates on raw materials and manufactured products have induced a migration of glass factories toward densely populated centers and the creation of new standards of place value for substandard sand deposits. This migration has been facilitated by the construction of pipelines to bring cheaper natural gas and liquid fuel to large cities and may be further speeded by the adoption of modern mineral dressing methods to permit economical utilization of local raw materials. Our national resources of naturally high-grade silica sand are abundant, but most of the best deposits are situated so far from large centers of population that it now costs $3.00 to $5.00 a ton to carry the sand from the mine to the optimum site for a factory making beverage bottles, scientific and electrical glassware, and miscellaneous blown and pressed ware for local markets. It follows that on the basis of freight savings alone it may be economical to use material from inferior deposits closer to the factory even though the f.O.b. cost of mining and treatment may be much higher than at a higher grade deposit farther away. Particle-size distribution is a feature of glass-sand specifications. As long as glassmakers insisted upon having their sand principally coarser than 100-mesh, any material was automatically excluded from consideration that could not be purified without fine grinding or other drastic treatment which greatly reduces the size of the quartz grains. However, re- cent large-scale tests in Norway' showed that finely-crushed quartz used in a soda-lime-silica glass batch caused it to melt and become refined appreciably faster than when ordinary coarse (Belgian) sand was employed. Dust losses were negligible. Since the pre-war cost of sand was only 3 pct of the manufacturing cost of the glass, it was deemed economical to pay twice as much for crushed quartz because of the saving in wear and tear on the furnace walls. In the author's unverified opinion, sand ranging in size between 150 and 400-mesh would probably be superior to that in the more usual range of 20 to 100-mesh. Former objections to fine sand (other than dust) were doubtless fostered by the natural concentration of impurities in the finer sizes. Ordinarily the most objectionable, as well as the commonest, impurity from the standpoint of the manufacturer of any but the cheaper qualities of colored glass is iron. No natural sands are really pure silica. Even water-clear quartz crystals are likely to contain impurities in the form of solid solution as well as inclusions which cannot be eliminated except by chemical treatment that will break down the silica lattice. Pure white pegmatite quartz will usually analyze 0.01 pct or more of iron oxide, and ordinary vein quartz and the quartz grains in most

Jan 1, 1951

By Paul M. Tyler

Higher freight rates and better methods of beneficiation now may make it more economical to open inferior deposits closer to a glass factory than to work higher-grade deposits farther away. Nature of impurities and special treatments as well as common practice of sulphuric acid leaching are described. HISTORICAL concepts of the economics of the glass-sand industry are changing rapidly. The greatly expanded demand for glass containers combined with higher freight rates on raw materials and manufactured products have induced a migration of glass factories toward densely populated centers and the creation of new standards of place value for substandard sand deposits. This migration has been facilitated by the construction of pipelines to bring cheaper natural gas and liquid fuel to large cities and may be further speeded by the adoption of modern mineral dressing methods to permit economical utilization of local raw materials. Our national resources of naturally high-grade silica sand are abundant, but most of the best deposits are situated so far from large centers of population that it now costs $3.00 to $5.00 a ton to carry the sand from the mine to the optimum site for a factory making beverage bottles, scientific and electrical glassware, and miscellaneous blown and pressed ware for local markets. It follows that on the basis of freight savings alone it may be economical to use material from inferior deposits closer to the factory even though the f.O.b. cost of mining and treatment may be much higher than at a higher grade deposit farther away. Particle-size distribution is a feature of glass-sand specifications. As long as glassmakers insisted upon having their sand principally coarser than 100-mesh, any material was automatically excluded from consideration that could not be purified without fine grinding or other drastic treatment which greatly reduces the size of the quartz grains. However, re- cent large-scale tests in Norway' showed that finely-crushed quartz used in a soda-lime-silica glass batch caused it to melt and become refined appreciably faster than when ordinary coarse (Belgian) sand was employed. Dust losses were negligible. Since the pre-war cost of sand was only 3 pct of the manufacturing cost of the glass, it was deemed economical to pay twice as much for crushed quartz because of the saving in wear and tear on the furnace walls. In the author's unverified opinion, sand ranging in size between 150 and 400-mesh would probably be superior to that in the more usual range of 20 to 100-mesh. Former objections to fine sand (other than dust) were doubtless fostered by the natural concentration of impurities in the finer sizes. Ordinarily the most objectionable, as well as the commonest, impurity from the standpoint of the manufacturer of any but the cheaper qualities of colored glass is iron. No natural sands are really pure silica. Even water-clear quartz crystals are likely to contain impurities in the form of solid solution as well as inclusions which cannot be eliminated except by chemical treatment that will break down the silica lattice. Pure white pegmatite quartz will usually analyze 0.01 pct or more of iron oxide, and ordinary vein quartz and the quartz grains in most

Jan 1, 1951

By Frederick Laist

THE object of the construction and operation of the 80-ton leaching plant was to test out the leaching of sand tailings on a large scale and, if possible, determine a definite method of operation, and the best construction for the larger unit which is now being built. It was also of importance to obtain some practical data as regards cost items. The 80-ton plant was not expected to be a commercial success. A leaching process on such low-grade material necessitates the treatment of a very large tonnage.. Such a cost item as labor, for instance, is entirely out of proportion in such a small plant. The amounts of acid, salt, scrap iron, and fuel for roasting, however, are the same per ton in an 80-ton plant as in a 2,000-ton plant. These costs were definitely established. Also the percentages of extraction and grade of tailings were definitely determined for a large-sized unit. Following will be found the description of the plant, the method of operation and the results obtained. A -general view of the plant is given in Fig. 1. The bins consisted of one coal bin with a capacity of 30 tons, one salt bin with a capacity of 20 tons, and a "mill tailings" or feed bin which held 70 tons. The coal and salt bins were placed with their floor level even with the firing floor of the roasting furnace. The feed bin, to give it greater capacity, discharged about 12 ft. below this point, to a 12-in. conveyor belt, which, in turn discharged into a pug mill. This fed another conveyor belt, which discharged into the top hearth of the furnace. The furnace was an ordinary six-hearth MacDougall roaster, .20-ft. in diameter. It was equipped with two fire boxes, on opposite sides, the flames entering on the third hearth. An induced draft was obtained with a Buffalo blower, worked as an exhaust fan. Just above the top hearth was an annular flue around the whole circumference of the furnace. Slots at intervals of about 18 in. led from this flue down into the top hearth. The slots were 2 in. wide by 6 in. long. This annular flue proved rather unsatisfactory, as it tended to fill up with flue dust and choke the slots, thus cutting down the draft. The volume of gas through

Jan 8, 1914

By Robert Glass Cleland

DURING the closing month of 1929, Walter Douglas found his health impaired by the strain of many difficult years of alternating prosperity and depression, and in April 1930 resigned the presidency of Phelps Dodge. Douglas was succeeded by Louis S. Cates. The choice broke a tradition that had been held inviolate for almost a hundred years. For the first time in its long history, the direction of the company's policies and operations was entrusted to a stranger without previous Phelps Dodge background, training, experience, or tradition. But the choice was wisely made. For a time at least, the path of Phelps Dodge was to run through rough and dubious ground, and no man in the United States was better qualified to blaze that trail than Louis Cates. After graduating from the Massachusetts Institute of Technology, Cates had served a rough and financially un- profitable apprenticeship at one of the century-old mining camps of Chihuahua, Mexico. Following his return to the United States and a brief interlude as assistant to the president of the National Steel and Wire Company, he had resumed his mining career, first as timekeeper, then as

Jan 1, 1952

By Frederick Laist

This paper is a review of Montana reverberatory smelting practice covering a period of approximately fifty years, during which time the small furnaces that had been in use elsewhere for a century or more, without much change in size or method of operation, were revolutionized and were developed into the huge, continuously operating furnaces of today. A study of Montana practice is particularly instructive, since reverberatory furnace smelting for the production of copper matte received more attention there than in other smelting centers and many of the improvements that finally resulted in establishing the preeminence of this type of furnace over the blast furnace originated in Butte and Anaconda. The material in this paper was gathered from original sources, wherever possible, and all of the illustrations were made from original drawings, some of which have been simplified and worked over so as to be better adapted for reproduction. The aim has been to trace the modifications in design and operation which took place during the period in question—the reasons for the changes and the results achieved through them. For developments during the first twenty years, the technical literature of the period was freely drawn upon, but I am particularly indebted to W. M. Kelly, general foreman at the Anaconda Works, whose recollections and personal experience extend back to the early days. I wish to acknowledge the use made of publications of E. P. Mathewson, William Wraith and L. V. Bender, as well as information obtained here and there from friends and associates. The resume of Great Falls practice was compiled from correspondence, notes and drawings found in the files of the old Boston & Montana Co. and the tabulations showing the performance of furnaces at Butte and Anaconda are adapted from plant records and correspondence in the files of the Anaconda Copper Mining Company. The paper deals with some interesting comparisons that were made between furnaces operated under various conditions in connection with the Carson Case, and results obtained when using natural gas as fuel. This part of the paper may be considered original to the extent that these results have not previously been published.

Jan 1, 1934

By E. A. Anderson, P. W. Ramsey, E. J. Boyle

Only a few zinc-rich alloys are in commercial use today, and these depend largely upon solid solution effects for their desirable properties. Typical examples are rolled alloys containing about 1 Per cent of copper, with Or without additions of elements such as magnesium, and die-casting alloys containing about 4 Per cent of aluminum with small amounts of magnesium and sometimes copper." A number of elements are capable of forming alloys with zinc, Which do not enter into solid solution except in limited degree. The Present Paper considers One such system, zinc-titanium, in which alloys having potentially useful properties in the rolled form were found. Zinc-titanium System The zinc-rich portion of the zinc-titanium diagram, as determined by standard methods in this laboratory,3 is shown in Fig. I. The portion of the diagram of significance in the present work extends from 0 to about 0.5 per cent titanium. Gebhardt4 places the eutectic at 0.45 per cent, but the evidence in favor 01 the composition given in Fig. I is considerable. The solid solubility of titanium in zinc is very low indeed and possibly places these alloys in a somewhat different category from the commercial alloys previously mentioned. Alloy Preparation All of the alloys used in this work were prepared with titanium metal of 98.9 per cent purity. In most cases high-purity zinc (99.99 per cent) was used but sufficient work was done later to show that zinc of about high-grade composition would yield approximately equal results. Best practice in alloy manufacture seemed to be to prepare a zinc-titanium hardener by adding about 4 per cent of titanium metal to molten zinc and holding under a boric acid cover, with frequent stirrings, at 750°C, until the titanium had dissolved. The desired titanium content could readily be obtained in the final alloy by adding the calculated amount 01 this hardener. The alloys proved to be quite stable in composition, unless overheated, and, with reasonable temperature control, required no cover or flux to prevent loss of titanium content by oxidatioll, Hot-rollig Tests Three alloys containing, by analysis, 0.05, 0.12, and 0.23 per cent titanium were selected for preliminary study. Slab cross sections a a marked grain refinement* (Fig. 2) in the two alloys of highest titanium content. Cast unalloyed zinc has a columnar grain structure similar to that shown for the alloy of lowest titanium content but without the scatter of small equiaxed grains at the center.

Jan 1, 1944

By P. W. Ramsey, E. J. Boyle, E. A. Anderson

Only a few zinc-rich alloys are in commercial use today, and these depend largely upon solid solution effects for their desirable properties. Typical examples are rolled alloys containing about 1 Per cent of copper, with Or without additions of elements such as magnesium, and die-casting alloys containing about 4 Per cent of aluminum with small amounts of magnesium and sometimes copper." A number of elements are capable of forming alloys with zinc, Which do not enter into solid solution except in limited degree. The Present Paper considers One such system, zinc-titanium, in which alloys having potentially useful properties in the rolled form were found. Zinc-titanium System The zinc-rich portion of the zinc-titanium diagram, as determined by standard methods in this laboratory,3 is shown in Fig. I. The portion of the diagram of significance in the present work extends from 0 to about 0.5 per cent titanium. Gebhardt4 places the eutectic at 0.45 per cent, but the evidence in favor 01 the composition given in Fig. I is considerable. The solid solubility of titanium in zinc is very low indeed and possibly places these alloys in a somewhat different category from the commercial alloys previously mentioned. Alloy Preparation All of the alloys used in this work were prepared with titanium metal of 98.9 per cent purity. In most cases high-purity zinc (99.99 per cent) was used but sufficient work was done later to show that zinc of about high-grade composition would yield approximately equal results. Best practice in alloy manufacture seemed to be to prepare a zinc-titanium hardener by adding about 4 per cent of titanium metal to molten zinc and holding under a boric acid cover, with frequent stirrings, at 750°C, until the titanium had dissolved. The desired titanium content could readily be obtained in the final alloy by adding the calculated amount 01 this hardener. The alloys proved to be quite stable in composition, unless overheated, and, with reasonable temperature control, required no cover or flux to prevent loss of titanium content by oxidatioll, Hot-rollig Tests Three alloys containing, by analysis, 0.05, 0.12, and 0.23 per cent titanium were selected for preliminary study. Slab cross sections a a marked grain refinement* (Fig. 2) in the two alloys of highest titanium content. Cast unalloyed zinc has a columnar grain structure similar to that shown for the alloy of lowest titanium content but without the scatter of small equiaxed grains at the center.

Jan 1, 1944

By Paul F. Kerr, Donald M. Hausen

Fine colloidal gold near Carlin, Nevada is disseminated in leached carbonate strata of the Roberts Mountains Formation in the Lynn "window" of the Roberts Mountains thrust fault. The ore body is generally stratiform and is more or less conformable to altered beds near the top of the formation, underlying Devonian limestones. Two sequences of mineralization are recognized: ( 1) an earlier base metal-barite assemblage related to early Cretaceous intrusives (121 ± 5 m.y.), and (2) a later low temperature Au-As-Hg-Sb assemblage of near surface emplacement. The earlier sequence consisting of sparse galena and sphalerite in barite with anomalous amounts of zinc, lead, nickel and copper, associated with dikes of dacitic composition, is of little economic importance. The later sequence of gold, realgar, cinnabar, and stibnite associated with extensive silicification and argillic alteration of limestone beds, has resulted in important deposits of gold. Argillic alteration of hydrothermally leached carbonate strata has provided the environment in which the most prominent gold deposition took place. Carbonate minerals in the limestone host rock have been replaced by microcrystalline quartz and chalcedony to form stratiform silica masses and recrystallized lenses of euhedral quartz. Zones of porous silicification are light gray to white, elliptical in shape, and more or less follow bedding. Silicification is bordered by argillic alteration and pyritization. Deposition of gold usually lies in a zonal pattern that encircles chimney areas of silicification.

Jan 1, 1968